Edited by
Maurice C. Fuerstenau
Graeme Jameson
and
Roe-Hoan Yoon
Published by
Society for Mining,
Metallurgy, and Exploration, Inc.
© 2007 by the Society for Mining, Metallurgy, and Exploration.
All rights reserved. Electronic edition published 2009.
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Society for Mining, Metallurgy, and Exploration, Inc. (SME)
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Copyright © 2007 Society for Mining, Metallurgy, and Exploration, Inc.
Electronic edition published 2009.
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ISBN-13: 978-0-87335-280-2
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flotation0.book Page iii Tuesday, January 2, 2007 7:36 PM
Contents
CONTRIBUTORS
PREFACE
PART 1
v
ix
HISTORICAL ASPECTS OF FLOTATION
1
A Century of Developments in the Chemistry of Flotation
Processing
History of Flotation Technology
PART 2
FLOTATION FUNDAMENTALS
FLOTATION CHEMISTRY
95
133
179
227
259
283
339
373
Flotation Reagents—A Critical Overview from an
Industry Perspective
Sulfide Mineral Flotation
Flotation Chemistry and Technology of Nonsulfide Minerals
Depressants in Nonsulfide Mineral Flotation
Flotation of Precious Metals and Their Minerals
Coal Flotation
PART 4
FLOTATION CELLS, MODELING, AND SIMULATION
FLOTATION PLANT PRACTICE
869
iii
425
465
555
575
611
637
681
739
757
779
Plant Practice: Sulfide Minerals and Precious Metals
Plant Practice: Nonsulfide Minerals
INDEX
375
635
Mechanical Froth Flotation Cells
Column Flotation
Optimal Designs for Homogeneous, Countercurrent
Flotation Processing Networks
Modeling and Simulation of Industrial Flotation Processes
PART 5
65
93
Some Aspects of Flotation Thermodynamics
The Nature of Hydrophobic Attraction Forces
Adsorption of Surfactants and its Influence on the
Hydrodynamics of Flotation
Pulp and Solution Chemistry
The Physics and Chemistry of Frothers
Surface Characterization and New Tools for Research
The Flotation of Fine and Coarse Particles
PART 3
3
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781
845
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Contributors
N.A. Abdel-Khalek
Central Metallurgical Research and
Development Institute (CMRDI)
Cairo, Egypt
S. Chryssoulis
Advanced Mineral Technology Laboratory
London, Ontario, Canada
Jan J.I.R. Cilliers
Department of Earth Science and
Engineering
Royal School of Mines, Imperial College
London, England, United Kingdom
Hans Allenius
Outokumpu Technology Minerals
Espoo, Finland
Armando C. Araujo
Department of Mining Engineering
Federal University of Minas Gerais
Belo Horizonte, Brazil
William Ducker
Particulate Fluids Processing Center
Faculty of Engineering
The University of Melbourne
Victoria, Australia
Barbara J. Arnold
PrepTech Inc.
Apollo, Pennsylvania
Robert C. Dunne
Newmont Australia Ltd.
West Perth, Western Australia
Seher Ata
Center for Multiphase Processes
University of Newcastle
Callaghan, New South Wales, Australia
A. El-Midany
Central Metallurgical Research and
Development Institute (CMRDI)
Cairo, Egypt
Cesar I. Basilio
Thiele Kaolin Company
Sandersville, Georgia
Hassan El-Shall
Center for Particle Science and Technology
University of Florida
Gainesville, Florida
Trevor Bilney
Kanowna Belle Gold Mine
Boulder, Western Australia
W.J. Bruckard
CSIRO Minerals
Clayton South, Victoria, Australia
Jan Christer Eriksson
Department of Chemistry, Surface Chemistry
Royal Institute of Technology
Stockholm, Sweden
Subhash Chander
Department of Energy and
Geo-Environmental Engineering
Pennsylvania State University
University Park, Pennsylvania
K. Fa
Department of Metallurgical Engineering
University of Utah
Salt Lake City, Utah
v
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flotation0.book Page vi Tuesday, January 2, 2007 7:36 PM
Mike Fairweather
M.J. Fairweather & Associates
Rossland, British Columbia, Canada
Stephen Grano
Ian Wark Research Institute
University of South Australia
Mawson Lakes Campus
South Australia
James A. Finch
Department of Metallurgical Engineering
McGill University
Montreal, Quebec, Canada
Michael Habner
Kalgoorlie Consolidated Gold Mine
Kalgoorlie, Western Australia
Daniel Fornasiero
Ian Wark Research Institute
University of South Australia
Mawson Lakes Campus
South Australia
Gregory J. Harbort
Julius Kruttschnitt Mineral Research Centre
Indooroopilly, Queensland, Australia
Martin C. Harris
Department of Chemical Engineering
University of Cape Town
South Africa
Eric K.S. Forssberg
Division of Mineral Technology
Lulea University of Technology
Lulea, Sweden
Thomas W. Healy
Particulate Fluids Processing Center
Faculty of Engineering
The University of Melbourne
Victoria, Australia
Douglas W. Fuerstenau
Department of Materials Science and
Engineering
University of California
Berkeley, California
John A. Herbst
Metso Minerals Optimization Services
Colorado Springs, Colorado
Maurice C. Fuerstenau
Department of Materials Science and
Engineering
University of Nevada
Reno, Nevada
Ronaldo Herrera-Urbina
Chemical Engineering and Metallurgy
University of Sonora
Hermosillo, Sonora, Mexico
A.R. Gerson
Ian Wark Research Institute
University of South Australia
Mawson Lakes Campus
South Australia
G.A. Hope
Faculty of Science and Technology
Griffith University
Nathan, Queensland, Australia
Craig Goodall
Lonmin Platinum
Marikana, South Africa
Graeme J. Jameson
Center for Multiphase Processes
University of Newcastle
Callaghan, New South Wales, Australia
Barun K. Gorain
Corporate R&D/Technical Services
Barrick Gold Corporation
Toronto, Ontario, Canada
N.W. Johnson
College of Engineering
University of Queensland
Brisbane, Queensland, Australia
Brian D. Gotts
Potash Corporation of Saskatchewan
Allan, Saskatchewan, Canada
vi
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Bert Knopjes
Lonmin Platinum
Marikana, South Africa
D.R. Nagaraj
Minerals Processing Chemicals Division
Cytec Industries Inc.
Stamford, Connecticut
Janusz S. Laskowski
Mining and Mineral Processing Engineering
University of British Columbia
Vancouver, British Columbia, Canada
J. Nalaskowski
Department of Metallurgical Engineering
University of Utah
Salt Lake City, Utah
R. Lastra
Mining and Mineral Sciences Laboratories
Natural Resources Canada (CANMET)
Ottawa, Ontario, Canada
Anh V. Nguyen
Center for Multiphase Processes
University of Newcastle
Callaghan, New South Wales, Australia
Gerald H. Luttrell
Mining and Minerals Engineering
Virginia Polytechnic Institute and State
University
Blacksburg, Virginia
Heikke Oravainen
Outokumpu Technology Minerals
Espoo, Finland
Richard Peaker
Metso Minerals
York, Pennsylvania
Alban J. Lynch
Julius Kruttschnitt Mineral Research Centre
University of Queensland
Indooroopilly, Queensland, Australia
Antonio E.C. Peres
Federal University of Minas Gerais
Belo Horizonte, Brazil
Sharad Mathur
Technical Center
Engelhard Corporation
Gordon, Georgia
A.R. Pratt
Mining and Mineral Sciences Laboratories
Natural Resources Canada (CANMET)
Ottawa, Ontario, Canada
R. McEachern
Potash Corporation of Saskatchewan
Allan, Saskatchewan, Canada
Robert J. Pugh
Chemical and Engineering Industries Section
Institute for Surface Chemistry–YKI
Stockholm, Sweden
Thomas P. Meloy
West Virginia University
Morgantown, West Virginia
Srinivasa Raghavan
Department of Materials Science and
Engineering
University of Arizona
Tucson, Arizona
J. Mielczarski
Laboratoire Environment et Mineralurgie
Vandoeuvre-les-Nancy, France
Jan D. Miller
Department of Metallurgical Engineering
University of Utah
Salt Lake City, Utah
John Ralston
Ian Wark Research Institute
University of South Australia
Mawson Lakes Campus
South Australia
vii
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K.H. Rao
Division of Mineral Processing
Lulea University of Technology
Lulea, Sweden
X. Wang
Department of Metallurgical Engineering
University of Utah
Salt Lake City, Utah
S.A. Ravishankar
Minerals Processing Chemicals Division
Cytec Industries Inc.
Stamford, Connecticut
John Watt
Division of Minerals
CSIRO Minerals
Melbourne, Victoria, Australia
Geoff Senior
BHP Billiton Nickel West
Perth, Western Australia
Asa T. Weber
Dorr-Oliver Eimco
Salt Lake City, Utah
W.M. Skinner
Ian Wark Research Institute
University of South Australia
Mawson Lakes Campus
South Australia
Mark C. Williams
West Virginia University
Morgantown, West Virginia
James T. Woodcock
CSIRO Minerals
Clayton South, Victoria, Australia
Robert Snow
Beneficiation and Mining
Florida Institute of Phosphate Research
Bartow, Florida
Ronald Woods
School of Science
Griffith University
Nathan, Queensland, Australia
Ponisseril Somasundaran
Henry Krumb School of Mines
Columbia University
New York, New York
Juan Yianatos
Department of Chemical Engineering
Santa Maria University
Valparaiso, Chile
G.J. Sparrow
CSIRO Minerals
Clayton South, Victoria, Australia
Roe-Hoan Yoon
Center for Advanced Separation Technologies
Virginia Polytechnic Institute and
State University
Blacksburg, Virginia
Roger StC. Smart
Applied Center for Structural and
Synchrotron Studies
University of South Australia
Mawson Lakes Campus
South Australia
Lui Zhang
Akzo Nobel Chemicals Inc.
Dobbs Ferry, New York
G. Strathdee
Potash Corporation of Saskatchewan
Allan, Saskatchewan, Canada
Patrick Zhang
Beneficiation and Mining
Florida Institute of Phosphate Research
Bartow, Florida
Frank P. Traczyk
Dorr-Oliver Eimco
Salt Lake City, Utah
viii
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Preface
The year 2005 marked the 100th anniversary of Sulman and Picard’s U.S. patent award
(No. 793,808) that prescribed the use of air bubbles for flotation. 1905 was also the year
when the Potter process was introduced to flotation in the minerals industry. The production of sphalerite concentrate at Broken Hill in Australia was the first major commercial
application of froth flotation. Following that initial application, froth flotation quickly
spread to the United States and the rest of the world, where it remains an essential separation step in the beneficiation of minerals and coal. Its scope is continually broadening to
other applications such as environmental control, bitumen extraction from tar sands, and
recycling.
Recognizing its significance, a group of flotation researchers and practitioners met in
2001 to consider ways for commemorating this important anniversary. The idea was initiated by a group of individuals, including D.W. Fuerstenau from the University of California,
Berkeley; M.C. Fuerstenau from the University of Nevada, Reno; and Roe-Hoan Yoon from
Virginia Tech. They were joined by D.R. Nagaraj, Cytec Industries; J.A. Herbst, Metso
Minerals; J.-P. Franzidis, Julius Kruttschnitt Mineral Research Centre; J.A. Ralston, Ian
Wark Research Institute, University of South Australia; and G.J. Jameson, University of
Newcastle.
Two international initiatives were launched—a symposium and this commemorative
volume.
Managed by the Australasian Institute of Mining and Metallurgy, the Centenary of Flotation Symposium was held in June 2005 in Brisbane, Australia. It was a great success,
attracting more than 450 delegates and 149 presentations from around the world. The conference fostered in-depth discussion of recent research and up-to-date descriptions of
advanced plant practice. A CD of the conference proceedings is included with this volume.
This commemorative volume, published by Society for Mining, Metallurgy, and Exploration, is a comprehensive resource detailing the state of the art of flotation. The book is the
continuation of a distinguished series published by SME. The sequence began with Froth
Flotation: 50th Anniversary Volume (1962), edited by D.W. Fuerstenau, to celebrate the first
50 years of flotation in the United States; followed by the A.M. Gaudin Memorial Volume
(1976), edited by M.C. Fuerstenau. The continuing involvement of the Fuerstenau brothers
in these important volumes over such a long time span is particularly noteworthy.
The chapters in the book are written by experts in the various disciplines and cover all
aspects of flotation, from fundamental research to industrial practice. Coverage includes the
historical aspects of flotation; flotation fundamentals; flotation chemistry; flotation cells,
modeling, and simulation; and flotation plant practice. The book is an invaluable reference
for industry practitioners, researchers, and graduate students.
Sincere appreciation is extended to all who have contributed to the various chapters.
Despite its longevity, the field of flotation is quite active and rapidly changing. The editors
and SME are fortunate to have contributions from so many leaders in the industry for this
milestone project.
ix
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PART 1
Historical Aspects of Flotation
1
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A Century of Developments in the
Chemistry of Flotation Processing
Douglas W. Fuerstenau
A B S T R AC T
This chapter reviews some aspects of the significance of flotation during its early days, and particularly the development of the understanding of how flotation separations can be made by the
utilization of chemical reagents that interact with mineral surfaces. The second quarter of the
flotation century saw the development of most of the reagents and reagent schemes still used
today in flotation technology. Most of this chapter is concerned with a review of the fundamentals of flotation chemistry research, particularly the surface chemistry on which flotation is based.
The first decades of fundamental flotation research were oriented toward sulfide minerals, followed by extensive investigation of the flotation of oxide and silicate minerals, and then the sparingly-soluble salt minerals. More recent application of electrochemical and surface probe
techniques brought attention again to the flotation chemistry of sulfide minerals. Topics presented here are necessarily limited to broader aspects of sulfide mineral surface chemistry and the
role of oxidation in collection processes, the interfacial chemistry of oxide and silicate mineral flotation and the role of the electrical double layer and hydrocarbon chain association, and the
influence of aqueous solution chemistry on the flotation of sparingly-soluble salt minerals.
INTRODUCTION
No metallurgical process developed in the 20th century compares with that of froth flotation and the profound effect it had on the mineral industry. Most of the early developments
in flotation processing originated in Australia between 1900 and 1910. In the bulk oil processes that preceded froth flotation, generally the separation was aided by levitation of the
oil/mineral mass, either through the entrainment of air during mixing or by reduction of
pressure to generate bubbles, or by the addition of sulfuric acid to generate carbon dioxide
bubbles from carbonate minerals in the ore. Working independently as well as for Minerals
Separation Ltd., A.H. Higgins in London and G.A. Chapman at Broken Hill, Australia,
found that by reducing the oil content (oleic acid) to less than 1% and agitating the ore, the
mineral-laden bubbles rose to the surface (Rickard 1916). Modern flotation is attributed to
the resulting basic patent of Minerals Separation Ltd., where the aid of chemically generated
gas bubbles was definitively discarded in favor of air bubbles (Sulman, Picard, and Ballot
1905). The first operations in Australia simply involved bulk flotation to recover the fine
particles that were left behind in gravity concentration plants.
Froth flotation as it is known today is the process that had its beginnings 100 years ago
in Australia, but a graphite flotation process preceded it by nearly three decades. As stated
by Sutherland and Wark in 1955:
The brothers Bessel (1877) patented a true flotation process for the concentration
of graphite ores.… The modern flotation process differs little in principle from the
3
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4
HISTORICAL ASPECTS OF FLOTATION
Bessel process.… However, the work of the Bessel Brothers was forgotten, and the
modern process was evolved as the result of the work of many later investigators.
A German patent was issued to Gebrueder Bessel in Dresden, and using only 305
words, this 1877 patent outlined a process for the flotation of graphite from ores using 1%
to 10% of a nonpolar oil and listing 16 or more sources for the oil. After the ground ore was
mixed with the oil, this mass was added to water and the slurry then raised to boiling temperature. According to the patent, the graphite flakes attached to bubbles, rose to the surface, and were skimmed off to make the separation. From an ore containing 40% graphite,
the Bessel brothers produced a concentrate containing more than 90% graphite in an operation near Passau (Graichen et al. 1977). The market of their product was for the production
of graphite crucibles for smelting. In an attempt to reduce costs, in 1886 they patented
another gas-generating method for the process by adding acid with carbonates or metals
(Gebrueder Bessell 1886). About that same time, Ceylon graphite was discovered to be of
higher quality, which led to the demise of the Bessel graphite operation, and subsequently to
the disappearance of their process from the technical world.
In 1911, James M. Hyde installed the first flotation operation in the United States at
Basin, Montana, for the Butte and Superior Copper Company (Rickard 1916). Within 2 months,
Minerals Separation filed suit in the U.S. District Court in Montana for infringement
against Patent No. 835,120. This sparked the beginning of litigation in the early days of
modern flotation. Litigation affected the widespread adoption of flotation processing,
which is reflected in a paper by Barker (1928), who wrote:
Although flotation was known to be a successful process prior to 1912, Utah Copper Co.’s ores were not entirely treated by this process until 1923. Experiments had
been conducted, of course, prior to that time, and in February, 1917, the first unit
of the Arthur plant was changed over from gravity concentration to flotation….
The reasons for the delay in adopting flotation at these plants were, first, that it was
decided to await the outcome of the litigation with the Butte & Superior Mining
Co., which began with an injunction served on the plant on Oct. 3, 1911. This litigation continued for years.
After conversion of Utah Copper Co.’s operations to total flotation processing, the cutoff grade in mining was reduced and their reserves were enormously increased.
Data gathered by the U.S. Bureau of Mines shows the growth of flotation in the United
States, this growth being related to the development of selective flotation reagents and to
the increasing demand for mineral products (Varley 1928; Merrill and Pennington 1962;
Cooper 1980). Table 1 summarizes ore tonnages treated by flotation in the United States for
some representative years. The increase in ore tonnage processed by flotation in 1923 as
compared with that processed in 1919 resulted from the introduction of chemical flotation
reagents. Similarly, the marked increase in concentration ratio resulted from the advent of
selective flotation brought about by the introduction of these new chemical reagents, as will
be discussed later. In the early years, essentially only sulfide ores were treated by flotation,
but subsequently, processing other kinds of ores resulted from the development of new
reagents and reagent schemes. The huge increase in flotation processing in the United States
by 1960 resulted not only from increased copper ore production but also from extension to
other commodities, particularly phosphate and potash ores, as shown in Table 2. By 1980
there was a very significant increase in copper ore (due to lower grade) as well as in phosphate and iron ore flotation.
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
TABLE 1
5
Magnitude of the flotation industry in the United States for selected years
Year
1919
1923
1926
1960
1980
Ore Treated, Mt
24.08
34.29
46.16
179.86
404.34
Concentrates Produced, Mt
2.82
1.93
3.04
19.50
71.93
Concentration Ratio
8.6
17.7
15.2
10.8
5.6
TABLE 2 Types of ore treated by flotation and concentrates produced in the United States
(in million metric tons)
1926
Type of Ore
Copper
Lead-zinc
Gold-silver
Iron
Phosphate
Potash
Coal
Feldspar-mica
glass sand
Misc. industrial
minerals
Treated
39.89
5.57
0.48
0.23
1960
Concentrates
2.17
0.84
0.03
0.02
Treated
133.38
7.43
0.12
1.39
19.03
10.87
3.73
1.67
2.23
1980
Concentrates
4.82
0.49
0.003
0.54
6.37
2.83
2.54
1.06
0.83
Treated
211.61
11.39
0.10
37.88
108.70
12.93
11.70
11.58
0.58
Concentrates
4.67
0.84
0.005
21.48
26.63
2.99
6.86
8.51
0.37
In 1928, A.T. Tye wrote a landmark paper in which he described in detail not only how
selective flotation success was achieved in treating the problem ore at Cananea (Mexico) but
also the benefits to the Cananea Smelter of lowering the pyritic iron contamination in the
flotation concentrates. In 1923, with a combined gravity and bulk flotation flowsheet using
coal tar and pine oil as reagents, the grade of the concentrate was only 4.4% Cu, but by 1925
with selective flotation using xanthate, pine oil, and lime under very controlled conditions
necessitated by the soluble salts in the water, the flotation concentrates averaged 17.7% Cu.
Copper recovery by flotation increased from 87.4% to 91.2%, but overall recovery in the
smelter increased from 91% to 97% because of lower copper losses in the reduced amount of
slag. Further economic benefits resulted because much of the smelter could be shut down as
a result of the reduced tonnage of smelter feed. It is of interest that in discussion of this 1928
paper, G. Oldright suggested the promise of treating copper concentrates hydrometallurgically
instead of smelting them.
Mining geologist P. Billingsly (1928) expressed how flotation greatly expanded the role
of the exploration geologist:
The mining geologist searches for materials which the metallurgist can utilize, and
only such; and whenever an advance in metallurgy opens the gates for new materials, the geologist’s problem is correspondingly modified…. The metallurgist has
been the geologist’s best friend, and the geologist in turn has been able to help convert the metallurgist’s ideas into the concrete form of an increased ore supply.
Many authors of papers in Rickard’s edited classic 1916 monograph, The Flotation Process, asked questions about and speculated on the underlying phenomena involved in the flotation process. The overall objective of this chapter, therefore, is to show how many of those
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HISTORICAL ASPECTS OF FLOTATION
questions have been answered through research carried out during the last nine decades of
the flotation century. This chapter will primarily discuss the behavior of some typical
reagents and the surface chemistry involved in producing hydrophobic surfaces on minerals,
leaving factors that affect flotation kinetics to other presentations.
F L O TAT I O N R E A G E N T D E V E L O P M E N T
The success of any flotation separation depends on the range of chemical reagents added to
the system to control the surface behavior of minerals in the ore. Early flotation reagents for
sulfide mineral flotation were an almost unlimited range of various oils: coal tar derivatives,
crude petroleum, wood tars, and pine oils. Oleic acid could not be used where the gangue
minerals were calcareous. The coal tar derivatives contained sulfur compounds that probably possessed a certain amount of affinity for the sulfide minerals. Metallurgists were struggling to make separations between lead and zinc, copper minerals and pyrite, and oxide
minerals. Flotation entered a new era in 1921 when Perkins patented the slightly-soluble
thiocarbanilid as the first nonoily chemical collector for sulfide mineral flotation. James
Bean’s (1971) recollections illustrate the significance of this to a mill operator:
Thiocarbanilid for the first time gave the laboring metallurgist something that he
could add which would improve the collection of the sought-for mineral without,
at the same time, increasing the frothing to an uncontrollable degree. That this was
no small triumph was demonstrated practically to me while I was flotation operator at the Arthur mill of Utah Copper Company which at the time (1922) was
using Utah Copper’s own particular concoction of Barrett Oils and sulfur stewed
up together. Late on a sleepy afternoon an operator unduly increased the “oil”
being fed, hoping to lower mill tailing, but when the rougher froth got through
two cleaning steps neither the launders nor the floors could hold the resulting froth
and it literally ran out of the windows over a length of perhaps 40 feet and to a depth
of 3 or 4 inches. Years later I could still mark the area as I passed by on the highway.
Flotation reagents fall into six broad types: frothers, collectors, modifiers, activators,
depressants, and flocculants (natural and synthetic polymers). The frother is added to control bubble size and froth stability. Collectors are surface-active organic reagents that impart
hydrophobicity to minerals when they adsorb at mineral surfaces. The function of all other
reagents is to attain optimal conditions for selective separation of the minerals in an ore.
Activators are chemicals that enhance collector adsorption onto a specific mineral, whereas
depressants are reagents that prevent collector adsorption or prevent bubble attachment to
unwanted mineral surfaces. Modifiers constitute a broad range of inorganic and organic
compounds that modulate the flotation environment. Flocculants are added for assisting
dewatering of the flotation concentrates and are used in the selective flocculation/flotation
processing of nonmagnetic taconites. The great step forward that revolutionized the industry came with the 1925 patent of Keller for water-soluble xanthates as sulfide mineral collectors, followed by the patent of Whitworth (1926) for dithiophosphates. Table 3 provides a
brief glimpse of the amount and kinds of reagents used in the United States in two different
eras: 1925–1926 and 1980 (Varley 1928, Cooper 1980). In 1925, various oils were still used
as the collector with a large consumption of sulfuric acid to attempt selective flotation. In
1926, the change to xanthate collectors took hold, the use of oily collectors dropped sharply,
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
TABLE 3
7
Reagents used for flotation in the United States (in metric tons)
Reagent Amount
Ore Treated
Frothers
Collectors
Oils
Chemicals
Modifiers
Acids
Alkalis
Other
Activators
Depressants
Flocculants
1925
41,259,000
2,195
1926
41,616,000
2,935
1980
440,361,000
12,489
8,818
1,875
2,665
1,896
115,218
108,883
18,157
1,695
NA*
3,210
754
NA
2,061
75,701
NA
4,962
1,104
NA
35,169
413,055
28,735
3,925
33,389
18,069
*NA = Not available.
and that of alkalies dramatically increased in order to achieve the high pH necessary for
selective sulfide flotation. In 1925–1926, sulfidizing agents to treat oxidized lead and copper
ores accounted for about three-fourths of the activator consumption. By 1980, the total tonnage of ore treated was nearly ten times that in 1926. The application of flotation to processing nonmetallic ores resulted in higher reagent consumption because of amines, soaps, and
sulfonates being used as collectors along with various depressants. Oil consumption was
again high because of its use in phosphate and coal flotation.
Along with the quest for suitable organic chemicals having the ability to collect the
desired mineral in the froth, early flotation operators also tried to find agents to aid or
inhibit mineral floatability. Their discoveries, associated with the use of inorganic compounds in flotation, made possible the remarkable success achieved at present in the separation of sulfide minerals from each other, and in the concentration of oxides, silicates, and
salt-type minerals. A chronological account outlining some of these findings is presented in
Table 4, which also lists the flotation function of each chemical reagent. As this brief historical survey shows, most of the reagents used or known today were introduced during the first
half-century of flotation.
Reagent development had a great deal to do with improvement in the effectiveness of
flotation. The invention of Dow Chemical Company’s Z-200, a dialkyldithionocarbamate
by G.H. Harris and B.C. Fischback (1954), is undoubtedly the most significant sulfide flotation reagent development since the invention of xanthate as a flotation collector by Keller
and dithiophosphates by Whitworth shortly thereafter. The impact of Z-200 on sulfide ore
flotation, and particularly copper ore flotation, can be illustrated with data for 1979 as an
example (Harris, personal communication). In 1979, according to Harris, 4,500,000 kg of
this reagent (and its reproduction by other producers) were sold worldwide. At a reagent
consumption of 0.01–0.02 kg/t, there is an increase in copper flotation recovery of +2 percentage points. With the treatment of 300 Mt of copper sulfide ores worldwide at a grade of
0.7% Cu, this means that the invention of Z-200 gave the world an additional 40 million kg
of copper in 1979 alone.
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HISTORICAL ASPECTS OF FLOTATION
TABLE 4
Year
1901
1905
1909
1910
1912
1913
1921
1922
1923
1924
1925
1926
1928
1931
1934
1935
1952
1954
1965
1985
Milestones in the development of flotation reagents
Chemical Reagent
Sulfuric acid
Salt cake (NaHSO4)
Oils
Sodium sulfide
Ketones, aldehydes
Alkalis
Sodium dichromate
Sulfur dioxide
Copper sulfate
Thiocarbanilid
Cyanide
Alkali sulfides
Soluble sulfites
Soaps
Alkali xanthates
Dithiophosphates
Sodium silicate
Sodium carbonate
Starch
Alkyl sulfates
Amines
Polypropylene glycols
Thionocarbamates
Hydroxamates
Alkoxycarbonyl adducts
Function
Gas-bubble generator
Gas-bubble generator
Collectors for sulfide minerals
Activator for oxidized heavy-metal minerals
Soluble frothers
Sphalerite depressants
Galena depressant
Sphalerite depressant
Sphalerite activator
Slightly soluble chemical collector
Sphalerite and pyrite depressant
Sphalerite and pyrite depressants
Sphalerite depressant
Collectors for nonsulfide minerals
Soluble collectors for heavy-metal minerals
Collectors
Depressant
pH regulator
Depressant
Nonmetallic mineral collectors
Cationic collectors
Water-soluble frothers (polyethers)
Sulfide mineral collectors (copper)
Chelating agent for collector of Cu, Fe oxide
Collectors/modifiers for sulfides and nonsulfides
B R I E F C H R O N O L O G Y O F F L O TAT I O N R E S E A R C H
Many early efforts at understanding flotation were directed toward explaining differential
flotation in terms of the relative occlusion of gases, which would be driven out to nucleate
bubbles, thereby giving rise to selective flotation. In 1916, bubbles were considered to be at
the heart of flotation science, and Rickard (1916) postulated how progress in flotation
would be made: “…we know that the key to the flotation process is to be found not in the oil,
the acid, or the apparatus, but in the bubbles. The man who understands the physics of a
soap bubble has mastered the chief mystery of flotation.” As important a component as they
are in the process, bubbles usually play an inert role in flotation and merely provide a means
for levitating the desired mineral particles into a froth layer. Although industrial operators
and reagent manufacturers devoted effort toward finding cheaper chemicals that might act
as frothing agents and might alter froth characteristics, through the years bubbles have never
received the attention from flotation researchers speculated on by Rickard. However, in
1934 Gaudin commented:
Developments in flotation have been so rapid that one of the essential factors at
play—namely, the chemical effects of dissolved gases—has received scant attention. Recent theories have shown that gases are of extreme importance in many
instances. It is not unlikely that control of flotation can be exercised through control of the gases.
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
9
In a 1915 paper, Ralston suggested that flotation might result from the electrical attraction between negatively charged air bubbles and positively charged mineral particles, and
this postulate was actively debated for some years before eventually being discarded (Ralston
1916). However, today it is known that potential differences between bubbles and particles
enter into the kinetics of bubble-particle attachment.
The role of interfaces in flotation had been considered by Sulman by 1912 (see Rickard
1916) when he published the results of an investigation carried out at Minerals Separation
to determine the magnitude of the contact angle that various minerals needed before they
were wetted and would sink when brought into contact with a water surface. About these
phenomena, Ralston wrote in 1915 of adsorption changing contact angles and the properties of interfacial films:
A glance at Clerk Maxwell’s famous paper on capillarity upon which Reinders’
work is based, will suggest immediately the explanation of a contact angle, and that
it is the result of a certain equilibrium of interfacial tensions of air, water, and
solid.… There can be no doubt that there is a close parallelism between the angle of
hysteresis of the contact angle and the ability of a mineral to float.… To go into this
a little farther, we ought to consider the properties of the surface layers of the substances involved.… One important property of this film is that it will often take up
dissolved substances in different proportion from the amounts in which they are
taken up in the bulk solution, and there always is a definite equilibrium between
the two.… The properties of these interfacial films have been found to be greatly
modified by small amounts of dissolved substances. The importance of the study of
interfacial films becomes obvious.
The first direct application of thermodynamics to systems similar to flotation was that
of von Reinders (1913). Based on Maxwell’s capillarity equations, von Reinders deduced
how fine solid particles would be distributed between oil and water phases. For example,
using γ to represent the interfacial tensions at the oil–water (ow), solid–water (sw), and
solid–oil (so) interfaces, von Reinders showed that the solid will disperse in the aqueous
phase if γso > γow + γsw. Analogous relations give conditions under which the solid will disperse in the oil phase or concentrate at the oil–water interface. The three interfacial tensions
are interrelated with contact angles by the Young equation. Ralston suggested that von
Reinders’ relations might explain how interfacial tensions control flotation. In 1917, Taggart and Beach fairly lucidly applied these concepts directly to flotation. Several decades
would elapse before thermodynamics would become a fairly widely used tool for the analysis
of flotation phenomena.
In 1917, Anderson suggested that adsorption might play a dominant role in flotation.
Anderson discussed the Gibbs adsorption equation in relation to frother adsorption at the
air-water interface and, interestingly, stated: “An electric charge on an adsorbed substance
probably would considerably influence the amount adsorbed.” In 1920, Langmuir showed
that oleic acid created large contact angles on cleaved calcite and galena but only small
angles on clean glass and cleaved mica. Oleic acid was irreversibly adsorbed on calcite and
galena but not on glass and mica. He suggested further research with other kinds of reagents
on clean mineral surfaces. In 1928, Taggart described the results of adsorption tests on sulfide minerals that related the structure of the adsorbate to its ability to act as a flotation collector. He wrote that powdered sulfide minerals abstracted 90% of the thiocarbanilid in a
solution and captive-bubble experiments showed the sulfide to be hydrophobic. The
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10
HISTORICAL ASPECTS OF FLOTATION
adsorbed thiocarbanilid could be leached off with ethyl alcohol. Similar experiments with
thiourea (which has the same formula as thiocarbanilid but without the phenyl groups
attached to the nitrogen) showed that sulfide minerals adsorb thiourea almost as much as
thiocarbanilid but there is no collecting action. With carbanilid, there was no removal from
solution. These experiments led to Taggart’s formulation of the definition of the molecular
structure needed for a soluble flotation collector, namely, that it must possess both a polar
group that binds it to the surface and a nonpolar group that can orient away when adsorbed
at a mineral–water interface.
NHC6H5
S
C
NHC6H5
O
NHC6H5
Thiocarbanilid
C
NH2
S
NHC6H5
Carbanilid
C
NH2
Thiourea
What might be considered to be the first adsorption isotherm of a soluble flotation collector on a mineral are the results published by Taggart, Taylor, and Knoll in 1930 for the
abstraction of potassium ethyl xanthate (KEX) by ground galena as a function of reagent
concentration in solution. It would be some time before adsorption isotherms could reliably
be determined quantitatively and several years until methods were developed for determining specific surface area, for radioactively marking adsorbates, for spectrophotometrically
measuring reagent concentrations in solution, and for quantitatively analyzing infrared
absorption spectra. Because nearly all of the early flotation operations involved sulfide ores,
the behavior of sulfide minerals received nearly all of the initial research attention.
Although such early researchers as Fahrenwald, Sulman, and Taggart carried out a number of experiments to elucidate flotation phenomena, the founder of the scientific basis of
flotation was A.M. Gaudin. The first systematic research that opened the way toward understanding the chemistry of the flotation process was the extensive, dedicated investigation
initiated in 1926 at the University of Utah under Gaudin, using high-purity single minerals
in a miniature flotation cell (50 g of pure 100 × 600 mesh cleaned samples) that had been
developed at the Utah Engineering Experiment Station by Gates and Jacobsen (1925). In
1928, Gaudin described their laboratory approach:
It is a generally recognized scientific principle that to investigate a certain set of
phenomena one variable must be allowed to vary at one time while other variables
are kept strictly constant. Therefore, to obtain consistent results in flotation
research, pure minerals having a definite size should be used either by themselves or
as artificial mixtures. These minerals should have an especially clean surface,
cleaned in standard fashion, and the test should be run in a standard machine
cleaned in standard fashion, for a standard length of time after a standard preagitation period at a definite temperature. All reagents should preferably be added in
solution to eliminate the necessity for conditioning. Distilled water should be used
throughout.
With these guiding principles, this early work by Gaudin and his colleagues was the
beginning of the modern approach to research in flotation chemistry. In the author’s opinion, Gaudin was indeed the father of fundamental flotation research as it is known today.
Figure 1 illustrates the quality of flotation experiments conducted with carefully cleaned
mineral samples and high-purity reagents (Gaudin et al. 1928). In Gaudin and colleagues’
original paper, the flotation recovery of 100 × 600 mesh galena was presented as a linear
function of the fatty acid addition in pounds per ton. By recalculating the published results
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
Flotation Recovery, %
100
C12
Galena
0.2 kg/t Terpineol
C11
C10
C9
C8
11
C7
80
60
C6
40
20
C6
C7
C8
C9
C10
C11
C12
Caproic Acid
Enanthic Acid
Caprylic Acid
Pelargonic Acid
Capric Acid
Undecanoic Acid
Lauric Acid
0
0.01
0.1
1.0
10.0
Fatty Acid Addition, mol/t
Adapted from Gaudin et al. 1928.
FIGURE 1 Flotation of 100 × 600 mesh galena with fatty acids of different alkyl chain lengths
ranging from 7 to 12 carbon atoms, with the reagent addition in mol/t (mol wt of C11 lauric acid
is 200)
in terms of moles per metric ton and replotting those data in a semilogarithmic manner as
shown in Figure 1, their results show the very systematic effect of the number of carbon
atoms in the alkyl chain expected by the Traube rule. Because the carbon atom in the carboxyl head group is not part of the alkyl chain, lauric acid (mol wt 200) is given as an 11-carbon
reagent in Figure 1. This systematic chain-length effect indeed substantiates the validity and
care taken in their work. Interestingly, Gaudin never continued using mini-scale flotation
cells in his research after he moved from the University of Utah to the Montana School of
Mines.
Flotation processing technology did not come into being as a result of an intensive fundamental research effort, but, in a manner similar to the development of so much of the
other technology used in the processing of raw materials, it was developed over the years
through much empirical and intuitive work on complex ores. Fundamental understanding
of flotation resulted from careful experimentation with well-controlled systems, later followed by a firm grounding in physicochemical principles, including thermodynamics, surface and colloid chemistry, and electrochemistry. Major headway in understanding the
flotation chemistry of sulfide mineral flotation started shortly before 1930, and that of nonmetallic mineral flotation shortly before 1950. Prior to about 1950, most of the fundamental investigations were directed toward the flotation chemistry of sulfide mineral
separations. To achieve the desired separations from complex ores, the early research (1925–
1935) was mainly centered on interactions between mineral surfaces and sulfhydryl flotation reagents. The leading researchers, chronologically, in that era were Taggart and Gaudin
in the United States, and Wark in Australia. The key issues were the mechanism of interaction between the reagent and the mineral surface (by Taggart and by Gaudin), identification
of species responsible for flotation (by Gaudin), and the assessment of chemical conditions
for floatability (by Wark ). About mid-century most of the research shifted to oxides, particularly
quartz, corundum, hematite, rutile, and silicates. In the last quarter of the flotation century,
much attention was directed toward the flotation chemistry of the sparingly-soluble salt
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12
HISTORICAL ASPECTS OF FLOTATION
TABLE 5 Selected experimental techniques that provided significant advances toward
understanding flotation chemistry
Major Experimental Techniques
Single-mineral laboratory flotation mini-scale
50-gram scale
5-gram scale
Modified Hallimond tube flotation
Captive-bubble contact angle determination
Adsorption density/isotherm
Electrokinetics (zeta potentials)
Infrared spectroscopy
Ex situ
In situ
Researchers
J.F. Gates and L.K. Jacobsen
A.M. Gaudin
M.C. Fuerstenau
A.F. Hallimond
E.W. Ewers
D.W. Fuerstenau
H.S. Choi and I. Iwasaki
A.F. Taggart, T.C. Taylor, and C.R. Ince
I.W. Wark and A.B. Cox
R.H. Ottewill
J.S. Laskowski
S. Chander and D.W. Fuerstenau
A.M. Gaudin
P.L. deBruyn, I. Iwasaki, and G.A. Parks
P. Somasundaran and D.W. Fuerstenau
J.M. Cases
S.C. Sun and A.M. Gaudin
A.S. Buchanan and D.J. O’Connor
D.W. Fuerstenau
M.E. Wadsworth and A.S. Peck
J. Leja and G.W. Poling
J.D. Miller
J.A. Mielczarski
J.D. Miller
I.W. Wark and A.B. Cox
Electrochemistry
Rest potential
J.C. Nixon and S.G. Salamy
Polarization: voltammetry
J.T. Woodcock and M.H. Jones
Impedance spectroscopy
R. Woods
W.J. Trahar
S. Chander and D.W. Fuerstenau
P.A. Richardson
minerals, particularly apatite, calcite, dolomite, and bastnaesite. Problems of energy supply
gave rise to research on coal flotation and coal desulfurization. With the advent of newer
electrochemical techniques, major effort resumed in the last quarter of the flotation century
to extensive investigation of sulfide mineral flotation phenomena.
Although numerous experimental methods have been applied to investigating flotation
phenomena, several techniques have been widely used and have been responsible for the
greatest progress. These are summarized in Table 5, together with the names of several of
those who developed or applied these techniques to the study of chemical phenomena
involved in flotation. Numerous other techniques have been devised and utilized through
the years to study flotation phenomena, but they are not included here because they may not
have yielded definitive results or may not have had the impact or widespread use of the seven
techniques given in Table 5. Examples of some of these techniques (and some of researchers
who used them) include vacuum flotation (R. Schuhmann and B. Prakash), bubble-pickup
(S.R.B. Cooke), induction time measurement (I. Sven-Nilsen; V.A. Glembotsky; R.H.
Yoon), film flotation (M.C. Williams and D.W. Fuerstenau), microcalorimetry (O. Mellgren),
and radiography (I.N. Plaksin). There has been worldwide interest in surfactant adsorption
behavior at solid–water interfaces in recent years, resulting in many new tools having been
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
13
used to probe the detailed nature of adsorbed surfactant and polymer films. For example, P.
Somasundaran and his graduate students at Columbia University have extensively used such
newer molecular-level-information-yielding techniques as absorption, emission, magnetic
resonance, and scattering spectroscopic techniques (fluorescence, electron spin resonance,
excited state resonance, Raman, etc.) along with adsorption, flotation, flocculation, and
electrokinetic studies to gather information about the microscopic properties of the
adsorbed surfactant and polymer films. X-ray photoelectron spectroscopy, or XPS (A.N.
Buckley and R. Woods), has been used to identify chemical species at mineral surfaces. Secondary ion mass spectrometry, or SIMS (D.R. Nagaraj), has recently been utilized to clearly
show the nature of complexes adsorbed at mineral surfaces. Atomic force microscopy has
been applied to the study of the nature of adsorbed surfactant films (R.H. Yoon; T.W.
Healy; J.D. Miller).
Starting in the 1950s, two of the relatively simple techniques listed in Table 5 were
widely adapted to the study of flotation chemistry effects. When it became understood that
any ion that strongly adsorbs at a mineral–water interface is reflected in its effect on the zeta
potential, the use of zeta potential measurements in flotation surface chemistry spread rapidly, and particularly so because of the simplicity of electrophoresis techniques. The modified Hallimond tube permitted study of flotation response without changing the solution
composition (because no material is removed as a mineral-laden froth from the device during an experiment); this permitted direct correlation with the solution chemistry of the system. Almost all of the experimental investigations on flotation chemistry carried out during
the first half-century involved the use of a single experimental technique, such as flotation
testing, contact angle measurement, identification of surface species, determination of
adsorption isotherms, and so forth. However, using a number of different experimental
techniques to probe the behavior of the same system led to being able to make correlations
among various types of interfacial phenomena in flotation systems, and this led to a more
complete understanding of the surface chemical processes involved. An example of such a
correlation is given in Figure 2, which presents the zeta potential, adsorption density, contact
angle, and flotation response of quartz with dodecylammonium acetate (DAA) as collector
(D.W. Fuerstenau, Healy, and Somasundaran 1964). Here, two-phase mineral–water interfacial phenomena (adsorption density and zeta potential) correlate well with three-phase
behavior (contact angle and flotation response). The first such correlation was published in
1957 for the DAA–quartz system at constant collector concentration with pH as the variable; later results for the same system at constant pH but with collector concentration as the
variable are somewhat easier to explain and are therefore given in Figure 2. (The reasons for
the sharp breaks in the curves that occur at hemimicelle concentration [HMC] will be discussed in a later section.)
Major advances, particularly starting in the 1950s, were achieved through better understanding and application of the fundamental principles of surface and colloid chemistry,
particularly electrical double-layer phenomena, to flotation systems. In part, this was
strongly influenced by Professor J.Th.G. Overbeek’s year at the Massachusetts Institute of
Technology (MIT) with the mineral engineering group of Gaudin, and disseminated worldwide by the generations of students that followed. Detailed analysis of the thermodynamic
stability of minerals and reagents, speciation of complexes in aqueous solution, and solubility phenomena have also helped expand the understanding of different types of flotation
systems. All of this, combined with application of the many new techniques for probing
mineral–water interfaces at the molecular level, led to much of the research in the second
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14
HISTORICAL ASPECTS OF FLOTATION
5
Quartz
pH 6–7
Contact Angle
Adsorption Density
Zeta Potential
Flotation
80
+40
0.85
2
0.90
60
0
40
Zeta Potential, mV
0.80
Flotation Recovery, %
3
+80
Cosine θ
Adsorption Density, μmol/m 2
4
Hemimicelle Concentration
–40
1
0.95
20
–80
0
10–6
10–5
10–4
10–3
DAA Concentration, mol/t
Adapted from D.W. Fuerstenau, Healy, and Somasundaran 1964.
FIGURE 2 Correlation of adsorption density, contact angles, and zeta potentials with the
flotation of quartz at pH 6–7 as a function of DAA concentration
half of the flotation century being devoted to elucidating the detailed principles of mineral–
reagent interactions in flotation. In the sections that follow, some of the major advances in
understanding the flotation chemistry of various mineral systems will be briefly reviewed.
S U L F I D E M I N E R A L F L O TAT I O N C H E M I S T R Y
Because flotation was first applied to the recovery of sulfide minerals from ores, all of the
early research was conducted on sulfide minerals, particularly galena, sphalerite, chalcocite,
chalcopyrite, and pyrite. The first systematic investigations on sulfide mineral flotation were
the pure mineral flotation experiments of Gaudin and his associates at the University of
Utah (Gaudin et al. 1928). Their initial research was concerned with the behavior of galena.
Gaudin (1932) stated that
“…pure, unoxidized galena floats readily without the addition of a collecting agent,
a frother alone being required. This can be ascertained by grinding pure galena particles in water under anaerobic conditions, and floating immediately.… In practice
galena particles are more or less oxidized during grinding and classification, requiring
varying amounts of collecting agents.”
For fatty acids as collector, Figure 1 illustrates the quality of their results. Most of their
work was conducted with xanthates and other sulfhydryl reagents as the collector, and Figure 3 presents the results of Gaudin et al. (1928) for the flotation of galena with xanthates of
different chain length, but again with the xanthate additions being recalculated in terms of
moles per metric ton, rather than pounds per ton, and plotted semilogarithmically. The
amount of xanthate required for complete flotation with xanthates of two or three methylene groups is extremely low, merely about 0.1 mol/t of mineral, showing an extremely high
affinity for the surface that is not strongly dependent on chain length if the collector has
three or four carbon atoms. Although no specific numbers are available, nearly all of the
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
15
Flotation Recovery, %
100
80
60
40
Methyl
Ethyl
Propyl
Butyl
20
0.05 kg/t Terpineol
0.5 kg/t Sodium Carbonate
0
0.01
0.1
1.0
10.0
Potassium Alkyl Xanthate Addition, mol/t
Adapted from Gaudin et al. 1928.
FIGURE 3 Flotation of 100 × 600 mesh galena with alkyl xanthates of different alkyl chain
lengths with reagent additions expressed in mol/t (mol wt of KEX is 160)
added xanthate would have been abstracted from solution. In the case of fatty acid flotation,
there probably was considerable residual carboxylate left in solution.
Given that the surface of galena readily oxidizes, the oxidation products must enter into
the adsorption process. In 1934, Taylor and Knoll conducted a careful set of experiments to
quantitatively determine the exchange process involved in the uptake of ethyl xanthate by
galena, using an iodometric titration technique to determine the xanthate concentration in
solution. Taking one set of measurements as an illustration of their findings, with all concentrations being expressed as equivalent to 25 mg KEX per liter of solution, the original concentration of xanthate in solution was 200.0 mg/L, the amount of xanthate ion abstracted
was 58.3 mg, and the stoichiometric equivalent of reduced sulfur-oxygen ions emitted was
13.8 mg, 16.1 mg sulfate ions emitted, and 27.2 mg carbonate ions—or a total stoichiometric equivalent of 57.1 mg. Clearly, xanthate uptake by galena was exactly balanced by an
exchange with oxidation product ions at the surface.
In 1934, Wark and Cox presented some data on the contact angle of an air bubble on
galena as a function of the concentration of KEX in solution. Their data given as milligrams
of collector per liter have been converted to moles per liter (mol wt = 160) and are plotted
in Figure 4. The results tend toward the maximum contact angle of 60°, after increasing
sharply to about 50° at concentrations below 20 or so micromoles per liter. The 1928 results
from Gaudin et al. were recalculated in terms of micromoles per kilogram of 100 × 600 mesh
galena and are also plotted in Figure 4. This plot shows that about 400 μmol/kg of galena is
required to achieve 90% recovery. Assuming that most of the added xanthate was adsorbed,
in 1957 Gaudin estimated that roughly monolayer adsorption was achieved at this ethyl
xanthate addition. However, in that same year, Bogdanov et al. (1957) published a paper
that presented a summary of extensive work conducted in Russia on the adsorption of different reagents on various minerals using a number of radioactively marked adsorbates,
together with their effect on flotation response. Their results for the flotation recovery of
galena as a function of the percentage of monolayer coverage of ethyl xanthate are also plotted
in Figure 4. These experiments show the strong affinity of a sulfhydryl collector for the surface
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16
HISTORICAL ASPECTS OF FLOTATION
Contact Angle: μmol/L Solution
60
80
40
60
40
Contact Angle, degrees
Flotation Recovery, %
100
20
Flotation
Contact Angle
Adsoprtion
20
0
100
0
0
20
40
60
80
Collector Addition, [ μmol/kg galena][0.1]
Adsorption Density, % monolayer
FIGURE 4 Contact angle on galena as a function of the concentration of KEX expressed in
mol/L (data from Wark and Cox 1934), the flotation of galena as a function of the adsorption
density of ethyl xanthate expressed in terms of monolayer fraction (data from Bogdanov et al.
1957), and the flotation of galena as a function ethyl xanthate addition expressed in terms of
μmol/kg (data from Gaudin et al. 1928)
Flotation Recovery, %
100
80
60
40
Chalcocite
Pyrite
20
0.015 kg/t KAX
0.10 kg/t Terpineol
0
0
2
4
6
8
10
12
14
pH
Adapted from Gaudin 1929.
FIGURE 5
collector
Effect of pH on the flotation of 100 × 600 mesh chalcocite and pyrite with KAX as
of sulfide minerals and also show that experiments conducted with pure systems under controlled conditions can exhibit agreement among different measures of mineral-collector
interaction.
Regulation of pH has been the most important method for regulating flotation chemistry. In 1929, Gaudin first published the results of his measurements of the flotation of a
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
17
variety of minerals as a function of pH. Figure 5 presents the flotation recovery of chalcocite
and pyrite as a function of pH with 0.015 kg/t of potassium amyl xanthate (KAX) and
0.10 kg/t terpineol as frother. The results clearly show that there is sharp decrease in the flotation of pyrite as the pH is increased above 6 and that chalcocite remains fully floatable
under these conditions until the pH exceeds about 13. This was definitive work showing the
depressant role of pH in the flotation of sulfide minerals. Gaudin on several occasions commented that not patenting the use of pH control for selective flotation was one of his oversights.
In the early 1930s at the University of Melbourne, I.W. Wark (personal communication) initiated an important program to understand the flotation chemistry of sulfide minerals. Wark’s research group spent a prolonged effort, essentially a year or more, in first
refining the contact angle measurement technique of Taggart, Taylor, and Ince (1930) and
sample preparation so that reliable and reproducible results could be obtained. In 1934,
Wark and Cox published the first of a remarkable set of papers in which they presented their
classic diagrams showing the relationship between collector concentration and pH for conditions of incipient flotation, and for the behavior of a wide variety of modifiers and depressants with various collectors. Figure 6 presents one of their critical pH diagrams for three
sulfide minerals—namely, pyrite, galena, and chalcopyrite—with sodium diethyldithiophosphate as collector. In each case, flotation should occur under conditions to the left of
the curve. Diagrams such as these provide a means for predicting conditions under which
flotation separations can be made. If one considers that hydroxyl ions adsorb competitively
with collector ions, that the amount of collector adsorbed under conditions of incipient flotation is constant, and also that the standard free energy of adsorption is constant, then each
line in Figure 6 must be characterized by [X–]/[OH–] being constant. These critical pH
curves were a major contribution to early flotation theory and they show, for example, the
pH and collector concentration at which flotation does or does not take place. In discussion
of the 1934 Wark and Cox paper where KEX was used as the collector, Barsky (1934)
pointed out for their experiments that [X–][H+] was constant along their critical pH curves
and that the results could be interpreted as xanthic acid [HX] being constant along these
curves. Gaudin (1957) interpreted the results in terms of ion exchange between adsorbed
X– and OH– for surface sites.
Wark and co-workers (Sutherland and Wark 1955) also measured contact angles of various thiol collectors having a range of carbon atoms in their nonpolar groups. For example,
they found the contact angle of collectors having an ethyl group on the nonpolar chain
to be 60° on all sulfides. This included xanthate, mercaptan, dithiophosphate, disubstituted
dithiocarbamate, and others. Methyl xanthate and disubstituted dithiocarbamate produced
contact angles of 50°.
For nearly 25 years, there was spirited and ongoing debate about the mechanism of collection in sulfide mineral systems. Gaudin was a strong proponent of adsorption as the
means of collector uptake by minerals. In 1927 he wrote, “The mechanism by which xanthates float other sulfides than galena may involve an adsorption of xanthate ions without
further reaction.” On the other hand, Taggart was convinced that collectors coated mineral
surfaces by chemical reaction. In 1930, Taggart, Taylor, and Knoll wrote, “All dissolved
reagents which, in flotation pulps, either by action on the to-be-floated or on the not-to-befloated particles affect their floatability, by function of the reason of chemical reactions of
well recognized types between the reagent and the particle affected.” Taggart’s shortcoming
was his belief that the chemical theory of flotation was all-inclusive, even with regard to oils
on naturally hydrophobic minerals, and for collectors that do not form insoluble products
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18
HISTORICAL ASPECTS OF FLOTATION
700
Pyrite
Galena
Chalcopyrite
Collector Concentration, mg/L
600
500
400
300
200
100
0
2
3
4
5
6
7
8
9
10
11
pH
Adapted from Wark and Cox 1934.
FIGURE 6 Critical pH curves for the flotation of pyrite, galena, and chalcopyrite with sodium
diethyl dithiophosphate as collector
such as amines on minerals. In reality, Taggart’s chemical theory of collection is merely
exchange adsorption (as was shown by Taylor and Knoll [1934]). Overall, Wark was another
advocate of the adsorption theory of collector uptake. In 1934, Wark and Cox wrote, “We
find there is a strong connection between adsorption of xanthates and the solubility of the
heavy-metal xanthates, but we are unable to decide if this is an identity.” In 1950, Cook and
Nixon were as forceful in promoting the concept that sulfide mineral flotation takes place
by neutral molecule adsorption as Taggart had been in his promoting the idea of chemical
reaction. They wrote, “Assuming a complete or nearly complete monolayer of ‘ions’ on the
mineral particles, one would obtain a bulk concentrate with so much charge that it would
explode with greater violence than an equal weight of nitroglycerine!” M.A. Cook, an
expert in explosives and an outstanding solution physical chemist, did not think in terms of
the electrical double layer because in all cases of ion adsorption, counterion adsorption or
exchange adsorption keeps the system electrically neutral. Note that Cook’s neutral molecule theory is the same idea that Barsky (1934) had presented in his discussion of the critical
pH curves of Wark and Cox in 1934. There are many examples where the collector indeed
appears to adsorb in its neutral molecule form. In 1967, Steininger showed that the upper
pH limit for the flotation of sphalerite with a wide variety of thiol collectors was a function
of their pKa values. Such results indicate that the chemisorption of the neutral molecule may
indeed have a role in flotation in this mineral–collector system. Raghavan and Fuerstenau
(1975) demonstrated that the neutral hydroxamic acid molecule appears to be the active
adsorbing species in the hematite–hydroxamate system. However, as will be subsequently
discussed, when a cationic amine collector hydrolyzes to the neutral molecule species with
oxide minerals, flotation ceases. In 1957, Nixon wrote, “Prominent theories could be reconciled by the electrochemical approach.” In 1984, Woods summarized sulfide flotation as follows: “Electrochemical investigations of the interaction of the thiol collectors with sulfide
minerals have demonstrated that each of the three anodic processes—chemisorption, reaction
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
1,014
1,212
1,110
100
19
50
1,110–
1,210
1,020
1,112
B
1,140
0
50
990
987
50
0
1,300 cm–1
D
1,115
0
1,020
1,195
50
1,110
C
0
1,140
Absorption, %
986
0
50
A
1,200
1,100
1,000
E
900
800
700
Adapted from Leja, Little, and Poling 1962–1963.
FIGURE 7 Infrared spectra showing adsorption of ethyl xanthate onto an evaporated PbS film:
(a) bulk lead ethyl xanthate, solid on Nujol mull; (b) freshly evaporated PbS film after
atmospheric oxidation; (c) PbS film treated in aqueous solution of ethyl xanthate; (d) after
prolonged washing in ether; and (e) after washing in pyridine
to form a metal collector compound, and the formation of a dithiolate—plays a role in creating hydrophobic surfaces.”
New instrumentation permitted identification of species at the surface and quantification of energies involved in surface reactions. In a seminal study, Leja, Little, and Poling
(1962–1963) applied infrared spectroscopy to demonstrate the nature of collector species at
mineral surfaces. Figure 7 presents their classic infrared spectra showing the adsorption of
ethyl xanthate onto an evaporated lead sulfide (PbS) film. The top curve (a) in Figure 7,
taken from their work, shows the infrared spectrum of bulk lead ethyl xanthate, and the second spectrum (b) is for a lead sulfide film that has been oxidized in the atmosphere. After
exposing that film to xanthate in solution, they obtained the spectrum (c) that is virtually
identical to that of lead ethyl xanthate, showing that indeed a chemical compound is formed
at the surface. Washing with ether (d) removed some of the surface lead xanthate, but it took
a strong solvent, pyridine, to completely remove the xanthate, returning the spectrum (e)
back to that of a lead sulfide (oxidized) surface. Infrared spectroscopy has become a widely
used tool to study the nature of adsorbed films in flotation systems.
The energetics of the interaction of xanthate with galena was carefully determined by
Mellgren (1966) using microcalorimetry techniques. First, Mellgren reacted lead sulfate
with xanthate. Then he reacted xanthate with galena that had lead sulfate on its surface and
again measured the heat that evolved. Mellgren’s measurements of the heat of reaction for
these two cases gave identical results; namely, that the enthalpy is –22 kcal/mol Pb2+ in each
case. These measurements clearly indicate that the uptake of xanthate by oxidized galena is
energetically equivalent to the chemical exchange reaction forming lead ethyl xanthate from
lead sulfate. He conducted similar studies with lead carbonate. Mellgren also observed that
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20
HISTORICAL ASPECTS OF FLOTATION
at high pH, galena ceases to float because basic lead compounds form on the surface in preference to lead xanthate. Thus, depression in this system may also occur by a chemical reaction mechanism, that is, by the formation of a surface compound that is more insoluble than
the surface collector salt.
As early as 1931, Kamienski measured the rest potential of a galena in xanthate solution. In 1934, Wark and Cox presented a direct correlation of the potential of a copper electrode and the prevention of xanthate adsorption on chalcopyrite. However, about two
decades elapsed before electrochemical methods were again used in flotation research for
investigating mechanisms of thiol collector–sulfide mineral interaction. Salamy and Nixon
(1953) found that the concentration of xanthate determines the rest potential of the mercury electrode, which corresponds to the xanthate–dixanthogen reversible potential. They
concluded that the adsorbing species at the metal–solution interface is dixanthogen, the oxidation product of the xanthate ion. Definitive papers on xanthate flotation of pyrite, published independently by M.C. Fuerstenau, Kuhn, and Elgillani (1968) and by Majima and
Takida (1968), showed dixanthogen to be the surface species responsible for flotation in this
system. The former researchers used infrared spectroscopy and oxidation potential (Eh)
measurements, and the latter conducted polarization experiments in basic medium with a
pyrite electrode. It is interesting to note that Gaudin and Wilkinson wrote in 1933 that
“Pyrite, or ferric iron derived from it by oxidation, changes xanthate to dixanthogen; the
dixanthogen can be extracted from the mineral surface provided oxidation of the dixanthogen is prevented.” Their observation lay dormant for the next 35 years.
Numerous electrochemical investigations have been made in sulfide mineral flotation
systems, particularly conducting linear potential sweep voltammetry for various systems.
Chander and Fuerstenau (1975) first combined this technique with the measurement of
contact angles. Woods (1984), Woods and Richardson (1986), and Chander (1985) presented thorough reviews of the electrochemistry of sulfide mineral flotation. The cathodic
step usually involves the reduction of oxygen:
O 2 + 2H 2 O + 4e – = 4OH –
(EQ 1)
and is coupled with an anodic step involving oxidation of either the collector or the mineral.
The products of the anodic reaction depend on the mineral and the collector used, and the
pretreatment of the mineral surface. They have been identified as chemisorbed collector,
dithiolates, and metal–collector compounds. As suggested by Woods and Richardson
(1986), the reactions for the anodic oxidation of the collector can be written as follows:
A. Charge transfer chemisorption of a thiol ion (X–):
X – = X ( ads ) + e –
B. Oxidation of a thiol ion to the disulfide:
2X – = X 2 + 2e –
C. Formation of a thiol compound with a metal component of the mineral:
MS + 2X – = MX 2 + S + 2e –
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
21
This mixed potential model does not actually contradict earlier theories of sulfide mineral flotation. Gaudin’s and Wark’s adsorption models would be equivalent to Reaction A,
whereas Taggart’s chemical reaction theory would be represented by the anodic process
given by Reaction C, which can be a two-step reaction involving oxidation of the mineral
and metathesis.
In a classic paper, Allison et al. (1972) presented the results of a detailed study of the
products resulting from the interaction of sulfide minerals with thiol collectors. Table 6 presents a summary of their identification of surface products in ethyl xanthate–sulfide mineral
systems and a correlation of those results with the rest potential of the minerals. These
results clearly indicate that, with the exception of covellite, sulfide minerals that display a
rest potential below the corresponding reversible potential for the oxidation of the collector
to its disulfide (dixanthogen, in this case) will react with thiols, forming metal tholates.
Those sulfide minerals whose rest potentials are above this value oxidize the thiol to its
disulfide, and this product is now accepted as the collecting species in these systems. Similar
results were reported for the diethyl dithiocarbamate–sulfide systems. A necessary condition for sulfide mineral floatability is the thermodynamic stability of the hydrophobic entity
formed at the surface. If this compound forms at a rate that is too slow compared to the rate
at which the sulfide mineral oxidizes, however, the mineral will not float because surface oxidation may result in the formation of oxide-type materials that may not only impede electron transfer from the collector to the mineral but are also extremely hydrophilic. If the
species is decomposed by oxidation, strong oxidizing conditions would be detrimental to
flotation. Knowledge of the Eh of the system is therefore vital in sulfide mineral flotation. It
is possible to predict the oxidizing and reducing conditions necessary for a substance to be
thermodynamically stable by constructing the appropriate Eh–pH diagram.
Chander (1985) grouped sulfide minerals into two types—reversible and passivated—
based on their electrochemical behavior in aqueous solutions. The reversible group of sulfide minerals includes galena and chalcocite. The potentials for oxidation–reduction reactions can be predicted if metastability of elemental sulfur is considered. In the passivated
group of sulfide minerals in which Chander included pyrite, chalcopyrite, and bornite, the
reactions are irreversible. The surfaces of passivated sulfides are normally covered with a
layer of oxidation products. The potential of such surfaces cannot be predicted thermodynamically but are so-called mixed potentials. Chander also pointed out that the collecting
species for sulfides with sulfhydryl collectors such as xanthate can also be divided into two
categories. He observed that metal-xanthate salts form at the surface of reversible sulfides,
TABLE 6
Rest potential and reaction product of sulfide minerals with KEX solutions
Mineral
Sphalerite
Stibnite
Galena
Bornite
Chalcocite
Chalcopyrite
Molybdenite
Pyrite
Arsenopyrite
Rest Potential, V
–0.15
–0.13
+0.06
+0.06
+0.14
+0.16
+0.21
+0.22
+0.22
Reaction Product*
NPI
NPI
MX
MX
NPI (MX)
X2
X2
X2
X2
Adapted from Allison et al. 1972.
*MX = metal xanthate; NPI = not positively identified; X2 = dixanthogen.
Potential for X 2 + 2e – → 2X – at 0.625 mM KEX = 0.13 V.
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22
HISTORICAL ASPECTS OF FLOTATION
100
0.02 mM KEX
(0.014 mM for Cu2S)
pH 9.2
Flotation Recovery, %
80
Cu2S
60
Cu5FeS4
CuFeS2
FeS2
40
20
0
–0.6
–0.4
–0.2
0.0
0.2
Potential, V (vs. SCE)
Adapted from Richardson and Walker 1985.
FIGURE 8 Relationship between flotation recovery and conditioning potential for chalcocite,
bornite, chalcopyrite, and pyrite with KEX as collector at a concentration of 0.02 mM with the
exception of 0.0144 mM for chalcocite (potentials given versus the standard calomel electrode,
or SCE)
whereas dixanthogen is the oxidation product on passivated sulfides, although this remains
an area for further study.
Using in-situ electrodes in microflotation cells, Gardner and Woods (1974) and
Chander and Fuerstenau (1975) were the first to independently and simultaneously demonstrate that the electrochemical potential can control the flotation of sulfide minerals. Using
an electrochemical microflotation cell that incorporated a packed bed of conducting sulfide
particles as the working electrode to correlate interfacial electrochemical reactions with
flotation response, Richardson and Walker (1985) investigated the xanthate flotation of
bornite, chalcocite, chalcopyrite, and pyrite as a function of the electrochemical potential.
The flotation response of these sulfides, shown in Figure 8, is strongly dependent on the
conditioning potential and occurs only under moderately oxidizing conditions. At a certain
reducing potential, depending on the mineral, flotation ceases. Identification of the products
of the sulfide mineral–ethyl xanthate interaction was conducted by linear sweep voltammetry
and ultraviolet (UV) spectroscopy. They found that surface hydrophobicity appears to be metal
xanthates or surface analogs of metal xanthates formed by chemisorption on chalcocite and
bornite, dixanthogen on pyrite because the potential at which flotation begins is that at
which xanthate is oxidized to dixanthogen, and both metal xanthate and dixanthogen (at
higher potentials) on chalcopyrite. Trahar (1984) used chemical means to control the pulp
potential and observed similar behavior in the flotation of sulfide minerals.
In much of the early flotation literature, there are discussions of how low the adsorption
density of collector is for flotation. A subsequent detailed study of the flotation of chalcocite and pyrite using the electrochemical flotation cell, together with determination of the
amount of xanthate collector adsorbed, was carried out by Gebhardt and Richardson
(1987). Their results for the flotation of these two sulfides individually are plotted in Figure 9.
Collection of chalcocite occurs by chemisorption. The plots given in Figure 9 indicate that
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
23
100
14
Flotation Adsorption
Chalcocite
Pyrite
–5
Flotation Recovery, %
5 × 10
M KET, pH 9.2
10
60
8
Monolayer
6
40
4
Adsorption Density, μmol/m 2
12
80
20
2
0
–0.8
0
–0.6
–0.4
–0.2
0.0
0.2
Potential, V (vs. SCE)
Adapted from Gebhardt and Richardson 1987.
FIGURE 9 Flotation recovery and ethyl xanthate adsorption as a function of potential for
single-component chalcocite and pyrite beds at pH 9.2 in an electrochemical flotation cell at a
collector concentration of 0.05 mM
when complete flotation of chalcocite occurs, the adsorption density of xanthate is in the
range of their calculated monolayer, which would be the situation for any chemisorption
process. Flotation of pyrite does not begin until the potential for dixanthogen formation
occurs, and the dixanthogen coating that hydrophobizes pyrite could exceed monolayer coverage. The authors found that separation of mixtures of the two minerals by controlled
potential could be achieved but under conditions different from that for the minerals alone.
Chalcocite dissolution products had a deleterious effect on the chalcocite-pyrite separation
by activating pyrite so that it floated below the potential at which xanthate is oxidized to
dixanthogen, but dissolution products from pyrite had no effect on the flotation of chalcocite.
A C T I VAT I O N A N D D E P R E S S I O N O F S U L F I D E A N D OX I D I Z E D
MINERALS
With the exception of sphalerite and pyrrhotite (Sutherland and Wark 1955), sulfide minerals respond well to flotation with short-chain xanthate collectors. Even methyl xanthate will
float galena, as can be seen from the plots given in Figure 3. Although there is no flotation
with ethyl xanthate at low concentrations, Gaudin (1930) showed that sphalerite exhibits
some response with amyl xanthate and significant flotation with heptyl xanthate, reaching
100% recovery at a collector addition of 0.2 kg/t. M.C. Fuerstenau, Clifford, and Kuhn
(1974) conducted a detailed study of the flotation of sphalerite with short-chain (C2 to C6)
xanthates and found a regular relation between flotation and collector concentration in
solution. Flotation increased sharply once the necessary collector concentration was
reached, indicating zinc xanthate precipitation enters into the process. Fifty percent recovery
was achieved at an ethyl xanthate concentration in solution of 10 mM, whereas only 0.1 mM
hexyl xanthate was necessary to achieve the same degree of flotation. Flotation response correlates exactly with the solubility products of the various zinc xanthate. Most importantly,
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24
HISTORICAL ASPECTS OF FLOTATION
upon the addition of copper salts, sphalerite readily responds to flotation with low additions
of short-chain xanthates. As discovered by Leslie Bradford in 1913 in Australia, this means
of enhancing the floatability of sphalerite has become the classic example of activation in
flotation technology, and since then, the inorganic salt, copper sulfate, has proved to be
indispensable for effectively concentrating zinc sulfide ores by flotation. The general consensus on the activation of sphalerite by Cu(II) ions has been that the process involves an ion
exchange mechanism in which cupric ions are exchanged for zinc ions according to the following reaction, driven by the extremely low solubility product of copper sulfide (CuS):
ZnS ( s ) + Cu 2+ = CuS ( s ) + Zn 2+
(EQ 2)
In his 1930 paper, Gaudin commented that from a geological standpoint, the action of
copper sulfate and similar salts on sphalerite had long been known, with some experiments
having been conducted on the reaction as far back as 1911 by Rogers. The coating on
sphalerite was shown to be covellite. In 1930, Gaudin reported on the results of some experiments conducted to determine the rate of copper uptake by sphalerite and found copper
uptake to be rapid initially, followed by a slow process. He correctly attributed this to slow
diffusion as the copper sulfide coating thickens. About 20 years later, Gaudin again became
interested in this problem and with radioactive 64Cu showed that there indeed is a 1:1 Cu–Zn
exchange, that the first two or three layers rapidly exchange, and that the reaction rate then
continues by solid-state diffusion (Gaudin, Fuerstenau, and Mao 1959). Mellgren and colleagues (1973) measured the heat of adsorption of copper onto sphalerite under controlled
deoxygenated conditions and found the heat of adsorption to be virtually identical to the
heat of formation of CuS. If copper sulfate is added under alkaline conditions, hydrolyzed
cupric species form and are rapidly adsorbed. Because CuS is far more stable than copper
hydroxide, the adsorbed copper hydroxy compounds subsequently transform to CuS and
thereby activate the sphalerite. In 1930, Gaudin also conducted some preliminary experiments on the deactivation of sphalerite with sodium cyanide, and this early interest in deactivation reactions were taken up again by Mao (Gaudin, Fuerstenau, and Mao 1959), who
quantitatively investigated this phenomenon with 64Cu and several copper-complexing
agents.
Through contact-angle measurements and flotation experiments, Wark and Wark
(Sutherland and Wark 1955) examined the activation behavior of sphalerite with a wide
range of metal salts. Their results indicate that successful activators are salts of metals which,
alone or with their sulfides, readily respond to flotation with thiol collectors. Salts of lead,
cadmium, silver, and mercury fall within this category. D.W. Fuerstenau and Metzger (1960)
investigated the adsorption of Pb(II) on sphalerite and clearly showed that the addition of
zinc sulfate can prevent lead activation of sphalerite. In flotation plants, the oxidation of
galena provides sufficient lead ions to activate sphalerite. It is for this reason that the addition of zinc sulfate became standard practice in lead-zinc flotation mills.
In the flotation separation of complex sulfide ores, selective depression of different sulfide minerals is desired. Depressants include such inorganic reagents as hydroxyl ions,
sodium sulfide, sodium cyanide, and alkali sulfites, and their mechanism of action has been
the basis for much discussion through the years (Sutherland and Wark 1955; Gaudin 1957).
In his 1985 paper, Chander interpreted depressant mechanisms in terms of whether the
mineral is a reversible or passivated sulfide (MS). This is schematically shown in Figure 10.
As pointed out by Chander, if the hydrophobic entity at the surface is X2, depression of the
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
Reversible Sulfides
25
Passive Sulfides
1.0
1.0
X2
MX
Eh, V
Eh, V
X
0.0
–
MS
0.0
MS
–1.0
–1.0
0
7
14
pH
Flotation
Species
Depressant
Species
Metal/collector compounds
MX, MX2
0
7
14
pH
Oxidation product of collector
X2
(a) Mineral oxidation product
layer MO, M(OH)n, etc.
(a) Mineral oxidation product
layer MO, M(OH)n, etc.
(b) Hydrophilic layer of mineral–
depressant compound, MD
(b) Hydrophilic MD layer
(c) Removal of surface MX, MX2
by oxidation, reduction, or
hydrolysis
(c) Removal of surface X2
by reduction
Adapted from Chander 1985.
FIGURE 10 Conditions for the flotation and depression of reversible and passivated sulfide
minerals with thiol collectors
passivated sulfide mineral would occur if the mineral oxidizes to form a hydrophilic oxide or
hydroxide layer, if the depressant decomposes X2 by reduction, or if the depressant reacts
chemically or electrochemically to form a hydrophilic metal-depressant (MD) coating. If
MX is the hydrophobic entity, depression of a reversible sulfide mineral would occur if the
mineral oxidizes to a hydrophilic coating, if MX decomposes by oxidation or by reduction,
if MX decomposes by hydrolysis, or if the depressant reacts chemically or electrochemically
to form a hydrophilic coating. Each of the sulfide mineral–collector–depressant systems can
be interpreted in terms of the foregoing mechanisms.
Oxidized heavy-metal minerals such as anglesite (lead sulfate) and cerussite (lead carbonate) require large quantities of sulfhydryl collectors before they respond to flotation.
Comparison of xanthate collector required for the flotation of chalcocite with that for malachite is quite striking. To achieve 50% recovery of 100 × 600 mesh chalcocite with potassium heptyl xanthate (mol wt = 230), a collector addition of 0.0022 mol/t is required,
whereas for similar flotation of 100 × 600 mesh malachite, 4.4 mol/t is required (Gaudin
1957). The results of that same investigation permit comparison for KEX (mol wt = 160) at
20% recovery, namely 0.062 mol/t for chalcocite and 11.2 mol/t for malachite. These huge
differences show the disparity between collector adsorption on chalcocite and chemical
exchange reaction on malachite. Fleming (1952a) quantitatively studied the metathetical
exchange of xanthate for carbonate with cerussite. Gaudin (1957) discussed the results of a
study of the uptake of amyl xanthate by cerussite conducted years earlier, where a coating
several hundreds of ions thick was formed. He commented that because the new phase does
not have much crystal-chemical resemblance to the old solid phase (cerussite), the connection
between the substratum and coating is fragile. To overcome the high collector consumption,
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26
HISTORICAL ASPECTS OF FLOTATION
these oxidized minerals have long been sulfidized before adding the collector, a process that
converts the mineral surface to a sulfide. The most common activator for this purpose is
sodium sulfide (Gaudin 1957). The interaction of sodium sulfide with cerussite was investigated in detail by Fleming (1952b), who observed this to be a metathetical reaction with an
equilibrium constant of about 3 × 1016. Sulfidization is similar to copper activation of
sphalerite in that, initially, the uptake is rapid and the process then continues more slowly by
diffusion as the sulfide layer increases in thickness. In contrast to the Cu–ZnS (copper–zinc
sulfide) system, sulfidized coatings often do not adhere strongly (as noted previously by
Gaudin for the collector coating). Herrera-Urbina, Sotillo, and Fuerstenau (1999) conducted detailed flotation and pulp potential measurements on the sulfidization reaction and
flotation behavior of cerussite with KAX as collector. A summary of their complex findings
can be seen in Figure 11. The effect of adding sodium sulfide to the system manifests itself in
consecutive steps. The first addition of sulfide ions precipitates the lead ions in solution as
PbS particles that coagulate onto the cerussite. After all the dissolved lead has been precipitated, the sulfide concentration in solution reaches that for depression of PbS flotation but is
not yet high enough to initiate the metathetical exchange between carbonate and sulfide at
the cerussite surface. Upon further addition of sodium sulfide, the sulfidization reaction
begins to take place and flotation sharply increases. The surface reaction is rapid until the
PbS coating reaches about 7 monolayers, at which point the residual sulfide ions in solution
again reach the concentration for depression of PbS flotation, causing a pronounced cessation in flotation of the sulfidized cerussite. At this point, the concentration of sulfide ions
(as measured by an ion-specific electrode) in solution builds up sharply while the PbS layer
on cerussite continues to thicken, but at a reduced rate due to diffusion control. Thus, in
industrial flotation practice, the addition of sodium sulfide must be closely controlled,
which can be accomplished by measurement of pulp potential.
100
25
Cerussite
20
Flotation
Monolayers
Residual Sulfide
60
15
40
10
20
5
0
10–6
10–5
10–4
10–3
Number of Monolayers,
Residual Sulfide, M × 10 5
Flotation Recovery, %
80
0.05 mM KAX
5 mM KNO3, pH 9.5
0
10–2
Sodium Sulfide Addition, mol/L
Adapted from Herrera-Urbina, Sotillo, and Fuerstenau 1999.
FIGURE 11 Flotation recovery of cerussite at pH 9.5 with 0.05 mM KAX as collector, the
residual aqueous sulfide in solution, and the number of sulfide monolayers as a function of
added sodium sulfide in 5 mM potassium nitrate open to the atmosphere
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
27
G A S E S I N F L O TAT I O N
Many people involved in flotation considered gas bubbles to be the crux of flotation (Rickard 1916). In 1928, Adams published the results of a detailed study conducted to ascertain
whether the gas constituting the bubbles had any effect on the relative floatabilities of minerals. He carried out extensive investigations with the common sulfide minerals plus graphite and sulfur using a range of gases that included air, hydrogen, oxygen, nitrogen, and
carbon dioxide. He concluded from his experiments that the floatability of minerals
depends on the nature of the gas being used. However, in their extensive contact-angle investigation, Wark and Cox (1934) were somewhat surprised to find virtually the identical contact angle of 60° with ethyl xanthate collector for a range of gases, including air, oxygen,
hydrogen, nitrogen, carbon dioxide, and sulfur dioxide. In the 1950s, Plaksin extensively
studied the effect of nitrogen gases and dissolved gases on the flotation of various minerals,
particularly the effect of the oxygen in water (Plaksin 1959; Klassen and Mokrousov 1963).
Flotation in closed cells with nitrogen as the flotation gas is used industrially in the final separation stage of copper-molybdenum concentrates, the main savings being reduction in
sodium sulfide consumption. For the flotation of an auriferous pyrite that oxidizes so readily
that it rapidly loses its floatability, a nitrogen flotation system was installed at the Lone Tree
mine in Nevada (Simmons et al. 1999).
It was not until 1958, in a study of the effect of chemical reagents on the motion of air
bubbles in water, that D.W. Fuerstenau and Wayman published the first results which clearly
showed that the adsorption of organic surface-active agents (frothers) retarded the velocity
of bubbles. Previous investigations had shown that bubbles in distilled water, before they
begin to distort from spherical shape, rise faster than solids of the same specific gravity due
to circulation within the bubble. By adding a frother (terpineol) to distilled water, the rise
velocity of aged bubbles was found to slow down to that expected if they had been solid due
to the surface tension gradient at the surface set up by the adsorbed frother molecules. This
retardation was similar to that observed earlier for air bubbles in tap water. In distilled water
in the absence of terpineol, solutions containing about 20 mg/L of potassium hydroxide
(KOH), KCl, or KEX did not have any appreciable effect on the rise velocity of bubbles.
However, air bubbles in industrial flotation pulps full of surface-active materials would be
like bubbles in tap water and would not exhibit the rapid rise characteristic of bubbles in distilled water.
The primary role of gas bubbles in flotation, of course, is to attach to hydrophobic mineral particles, levitate them to the surface, and maintain a froth layer that lasts long enough
for effective recovery of the desired mineral particles. This requires organic compounds
(frothers) that create a surface tension gradient at the air–water interface. Long-chained surfactants such as amines, sulfonates, and fatty acids function not only as collectors but also as
frothers. Typical frothers are alcohols (ROH), either straight-chain with 6 to 9 carbon
atoms or branched-chain compounds with 6 to 16 carbons. Pine and eucalyptus oils represent typical cyclic alcohols, with the active molecule being terpineol. In more recent
decades, alkoxy-substituted paraffin-type frothers (such as triethoxy butane) contain no
hydroxyl groups but gain their polarity from ether linkages. Another class of frothers
includes the hydroxylated polyglycol ethers (King 1982). Interest in frothers has increased
because of overall economic factors. For example, Klimpel and Hansen (1988) carried out
considerable work on the effect of frother structure on the selectivity and recovery of minerals and found that with increasing branching, the maximum particle size that can be recovered decreases while at the same time selectivity increases. More research is needed in this area.
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28
HISTORICAL ASPECTS OF FLOTATION
O X I D E A N D S I L I C AT E M I N E R A L S A N D T H E E L E C T R I C A L
D O U B L E L AY E R
The immersion of a mineral into an aqueous solution produces a region of electrical inhomogeneity at the solid–solution interface; namely, an excess (+ or –) charge apparently fixed
at the solid surface is exactly balanced by a diffuse region of ions of equal but opposite
charge (termed counterions). This is the electrical double layer. Useful monographs that discuss the electrical double layer are those of Hunter (1981), King (1982), and Stumm (1992).
Electrical double-layer phenomena have a limited role in the collection of sulfide minerals
but are very important in nonmetallic mineral flotation, particularly with oxides and silicates. Figure 12 is a simple model of the electrical double layer at a mineral–water interface,
showing the charge on the solid surface and the counterions extending as a diffuse layer out
into the aqueous phase. This figure also shows the drop in potential across the double layer.
The closest distance of approach of hydrated counterions to the surface, δ, is called the Stern
plane. The total double-layer potential, or surface potential, is ψo, and the potential at the
Stern plane is ψδ. From the Stern plane out into the bulk of the solution, the potential drops
exponentially to zero. The electrokinetic or zeta potential, ζ, is the potential just outside the
Stern plane where the diffuse layer is able to slip relative to the solid surface. In the case of
ions that directly interact with surface sites, either chemically or by some other strong specific adsorption force, the adsorbed ions may lie closer to the surface at a plane called the
inner Stern plane. However, these discussions simply refer to adsorption in the Stern plane.
Several different parameters that quantify the electrical double layer are useful in interpreting flotation behavior, particularly the selective adsorption of collectors. This includes such
factors as the magnitude of the surface charge, the point of zero charge (PZC) of the mineral, interfacial potentials, thickness of the electrical double layer, specific adsorption of collectors, and ion exchange phenomena (D.W. Fuerstenau and Healy 1972).
The surface charge in systems of importance to flotation may arise in a number of ways.
For example, the surface charge on an oil droplet or an air bubble may result from the
adsorption of long-chained surfactant ions at the oil–water or air–water interface. In the
case of layer-silicate minerals (such as clays and micas), because of substitution of Al3+ for
Si4+ in the silica tetrahedra and Mg2+ for Al3+ in the octahedral layer of the crystal lattice,
the surfaces of these crystal faces carry a constant negative charge that is independent of
Stern
Layer
Surface
Charge, σo
Counterions, σd
δ
ψo
Shear
Region
ψδ
Solid
Potential
ζ
δ
1/κ
Distance
NOTE: Shown are the surface potential, the Stern layer potential, the zeta potential at the
slipping or shear plane, and the potential decrease to zero-out into the bulk solution. The
distance 1/κ is the center of gravity of the diffuse layer of counterions.
FIGURE 12 Simple schematic of the electrical double layer showing the surface charge on the
solid and the counterions adsorbed in the diffuse layer, with their closest distance of approach
being the Stern plane
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
29
solution conditions. For two of the layer-silicate minerals, talc and pyrophyllite, the charge
is exactly balanced within sheets that constitute the layer structure, the faces of these crystals
are uncharged, and the minerals are naturally hydrophobic. In the case of salt-type minerals
such as barite (barium sulfate), fluorite (calcium fluoride), argentite (silver sulfide), iodyrite
(silver iodide), and so forth, the surface charge arises from the preference of one of the lattice
ions for the solid relative to the aqueous phase. Equilibrium is attained when the electrochemical potential of these ions is constant throughout the system. Those particular ions
that are free to pass between both phases and therefore establish the electrical double layer
are called potential-determining (PD) ions. For oxide and silicate minerals, hydrogen and
hydroxyl ions have long been considered to be PD ions (although there remains a difference
in opinion as to how pH controls the surface charge on oxides), and adsorption/dissociation
of surface hydroxyls is generally assumed to be the mechanism of developing a surface charge:
M – OH + H + → M – OH 2+
M – OH + OH – → M – O – + H 2 O
(EQ 3)
If H+ and OH– are the potential-determining ions (PD+Z and express PD–Z are the PD
ions of valence z in more general terms) for an oxide mineral, then the surface charge (σo) is
simply given by
σ o = zF ( Γ H + – Γ OH – )
(EQ 4)
where F is the Faraday constant and the adsorption density (the quantity in the brackets) of the PD ions is in moles per square centimeters. This can be measured by titration of a
suspension of the mineral in water and is generally a complex function of the ionic strength
and the activity of PD ions (pH for oxides) in solution. If the adsorption of the positive PD
ion exceeds that of the negative PD ion, the surface of the mineral is positively charged, and
vice versa if the adsorption of the negative PD ion exceeds that of the positive PD ion. If
adsorption of counterions occurs only because of electrostatic interaction, then the diffuse
layer charge, σd, is oppositely equal to the surface charge, σo. Such counterions are called
indifferent ions. The relationships among surface charge, diffuse layer charge, and ψδ are
given for a symmetrical electrolyte with ions of valence z (where z = z+ = z–) by the Gouy–
Chapman equation, as modified by Stern (Hunter 1981):
σ d = – σ o = ( 8εε o RTC ) 1 ⁄ 2 sin h ( zFψ δ ⁄ 2RT )
(EQ 5)
where ε is the dielectric constant of the liquid, εo is the permittivity constant, R is the gas
constant, T is the temperature, and C is the concentration of indifferent electrolyte in solution. If the net adsorption density of PD ions is zero, the surface of the mineral is uncharged
and the solid is at its PZC.
With regard to flotation behavior, the single most important parameter that describes
the electrical double layer of a mineral in water is its PZC, which is determined by a particular value of the activity, a, of the PD cation of valence z. The surface potential is considered
to be zero at the PZC, and in the case of oxides (if the Nernst equation is assumed to be
valid), its value at any other pH is given by
RT
ψ o = -------- ln [ a ( H + ) ⁄ a ( H + ) PZC ] = 0.059 [ pH PZC – pH ] volt
zF
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(EQ 6)
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30
HISTORICAL ASPECTS OF FLOTATION
The importance of the PZC is that the sign and magnitude of the surface charge has a
major effect on the adsorption of all other ions, and particularly those charged oppositely to
the surface.
Typical values of the PZC for some of the oxide minerals are as follows (D.W. Fuerstenau and Healy 1972; King 1982): quartz (SiO2), pH 2; cassiterite (SnO2), pH 4.5; zirconia (ZrO2), pH 4; rutile (TiO2), pH 5.6; natural hematite (Fe2O3), pH 4.8–6.7; synthetic
hematite, pH 8.6; corundum (Al2O3), pH 9.1; and magnesia (MgO), pH 12. The PZCs of
silicate minerals are affected by crystal chemistry and selective leaching of metal cations or
silica from the surface. Typical values of the PZC of some silicate minerals are as follows
(D.W. Fuerstenau and Raghavan 1980; King 1982): kyanite (Al2SiO5), pH 7.8; zircon
(ZrSiO4), pH 5.8; olivine [(Mg,Fe)2SiO4], pH 4.1; almandine [Fe3Al2(SiO4)3], pH 5.8;
beryl [Be3Al2(Si6O18)], pH 3.4; spodumene [LiAl(SiO3)2], pH 2.6; rhodonite (MnSiO3),
pH 2.8; talc [Mg6(Si8O20)(OH)4], pH 3.6; muscovite [K2Al4(Al2Si6O20)(OH,F)4], pH 1;
and orthoclase [K(AlSi3O8)], pH 1.8.
A useful method for the study of adsorption phenomena in solid–liquid systems is the
measurement of electrokinetic potentials that result from the interrelation of mechanical
fluid dynamic forces with interfacial potentials. In making electrokinetic measurements, the
liquid phase is caused to move relative to the solid phase by the application of a mechanical
force (streaming potential) or by an electric field (electrophoresis). More recently, electroacoustophoresis has also been used to evaluate zeta potentials, particularly with concentrated suspensions. From the appropriate theory, one evaluates the potential at the slipping
plane, which generally is considered to be just outside the Stern plane. The potential at the
slipping or shear plane is the zeta potential, ζ, and is often assumed to approximate the Stern
plane potential, ψδ, although ζ < ψδ, as can be seen in Figure 12. The determination of zeta
potentials has been a powerful tool in delineating flotation adsorption phenomena. Early
on, not everyone believed in discussing electrokinetic phenomena in terms of zeta potentials.
Because of his concern with regard to the equations connecting electrophoretic mobilities to
zeta potentials, the distinguished surface and colloid chemist, Victor LaMer (1967), commented:
It is for these reasons that I feel strongly that no scientific purpose is served by converting mobilities into zeta potentials until the more complicated connecting equations have been verified. Of course if you have something to sell, zeta potential is a
much better advertising catch word than is electrophoretic mobility. The natives
are mystified and admire with great awe the black box which gives the results on
the dials.… This above shows that much of the recent ‘hullabaloo’ about zeta
potentials is meaningless.
The success of interpreting adsorption phenomena in terms and zeta potentials and all of
the examples given by Hunter (1981) proved LaMer to be wrong.
When counterions adsorb in the Stern plane through forces in addition to simple electrostatics, such counterions are considered to be specifically adsorbed. Flotation collectors for
oxide minerals, for example, are such ions. Such phenomena as covalent bond formation,
hydrophobic bonding, hydrogen bonding, solvation effects, and so forth, lead to specific
adsorption. Because of their surface activity, the charge in the Stern plane can exceed the surface charge, and the sign of ψδ is reversed. Actually, indifferent and specifically adsorbed
ions may lie in different planes because of ionic size and hydration. Chemisorbed ions may
lie closer to the surface because they are dehydrated, in what is termed the inner Stern plane.
These discussions will not differentiate between an inner and outer Stern plane. Stern first
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
31
suggested specific adsorption of counterions in terms of a Langmuir-type adsorption equation. For adsorption densities less than about 30% coverage, Grahame (1947) derived a
Boltzmann-type relation, now generally called the Stern–Grahame equation, expressed here
in terms of adsorption-free energies and adsorption densities, rather than in terms of charge
and potentials as performed by Grahame:
o ⁄ RT )
Γ δ = 2rC exp ( – ΔG ads
(EQ 7)
where Γδ is the adsorption density in moles per square centimeters, ΔGadso is the standard
free energy of adsorption, r is the effective radius of the adsorbed ion, and C has to be the
bulk concentration in moles per cubic centimeter. Another form of the relation expressing
adsorption in the Stern plane is given as fractional coverage:
C
Φ o ⁄ RT )
-----------= ---------- exp ( – ΔG ads
55.5
1–Φ
(EQ 8)
In this equation, C is the concentration of adsorbate in moles per liter, and 55.5 is the
number of moles of water in a liter. If ions are adsorbed at the Stern plane only because of
electrostatic interactions, then the standard free energy of adsorption is given by
o
o
ΔG ads
= ΔG elec
= zFψ δ
(EQ 9)
When an ion exhibits surface activity, such as the case for a flotation collector, then the
standard free energy of adsorption has additional terms:
o
o
ΔG ads
= zFψ δ + ΔG spec
(EQ 10)
where ΔGspeco represents the specific interaction terms. These can be comprised of various
contributions:
o
o
o
o
o
ΔG spec
= ΔG chem
+ ΔG ho + ΔG hpb
+ ΔG solv
+ ΔG hpb*
+ ....
(EQ 11)
where the individual terms represent changes in the standard free energy due to chemical
bonding, hydrogen bonding, hydrophobic bonding, and solvation effects, respectively. The
o
term ΔG hpb*
represents the specific adsorption phenomena through surfactant chain interaction with a hydrophobic solid, such as talc or graphite, and would be absent for a hydrophilic mineral, such as quartz. Depending on the mechanisms involved in the interaction of
the collector with the mineral surface, the contributions to the change in adsorption free
energy can be essentially zero or have a finite value. This approach will be used in explaining
the behavior of various types of collectors on nonmetallic minerals. There are two parts to
the usual ionic collector—the charged head group and the hydrocarbon chain—and both
can give rise to specific adsorption effects in the Stern layer.
P H Y S I S O R P T I O N O F F L O TAT I O N C O L L E C T O R S
As first clearly defined by Taggart, the molecular structure of chemical flotation collectors
requires that it has a polar group and a nonpolar hydrocarbon chain. The polar head group
may react chemically with metal sites at the mineral surface or it may adsorb merely because
of the sign of its electrical charge. The former type of interaction is chemisorption, as exemplified
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32
HISTORICAL ASPECTS OF FLOTATION
by xanthates on sulfide minerals and gold, whereas the latter type of interaction is one of
physical adsorption. In this case, such factors as the size of the head group, its electrical
charge, and whether the head group hydrolyzes are important. Of particular importance is
the alkyl chain of the collector given that it enters into the process because of hydrophobic
bonding phenomena.
When surfactant ions or molecules in solution become sufficiently concentrated, they
aggregate through association of their hydrocarbon chains into clusters containing about 50
to 80 surfactant ions called micelles. The charged heads orient outward to the solution so
that the chains are effectively removed from water. The driving force is as a consequence of
the desire for water molecules to hydrogen-bond with themselves, which results in a free
energy decrease of about 1 kT per CH2 group (or 1 RT or 0.6 kcal/mol) removed from
water (where k is the Boltzmann constant, R is the gas constant, and T is the absolute temperature). Thus, the number of ions in the micelle and the concentration at which micelles
form, the critical micelle concentration or CMC, depends on the number of carbon atoms
in the hydrocarbon chain of the surfactant. Based on his findings of how the adsorption of
dodecylammonium ions affected the zeta potential of quartz, in 1953 D.W. Fuerstenau first
proposed that similar phenomena can take place at the mineral–water interface. In their
1955 paper, Gaudin and Fuerstenau termed these surface aggregates hemimicelles because
the charged heads would be oriented toward the mineral surface (at least until the zeta
potential is reversed). As the adsorption density increases in the Stern plane, the adsorbed
surfactant ions come sufficiently close together that they begin to associate into twodimensional aggregates similar to micelle formation in bulk solution. When hemimicelles
begin to form, the free energy of adsorption is given by
o
o
ΔG ads
= zFψ δ + ΔG hpb
= zFψ δ + Nφ
(EQ 12)
where N is the number of CH2 groups in the hydrocarbon chain, and φ is the free energy
change on removal of one mole of CH2 groups from water. Through these same chain association effects, organic molecules (such as alcohols) can co-adsorb with adsorbed organic
ions. The plots given in Figure 2 illustrate the effect of hemimicelle formation on a number
of interfacial properties of quartz in aqueous solutions of DAA at pH 6–7, conditions under
which the quartz surface itself is negatively charged. At about 10–4 M DAA, there are abrupt
changes in the amount of aminium ions adsorbed and the zeta potential becomes sharply
positive. This is because the hydrophobic bonding contribution to the adsorption free
energy dominates, that is, Nφ > zFψδ . It is also seen that the onset of rapid flotation occurs
under conditions where hemimicelles begin to form, clearly indicating that good flotation
depends on strong collector adsorption in the Stern plane. If the driving force for the
adsorption of physisorbing surfactants were only electrostatic, flotation with such collectors
would be limited.
Careful delineation of surfactant adsorption phenomena under controlled ionic
strength was first conducted by Somasundaran and Fuerstenau (1966), and continued by
Wakamatsu and Fuerstenau (1968). Combining zeta potential measurements with adsorption phenomena in the alumina–sodium dodecylsulfonate system clearly showed the existence of a three-step isotherm, which later was shown to be a four-step isotherm when
surfactant concentrations were taken above that used in flotation, namely, all the way up to
the CMC. This so-called S–F isotherm can be illustrated in terms of the results obtained for
the adsorption of sodium dodecylsulfonate on alumina at pH 7.2 (Figure 13). The plots
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
33
9
7
12
pH
6
5
14
PZR
4
pH for C12
No. C for pH 7.2
3
Number of
Carbon Atoms
10
8
16
2
18
10.0
III
80
II
Contact Angle, degrees
1.0
Adsorption Density, μmol/m 2
–40
70
I
0.1
60
–20
50
40
30
0
20
10
0.01
Zeta Potential, mV
Contact Angle
Adsorption Density
Zeta Potential
20
0
40
Alumina
2 mM Ionic Strength
pH 7.2
0.001
10–5
10–4
10–3
10–2
Sodium Dodecylsulfonate Concentration, M
NOTE: Because the maximum contact angle is reached at the PZR, the dependence of the
PZR on alkylsulfonate chain length at pH 7.2 and on pH for dodecylsulfonate are given in the
upper portion of this figure.
Adapted from D.W. Fuerstenau and Wakamatsu 1973.
FIGURE 13 For alumina at pH 7.2 and 2 mM ionic strength, the effect of sodium
dodecylsulfonate on the collector adsorption density, zeta potential, and equilibrium contact
angle as a function of reagent concentration, showing three distinct adsorption regions
given in Figure 13 show the three regions of adsorption, which can be interpreted in terms
of the Stern–Grahame equations. In Region I, the sulfonate ions adsorb individually by electrostatic interaction and ion exchange with chloride ions, and the zeta potential is therefore
constant given that excess adsorption in the Stern plane is absent. Region II is characterized
by a sharp change in the zeta potential and a sharp increase in the adsorption of sulfonate
ions due to hemimicelle formation through the onset of the hydrophobic bonding contribution to specific adsorption. The boundary between regions II and III occurs precisely at the
concentration where the zeta potential reverses (the PZR), at which point the electrical
repulsion in the Stern layer begins to act against the hydrophobic bonding forces that are
responsible for continued sulfonate adsorption. In Figure 13, the equilibrium contact angle
(virtually identical to the liquid-receding contact angle) of an air bubble on alumina is also
plotted as a function of the equilibrium sulfonate concentration in solution (after the results
presented by D.W. Fuerstenau and Wakamatsu 1973). At the PZR, in all cases the contact
angle reached its maximum and remained such as the concentration of sulfonate was
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34
HISTORICAL ASPECTS OF FLOTATION
C18
C16
C14
C12
C10
C8
C6
C4
Quartz
100
Flotation Recovery, %
pH 6–7
RNH3 Acetate
80
60
40
20
0
10–7
10–6
10–5
10–4
10–3
10–2
10–1
100
Alkylammonium Acetate Concentration, M
Adapted from D.W. Fuerstenau, Healy, and Somasundaran 1964.
FIGURE 14 Flotation recovery of quartz as a function of the concentration of alkylammonium
acetates of various chain lengths at pH 6 to 7
increased. One interpretation is that after the PZR, with continued adsorption, some of the
surfactant ions adsorb in reverse orientation. This behavior was observed for sodium dodecylsulfonate at different pH values and also for alkylsulfonates of different chain lengths at
pH 7.2. These results are summarized in the upper portion of Figure 13, which shows the
concentration where the maximum contact angle (the PZR) is reached for sulfonates of different chain lengths at pH 7.2 and for dodecylsulfonate at different pH values. This shows
that the amount of collector can be reduced by increasing the alkyl chain length or by reducing the pH (alumina is positively charged below pH 9).
Because the formation of hemimicelles depends on the length of the hydrocarbon
chain, the flotation of quartz with alkylammonium salts should exhibit a regular dependence on chain length, similar to the well-known Traube rule. Figure 14 presents the flotation of quartz with amine collectors ranging from 4 to 18 carbon atoms. Because of the Nφ
term, collectors having a longer hydrocarbon chain (greater N) adsorb more strongly and
function effectively as flotation reagents at more dilute concentrations. It is interesting to
note that a 4-carbon amine requires about 1 molar residual concentration for complete flotation of quartz, which is in sharp contrast to such strongly chemisorbing sulfhydryl collectors as ethyl xanthate on sulfide minerals. The role of the hydrocarbon chain in galena
flotation with carboxylic acids is apparent by the results of Gaudin et al. (1928), given in
Figure 1.
The role of double bonds of flotation collectors has been the subject of numerous investigations and that will be presented when chemisorption phenomena is discussed later in
this chapter.
E L E C T R O S TAT I C M O D E L O F F L O TAT I O N
During the period from 1953 to 1956, D.W. Fuerstenau began to develop the concept that
flotation collectors which physically adsorb must function as counterions in the electrical
double layer, and that oxide mineral flotation with physisorbed anionic collectors should be
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
35
150
NaCl, M
10–5
10–4
10–3
10–2
10–1
PZC
Zeta Potential, mV
100
50
0
–50
Corundum
100
Flotation Recovery, %
80
60
40
Corundum
4 × 10–5M, C12
RNH3Cl
RSO4Na
RSO3Na
20
0
0
2
4
6
8
10
12
14
pH
NOTE: The upper curve gives the zeta potential of corundum as a function of pH for a range
of sodium chloride concentrations showing the PZC. The lower curve represents the flotation
response of corundum with 0.04 mM SDS, sodium dodecylsulfonate, and DAC as a function
of pH, showing the dependence of flotation on the PZC and the collector.
Adapted from Modi 1956; Modi and Fuerstenau 1960.
FIGURE 15
Dependence of the flotation of corundum on surface charge
appreciable only at pH values below the PZC and with cationic collectors only at pH values
above the PZC. These concepts, which have been termed the electrostatic model of flotation,
are briefly summarized here. Experiments to confirm these ideas were conducted by H.J.
Modi as part of his doctoral thesis at MIT (Modi 1956). He determined the PZC of corundum by electrokinetics and conducted Hallimond tube flotation experiments with a variety
of physisorbed collectors as a function of concentration and pH. Figure 15 presents the
results of the very first experiments carried out to test these concepts, which were first presented in Modi’s doctoral thesis in 1956 and then published by Modi and Fuerstenau in
1960. The upper part of Figure 15 presents the zeta potential of corundum (synthetic sapphire), determined by streaming potential measurements as a function of pH with various
additions of sodium chloride as indifferent electrolyte. All the curves intersect and cross at
zero zeta potential at about pH 9, which is the PZC of this material. Flotation experiments
were conducted at a solution concentration of 4 × 10–5 M with three different high-purity
collectors: dodecylammonium chloride (DAC), sodium dodecylsulfate (SDS), and sodium
dodecylsulfonate. The lower part of Figure 15 clearly shows that corundum responds to
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36
HISTORICAL ASPECTS OF FLOTATION
PZC
Corundum
Goethite
Ilmenite
Quartz
SDS DAC
pH 9.1
pH 6.7
pH 5.6
pH 2.4
10–4 M
Flotation Recovery, %
100
80
60
40
20
0
–3
–2
–1
0
1
2
3
4
pH–pHPZC
Adapted from D.W. Fuerstenau and Herrera-Urbina 1989.
FIGURE 16 Effect of the PZC on the flotation of several different oxide minerals with
physisorbing anionic (SDS) and cationic (DAC) collectors
anionic collectors at pH values below the PZC where corundum is positively charged, and
to the cationic collector at higher pH values where corundum is negatively charged. The
pKa of dodecylamine is 10.4; at pH 10.4, the DAC in solution will be 50% aminium ions
and 50% amine molecules. Under these conditions, flotation is maximal because of coadsorption of aminium ions and amine molecules. As the pH is raised to about 12, flotation
drops sharply and ceases at pH 12.6. Under these conditions, there are insufficient aminium
ions to bind the collector to the surface. This upper pH limit for flotation with primary
amine collectors is virtually universal. Experiments with sodium dodecyl xanthate showed
that this reagent also functions as a physisorbed collector for corundum in a manner similar
to any other anionic long-chained surfactant.
After returning to the University of Minnesota upon completing his doctorate at MIT
in 1957, Iwasaki and several co-workers carried out similar detailed research on the flotation
of a number of iron ore minerals with high-purity reagents (Iwasaki, Cooke, and Columbo
1960; Iwasaki, Cooke, and Choi 1960). The flotation response of oxide and silicate minerals
to these types of collectors is characteristically similar to that presented in Figure 15, as can
be seen from the plots given in Figure 16, which is a composite drawing showing the flotation response of four different oxides whose PZCs range from pH 2 to pH 9 with DAC and
SDS as collectors. Two factors are involved in the electrostatic model of flotation: adsorption on a surface oppositely charged to the collector and a hydrocarbon chain sufficiently
long to help anchor the physisorbed collector.
In an excellent investigation of the flotation of iron oxides with 12- and 18-carbon collectors, Iwasaki, Cooke, and Choi (1960) showed that the flotation response of hematite
with 12-carbon alkyl amines and sulfates is that predicted by the PZC. Their results, given
in Figure 17, show that the 18-carbon surfactants continue to function as collectors at about
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
37
+4
Mobility,
μm/sec/V/ cm
+2
0
–2
Hematite
–4
100
Flotation Recovery, %
RNH3
RSO4
80
10–4 M
C12
C12
60
C18
40
C18
20
0
0
2
4
6
8
10
12
14
pH
Adapted from Iwasaki, Cooke, and Choi 1960.
FIGURE 17 Influence of hydrocarbon chain length on the flotation of hematite as a function of
pH with alkyl sulfate and aminium collectors having 12 (open symbols) and 18 (solid symbols)
carbon atoms. The upper portion of this figure gives the electrophoretic mobility of the
hematite showing that its PZC occurs at pH 6.7
3 pH units above or below the PZC, respectively. This means that hydrophobic bonding
phenomena must be sufficient to overcome the electrical repulsion. Because each CH2
group contributes about 1.1 kT to the energy of hydrophobic bond formation, the increased
contribution to the free energy of adsorption on lengthening the alkyl chain from 12 to 18
carbons is about 6.6 RT/mol. As an approximation, the potential at the surface increases
177 mV (7 RT) when the pH is changed by 3 units. The potential in the Stern plane would
be less, but that potential is equivalent to 7 RT. Unfortunately, similar systematic flotation
studies were never conducted with intermediate chain lengths in this system.
C H E M I S O R P T I O N O F C O L L E C TO R S O N OX I D E S A N D S I L I C AT E S
If counterions are adsorbed only through such forces as electrostatic attraction and hydrophobic bonding (association between the hydrocarbon chains), the process is termed physical adsorption or physisorption. If the surfactant forms covalent bonds with metal atoms in
the surface, then the process is called chemical adsorption or chemisorption. As already discussed, examples of physical adsorption in mineral–water systems include alkylammonium
ions on quartz and other oxide minerals, and alkyl sulfates and sulfonates on alumina. Conditions can be such that lattice ions are displaced from their lattice positions by the adsorbate, giving rise to surface reaction (such as the uptake of xanthate by cerussite). Examples of
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38
HISTORICAL ASPECTS OF FLOTATION
chemisorption of collectors on oxide and nonmetallic minerals include the oleate–hematite,
hydroxamate–hematite, hydroxamate–MnO2, and oleate–fluorite systems. Infrared spectroscopic studies have helped confirm the existence of calcium oleate, ferric oleate, ferric
hydroxamate, and Mn-hydroxamate species on the surface, indicating chemisorption by
covalent bond linkages. Chemisorption is attractive for several reasons: (1) flotation selectivity should be enhanced if the collector ions/molecules bind to surface sites on a single
mineral; (2) there is reduced reagent consumption because of lower residual solution concentrations; and (3) fine particles should respond to flotation if the mineral surface is not
highly charged and if the collector is not adsorbed on the air bubbles because of the collector having a shorter hydrocarbon chain or to residual collector concentrations in solution
being more dilute.
In nonmetallic mineral flotation, a commonly used collector is oleic acid, which has
received considerable attention through the years. The results of an excellent and interesting
investigation of oleic acid interaction with a variety of oxide and complex silicate minerals
was presented by Polkin and Najfonow in 1964. In addition to flotation tests, they determined the amount of collector adsorbed with 14C-marked oleate, delineated chemisorption
reactions with infrared spectroscopy, conducted leaching studies to remove surface metal
ions, ascertained the effect of reagents and leaching on the zeta potentials of the minerals,
and considered various regulating or modifying reagents to increase selectivity. Figure 18
presents their results for the effect of pH on the flotation recovery of nine different minerals
with 1 kg/t of oleic acid as collector. This figure shows that the flotation response of these
minerals to oleate collector is approximately the same (except for albite, a feldspar). Experiments with radioactively marked oleate showed the formation of durable multilayers on the
eight minerals that readily float. They found that pretreatment of some of the minerals with
acids provided a means for achieving selective flotation because various polyvalent metal
100
Oleic Acid
90
1kg/t
5
80
Flotation Recovery, %
70
6,7
60
1
50
2
40
5
6,7
30
20
8
9
3
10
3
1
2
4
9
0
1
2
3
4
5
6
7
8
9
10
11
12
13
pH
Adapted from Polkin and Najfonow 1964.
FIGURE 18 Similarity in the influence of pH on the flotation of a wide variety of minerals with
1 kg/t oleic acid: (1) columbite, (2) zircon, (3) tantalite, (4) ilmenite, (5) rutile, (6) garnet, (7)
tourmaline, (8) albite, and (9) perovskite
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
39
ions were leached or removed from the surface of some of the minerals, which thereby
removed the means by which oleate attached to the mineral. The metal ions at the mineral
surface have a major role in chemisorption of oleate collectors, which will be discussed later
in this chapter.
The role of chemisorption in the flotation of hematite has received considerable attention through the years. Reagents, such as fatty acids and soaps, chemisorb on hematite, as is
readily seen by flotation and collector adsorption occurring several pH units above a mineral’s PZC. Also, the shift in the zeta potential as a function of pH in the presence of soaps
clearly shows strong adsorption until the pH is 2 or 3 units above the PZC. The first
detailed study of chemisorption in the oleate–hematite system was the infrared (IR) spectroscopic investigation by Peck, Raby, and Wadsworth (1966). They found the appearance
of a carbonyl band with maximum absorbance between 1,520 and 1,530 cm–1 that was the
result of the formation of a surface cation–collector anion compound, with the cation
bonded to the mineral structure and the anion to the cation. They also conducted flotation
experiments as a function of pH, using 25 g of –65 mesh specular hematite with 6.3 mg
(2.2 × 10–5 mol) of oleic acid as collector in 275 mL of water, giving a starting collector concentration of 8 × 10–5 M (the residual solution concentration is unknown). Their results are
plotted in Figure 19 as a function of pH, the peak in absorbance is 0.30. Both the maximum
in flotation and IR absorbance occur at pH 7.9. With a titration method, Peck and colleagues determined the PZC of their specular hematite sample to be pH 7.7. They proposed
that the reactions of hematite with oleic acid can be expressed by the following equation:
M-OH + HOl → M-OH···HOl
M-OH···HOl → M-Ol + H 2 O
(EQ 13)
where HOl is oleic acid, M-OH are uncharged surface hydroxyls, and M-Ol are mineral surface sites with chemisorbed collector.
100
80
Hematite
Oleate
Relative Acid-Soap Concentration
Relative Infrared Absorbance
Flotation Recovery, %
Flotation
Infrared Absorbance
Flotation
Acid Soap
60
40
20
0
4
5
6
7
8
9
10
11
pH
FIGURE 19 Correlation of hematite flotation recovery and infrared absorbance of adsorbed
oleate as a function of pH (data from Peck, Raby, and Wadsworth 1966) and correlation of the
flotation of hematite with the concentration of oleate acid-soap concentration (data from
Kulkarni and Somasundaran 1980)
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40
HISTORICAL ASPECTS OF FLOTATION
Noting that there is almost a universal pH at which a wide range of minerals appear to
respond to flotation with oleic acid, Kulkarni and Somasundaran (1980) proposed that it is
the acid soap which is responsible for the strong adsorption of oleic acid or oleate soap in
slightly alkaline environments. They carried out detailed analysis of the composition of oleic
acid solutions and conducted a wide variety of adsorption and flotation experiments with
hematite. Titration experiments showed the PZC of their hematite sample to be pH 7.1.
Also in Figure 19, their results for the Hallimond tube flotation of hematite with 3 × 10–5 M
potassium oleate and 8 × 10–5 M potassium nitrate are plotted together with the calculated
concentration of the acid-soap dimer. The maximum in their plot corresponds to 4.8 × 10–8
M acid-soap dimer concentration.
From the results of a series of investigations on chemisorption in nonmetallic mineral
flotation, M.C. Fuerstenau and his colleagues (M.C. Fuerstenau and Palmer 1976) found
correlations between the flotation response and the pH at which metal ions at the surface of
the mineral hydrolyze. Figure 20 presents the results obtained by Palmer, Fuerstenau, and
Aplan (1975) for influence of pH on the flotation of chromite with sodium oleate as collector. Chromite ideally is FeO·Cr2O3, but isomorphous substitution of Mg(II) for Fe(II) and
Fe(III) for Cr(III) generally occurs in nature. The chromite used by Palmer and colleagues
assayed 41.7% Cr(III), 8.0% Al2O3, 3.7% Fe(III), 7.1% Fe(II), 8.% Mg(II), plus some minor
amounts of other elements. The two flotation peaks in the vicinity of pH 8 and pH 11
match the hydrolysis peaks of FeOH and MgOH, respectively. The peak at about pH 4 is
most likely due to physisorption of oleate anions on positively charged chromite, given that
the PZC occurs at about pH 7. Cr and Al probably do not participate in the surface hydrolysis reactions since Cr and Al are coordinated octahedrally, whereas the divalent cations are
coordinated tetrahedrally with oxygen.
Reaction with chemically hydrolyzed cations at the surface must differ from the chemisorption concept of Peck, Raby, and Wadsworth (1966) in that the hydrolyzed cation is
probably dislodged from its lattice site before reacting with the oleate ion. Hydrolysis of the
metal ion could free the metal ion from its lattice site and make it available for surface reaction. The interaction of both chrysocolla and hematite with K octylhydroxamate leads to a
100
Flotation Recovery, %
Chromite
80
60
40
1 × 10–4 M
5 × 10–5 M
20
Oleate
0
0
2
4
6
8
10
12
14
pH
Adapted from Palmer, Fuerstenau, and Aplan 1975.
FIGURE 20
Flotation of chromite as a function of pH and oleate concentration
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
41
change in color of the mineral surface, indicating that the surface hydroxamate product is
quite thick and suggesting that a surface reaction takes place, building up multilayers of
product. If a reagent chemisorbs, it can only reach monolayer coverage because the chemisorbing species interacts with a surface site. If the adsorbing species combines with a lattice
ion, the species dislodges the ion from the lattice; this process is termed adsorption with surface reaction. A surface reaction can take place by reagents reacting with hydrolyzed species,
as shown in the following reactions.
Chemisorption
M 2+ + OH – → M 2+ OH –
M 2+ + X – → M 2+ X –
Adsorption with surface reaction
M 2+ + OH – →
M 2+ + X – →
| ( MOH ) +
| ( MX ) +
Surface reaction
| ( MOH ) + + OH – →
| ( MOH ) + + X – →
| ( MX ) + + X – →
|M ( OH ) 2
| ( MX ) +
|MX 2
T H E N AT U R E O F T H E H Y D R O C A R B O N C H A I N I N FAT T Y A C I D
F L O TAT I O N
The role of double bonds of flotation collectors has been the subject of numerous investigations. Probably the most detailed investigation of the nature of the hydrocarbon (other than
chain length) in flotation has, perhaps, been that of Cooke, Iwasaki, and co-workers at the
University of Minnesota for fatty acid–iron ore flotation systems. In particular, they were
concerned with the degree of unsaturation in the chain of 18-carbon fatty acid collectors
and carried out extensive investigations with elaidic, oleic, linoleic, linolenic, and stearic acid
as collectors for iron oxide minerals (Iwasaki, Cooke, and Choi 1960). For hematite, the
degree of effectiveness at room temperature followed the order: elaidic > oleic > linoleic >
linolenic. At room temperature (25°C), contact angles on hematite with 3 × 10–5 M fatty
acid at pH 6 were found to be in degrees: stearic, 81; elaidic, 90; oleic, 86; linoleic, 80; and
linolenic, 75. Because stearic acid has limited solubility at room temperature, the researchers
also measured contact angles at 70°C, where stearic acid has appreciable solubility, and
obtained the following results: stearic, 103; elaidic, 91; oleic, 88; linoleic, 81; and linolenic,
80. The greater the degree of unsaturation in the alkyl chain, the greater the degree with
which water molecules interact with the chains and, hence, the less is their surface activity.
Therefore, in a chemisorbing system, the hydrocarbon chain also plays a significant role.
A very systematic study of the double bonds in the flotation of rutile was conducted by
Purcell and Sun (1963). This included determination of zeta potentials by means of streaming potential measurements and flotation response with a Hallimond tube. Because of the
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42
HISTORICAL ASPECTS OF FLOTATION
thoroughness and rigor of their measurements, Figure 21 was prepared from their results for
0.1 mM reagent concentration to illustrate how double bonds in oleate, linoleate, and linolenate affected flotation and the zeta potential of rutile. The PZC of their rutile sample was
pH 6.7. Chemisorption of the soaps causes the zeta potential to become much more negative than in sodium chloride solutions. As the pH is raised, linolenate joins the sodium chloride (NaCl) zeta potential curve at about pH 8, and the linoleate and oleate curves join the
NaCl curve at still higher pH values. It is at this point that the electrical repulsion of the
charged surface overcomes the adsorption tendency of the soap ions. Linolenate with three
double bonds can interact with water molecules more frequently and hence is repelled from
the surface at a lower pH. The lower portion of Figure 21 shows that the effect of pH on flotation response of rutile with these three collectors correlates exactly. Interestingly, in highly
100
Rutile
NaCl, 0.1 mM
Sodium Oleate
Sodium Linoleate
Sodium Linolenate
Zeta Potential, mV
50
0
–50
–100
100
Flotation Recovery, %
80
60
40
Rutile
20
Sodium Oleate, 0.1 mM
Sodium Linoleate
Sodium Linolenate
0
0
2
4
6
8
10
12
pH
Adapted from Purcell and Sun 1963.
FIGURE 21 Flotation of rutile in 0.1-mM sodium oleate, linoleate, and linolenate as a function
of pH to show the influence of double bonds. The upper portion shows the zeta potential of
rutile in 0.1-mM solutions of these chemisorbing collectors and also NaCl, which shows that
the PZC of this rutile sample occurs at pH 6.7.
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
43
acidic solutions, the flotation response of rutile to the three collectors is identical, with flotation sharply decreasing to zero at about pH 1. The three soaps form fatty acids that behave
the same.
At this point, brief mention is made of a planned attempt to utilize hydrocarbon chain
configuration to effect flotation separations. Schulman and Smith (1953) found that a
branched-chain fatty acid collector permitted separation of a cobalt mineral from a copper
mineral, but not with a straight chain collector. Specifically, after first investigating force/
area curves for monolayers with metal salts in the substrate, they found that both chalcopyrite and carrolite float together with caprylic acid (a carboxylic acid with a 7-carbon alkyl
chain), whereas with 2-ethyl hexanoic acid, only chalcopyrite floats.
R O L E O F E X P E R I M E N TA L M E T H O D O L O G Y I N F L O TAT I O N
FINDINGS
Figure 20 also shows that the collector concentration used in conducting an experiment can
mask results. For example, the dual peaks in the alkaline pH region for chromite flotation
are very apparent when the collector concentration is 5 × 10–5 M, but at 1 × 10–4 M, oleate
hydrophobicity is great enough to swamp the reduced floatability at pH 10.
Figure 22 illustrates how flotation time with the Hallimond tube can lead to different
insights into the flotation behavior of a mineral. In their experiments, Iwasaki, Cooke, and
Choi (1960) floated 100 × 150 mesh hematite with 10–4 M oleic acid for 5 minutes in a
Hallimond tube. As can be seen from the plot shown in Figure 22, this long period of flotation yields 100% recovery between the lower and upper pH limits. M.C. Fuerstenau,
Harper, and Miller (1970) floated 65 × 100 mesh hematite for 45 seconds in a Hallimond
tube. They investigated conditioning time and found 10-minute conditioning yielded
somewhat enhanced recovery over that obtained after 3 minutes of conditioning. Their
results for 10-minute conditioning with 10–4 M potassium oleate are also shown in Figure 22.
100
Flotation Recovery, %
80
60
40
Hematite
Oleate
20
Iwasaki
M.C. Fuerstenau
Somasundaran
0
0
2
4
6
8
10
12
14
pH
FIGURE 22 Illustration of how time for Hallimond tube flotation can accentuate or mask
various aspects of the results, based on the flotation of hematite by various investigators
(5 minutes by Iwasaki, Cooke, and Choi [1960]; 45 seconds by M.C. Fuerstenau, Harper, and
Miller [1970]; and 10 seconds by Kulkarni and Somasundaran [1980])
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44
HISTORICAL ASPECTS OF FLOTATION
By using this shorter flotation time, the decrease in hematite flotation between pH 5 and
pH 6 is very pronounced, followed by an increase again as the pH is decreased to below 5.
This second flotation region is due to the physisorption of oleate ions on the positively
charged mineral. On the other hand, Figure 22 also includes the results of Kulkarni and
Somasundaran (1980), who floated hematite with 3 × 10–5 M potassium oleate as collector
in a modified Hallimond tube for a flotation time of 10 seconds. This reduced collector
concentration and very short flotation time yields only the optimum pH for flotation. Thus,
in conducting flotation chemistry research, close attention should be given to the effect of
the range of such variables as reagent concentration, conditioning, and flotation time in
order to be certain that various effects are not masked.
C O M PA R I S O N B E T W E E N P H Y S I C A L A N D C H E M I C A L
ADSORPTION PROCESSES
In cases involving the physical adsorption of flotation collectors, one observes that flotation
is strongly controlled by the PZC of the mineral. In other words, flotation takes place when
the collector is ionic and when the mineral and collector are oppositely charged. This means
that electrokinetic measurements can quite readily delineate conditions for flotation with
physisorbing collectors. On the other hand, with chemisorbing collectors, the active species
can be an ion (as already shown for oleate) or can be a neutral molecule, as will be shown for
hydroxamic acid (a chelating agent). Moreover, when an anionic collector chemisorbs, it can
adsorb on a negatively charged mineral surface until the surface potential is made sufficiently
negative to prevent adsorption (by increasing the pH above the PZC in the case of oxides).
A detailed investigation of the behavior of manganese dioxide provides a comparison of
the complicated nature of physical and chemical adsorption in flotation processing (D.W.
Fuerstenau and Pradip 1984). The PZC of this manganese dioxide (gamma MnO2) occurs
at pH 5.6. Figure 23 (top, middle, and bottom) shows the effect of pH on the Hallimond
tube flotation response of this oxide at three concentrations of three different collectors,
namely sodium dodecylsulfonate, potassium octyl hydroxamate, and sodium oleate, respectively. The results shown in Figure 23 (top) indicate that the flotation of manganese dioxide
with the anionic sulfonate behaves as expected for a physisorbing collector. Flotation only
occurs when the pH is decreased below about pH 6, which happens in conditions where the
mineral carries a positive surface charge and hence adsorbs the anionic sulfonate ions as
counterions. Lower pH values (higher positive surface charge) are necessary for initiating
flotation at lower collector concentrations.
Figure 23 (middle) shows the effect of pH on flotation with a chelating agent (hydroxamate), which strongly coordinates with manganese ions at the mineral surface. Although
chelating agents have been investigated as flotation reagents for 60 years or more (Gutzeit
1946; Marabini, Cases, and Barbaro 1989; Somasundaran and Nagaraj 1984; D.W. Fuerstenau, Herrera-Urbina and McGlashan 2000), hydroxamates have received, by far, more
attention than any other single chelating agent. Coordination of the nitrogen-oxygen atoms
to manganese takes place as shown schematically:
Mn OH +
O
C
R
HO
N
H
O
C
R
O
N
H
Mn
+ H2O
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
45
100
MnO2
PZC
Flotation Recovery, %
80
10–6 M SDS
10–5
10–4
60
40
20
0
100
MnO2
Flotation Recovery, %
80
1 × 10–5 M HXm
1 × 10–4
3 × 10–4
60
40
20
0
100
MnO2
Sodium Oleate
Flotation Recovery, %
80
5 × 10–5 M
1 × 10–4
5 × 10–4
60
40
20
0
0
2
4
6
8
10
12
14
pH
Adapted from D.W. Fuerstenau and Pradip 1984.
FIGURE 23 Influence of pH and collector type on the flotation of manganese dioxide at the
various reagent additions (top: physisorbing SDS; middle: chemisorbing potassium octyl
hydroxamate; bottom: chemisorbing/physisorbing sodium oleate)
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46
HISTORICAL ASPECTS OF FLOTATION
In this chemisorbing system, flotation takes place at pH values where the surface of the
mineral is highly negatively charged. The maximum in floatability occurs at about pH 9,
similar to the case for hematite, which is approximately the pKa of hydroxamic acid. Thus,
optimum flotation appears to take place where the neutral molecules and hydroxamate
anions are about equal, indicating that both coadsorption and surface hydroxylation may be
responsible for the pronounced chemisorption at pH 9. As the pH is raised above 9, flotation decreases because the surface is so highly negatively charged that the hydroxamate
anions are repelled.
The flotation response of manganese dioxide with oleate collector is particularly complicated, similar to the behavior of hematite, with two flotation peaks being present in this
system. The peak in the alkaline region occurs under conditions where the solid surface is
highly negatively charged (PZC = pH 5.6); therefore, strong chemical adsorption forces
must be operative, typical of oleate–mineral systems previously discussed. Under these conditions, the carboxyl ion may be chemically binding to Mn surface sites or the acid-soap
dimer could be extremely surface-active. In the vicinity of pH 5, there is little flotation of
manganese dioxide with oleate. However, as the pH is lowered even more, a second flotation
maximum occurs. This must now be the result of physical adsorption of oleate anions on a
positively charged surface. Since the pKa of oleate is about pH 4.7, the actual adsorbing species must be oleate ions together with oleic acid molecules. At pH values less than about 3,
most of the oleate has been transformed to molecular oleic acid, which is not the reactive
species, and flotation ceases (similar to the case for amines at high pH). As can be seen from
the plots for oleate in Figure 23, the total concentration of oleate in the system must be sufficiently high to provide enough oleate ions for adsorption at low pH values. Liquid oleic
acid droplets form at low pH values and higher oleate concentrations, but that will not be
taken into account in this chapter.
M.C. Fuerstenau, Harper, and Miller (1970) compared the flotation of finely ground
hematitic ore using octylhydroxamate as collector with using oleic acid as collector. In a
cited example, the ore was ground to 70% minus 15 μm, and an addition of 0.2 kg/t of
hydroxamate collector resulted in a final concentrate recovery of 86% at a grade of 64% Fe.
The ability of a chelating agent such as hydroxamate to successfully float fine particles has
much to do with the fact that the bubbles will not be highly charged.
A C T I VAT I O N I N N O N M E TA L L I C M I N E R A L F L O TAT I O N
Activation in nonmetallic mineral (oxides and silicates) flotation is the result of strong
adsorption in the Stern plane of multivalent species that can reverse the sign of the zeta
potential and cause the formation of a triple layer: the first layer will be the charge on the
surface of the mineral itself; the second layer is the oppositely charged Stern layer; and the
third layer is collector counterions charged similarly to the mineral surface. Thus, activation
of a negatively charged oxide for flotation with an anionic collector requires the strong
adsorption of inorganic cations as activator. Activation in these systems does not produce a
new surface, as in the case of copper activation of sphalerite or the sulfidization of oxidized
lead or copper minerals. Industrially, calcium or magnesium salts are used for the activation
of a mineral such as quartz for oleic acid flotation. Hydrolyzed multivalent metal ions are
strongly adsorbed, and this affinity for a mineral surface has long been recognized. The
papers of James and Healy (1972) provide a method for quantification of these adsorption
processes.
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
47
In 1928, Gaudin et al. were the first to find that pH most markedly affects the flotation
of quartz activated by ferric and cupric ions when oleate is used as a collector. For quartz
activated with ferric chloride, sodium oleate as the collector, and terpineol as the frother,
they found that marked flotation occurs in the pH range of about 5 to 7. In 1934, Kraeber
and Boppel (in Sutherland and Wark 1955) found that pure quartz was not floated by a sulfonated castor oil but that it responded well if treated by a number of different heavy-metal
salts over specific pH ranges for each activating metal salt. Hydrolysis of the activating cation undoubtedly is important in the process. In the case of Fe(III) activation, they found a
rather wide pH range for quartz flotation for their system, pH 2 to 9. In 1966, Mackenzie
conducted a detailed investigation of the effect of ferric chloride on the zeta potential of
quartz, and for an addition of 5.7 × 10–5 M ferric chloride, he observed a marked increase in
the zeta potential at about pH 3, reaching a maximum at about pH 5, and then reversing
sign at pH 7.3. Hergt et al. (in Sutherland and Wark 1955) used contact angles to delineate
the critical bubble contact region for Fe-activated quartz with sodium hexadecylsulfate as
collector, and they observed the lower pH required for contact to be about pH 3.3 and the
pH at which contact ceased to be about 7.5. Schuhmann and Prakash (1950) presented the
results of a comprehensive investigation of activation in the soap flotation of quartz, with
vacuum flotation being their main research tool. For ferric chloride as activator and oleic
acid as collector, they found that the flotation range was between pH 3 and pH 12. Perhaps
with the vacuum flotation test procedure used by Schuhmann and Prakash, the ferric
hydroxide precipitate itself was responding to flotation and carrying with it the quartz particles. Subsequently, M.C. Fuerstenau and associates (e.g., in M.C. Fuerstenau and Palmer
1976) conducted detailed systematic investigations of hydrolyzing phenomena involved in
activation phenomena. Figure 24 presents the results of M.C. Fuerstenau and Palmer (1976)
for the flotation of quartz, with a sulfonate of mol wt 450 as collector, as a function of pH
for various activating metal ions. For clarity, only the initial flotation edge is shown in this
pH of Hydroxo Complex Formation
FeOH2+
AlOH2+
PbOH+
MnOH+
MgOH+
CaOH+
100
Flotation Recovery, %
80
60
Fe3+
Al3+
Pb2+
4
6
Mn2+
Mg2+
Ca2+
40
20
0
0
2
8
10
12
14
pH
Adapted from M.C. Fuerstenau and Palmer 1976.
FIGURE 24 Minimum flotation edges for the flotation of quartz as a function of pH with 0.1 mM
sulfonate collector and 0.1 mM metal ion activators
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48
HISTORICAL ASPECTS OF FLOTATION
figure. As can be seen, there is a distinct correlation of flotation response with the pH of the
hydroxy complex formation. However, in their work, M.C. Fuerstenau and Palmer found
both the upper and lower flotation edge to be steep, in contrast to the work previously discussed. With their system, the pH range for 90% flotation recovery with 10–4 M sulfonate
collector and 10–4 M metal salt as activator was as follows: Fe(III), pH 2.9–3.8; Al, pH 3.8–
8.4; Pb, pH 6.5–12.0; Mn(II), pH 8.5–9.4; Mg, pH 10.9–11.7; Ca, pH 12.0 and greater.
Their experimental techniques clearly delineate the activation pH for optimum floatability.
In silicate mineral flotation, activation by anions has been important. Specifically, fluoride has been widely used as an activator in the cationic flotation of feldspar from quartz,
and as a depressant in the anionic flotation of beryl and spodumene (Smith 1963). By measuring contact angle on quartz and microcline as a function of pH in the presence and
absence of sodium fluoride with DAC as collector, Smith showed that there is a specific pH
range in which microcline (a feldspar) is activated and quartz is depressed, as can be seen
from the results given in Figure 25. In the absence of fluoride, quartz and feldspar behave
identically with the cationic collector. Although the contact angles on quartz are not
affected by the addition of fluoride, those on microcline change significantly. The activation
Quartz
60
Microline
10–2 M NaF
No NaF
4 × 10–5 M DAC
50
Contact Angle, degrees
40
30
20
10
0
0
2
4
6
8
10
pH
Adapted from Smith 1963.
FIGURE 25 Contact angles on quartz and microcline (feldspar) in aqueous dodecylamine as a
function of pH in the presence and absence of NaF, showing the activation of feldspar at low pH
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
49
of microcline can only take place through strong specific adsorption of fluoride ions on the
aluminum sites at the feldspar surface, such that the surface carries a high enough negative
charge to attract the physisorbed cationic collector ions. The high insolubility of aluminum
fluoride suggests that fluoride ions may strongly chemisorb on the aluminum sites at the surface. Shergold, Prosser, and Mellgren (1968) found that inorganic anions function as activators for the flotation of hematite at pH 1.5 with 0.2 mM dodecylamine. At concentrations
greater than 3.5 mM, bubble-pickup tests showed firm bubble–particle attachment with
sodium fluoride (NaF) and NaCl but no attachment with sodium nitrate (NaNO3), sodium
thiocyanate (NaCNS), or sodium acetate (NaCH3COO), and weak attachment with
sodium sulfate (Na2SO4). Batch 1-kg flotation tests with a hematite ore and with synthetic
hematite–quartz mixtures showed excellent flotation separation with hydrochloric acid
(HCl) and also with sulfuric acid (H2SO4) at pH 1.5, with or without a small addition of
ferric chloride (FeCl3). Hematite recovery was about 90% at almost 100% hematite grade.
Evidently, a highly negative surface must be produced by adsorption of the activating anions
or surface complexes for the cationic collector to adsorb.
In the case of oxide and silicate minerals, because collector ions function as counterions
in the double layer, their adsorption density will depend on competition with any other
counterions in solution. Thus, the presence of excessive amounts of dissolved salts can
inhibit flotation because inorganic ions similarly charged to the collector can then act as a
depressant. In the case of the flotation of goethite with quaternary amine salts at pH 11,
adding 0.03 M NaCl will reduce flotation to about nil (Iwasaki, Cooke, and Colombo
1960). Onoda and Fuerstenau (1965) carried out a detailed study of the depression of
quartz flotation with DAA as collector and showed that Ba2+ and Na+ both inhibit flotation, the effect being considerably greater with the divalent salt, as can be seen from the plots
given in Figure 26. Ion exchange as related to flotation would be controlled by the phenomena
involved in the Stern–Grahame equation. In the absence of specific adsorption, the
Quartz
100
60
80
60
40
40
Zeta Potential, –mV
Flotation Recovery, %
80
0.1 mM DAA
pH 6.5
20
Flotation ζ-Potential
20
NaCl
BaCl2
0
10–7
10–6
10–5
10–4
10–3
10–2
10–1
0
100
Added Salt Concentration, M
Adapted from Onoda and Fuerstenau 1965.
FIGURE 26 Effect of adding barium chloride and sodium chloride on the flotation and zeta
potential of quartz in 0.1-mM DAA solutions at pH 6.5, showing the depression of flotation by
ionic competition
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HISTORICAL ASPECTS OF FLOTATION
exchange is controlled directly by the ratio of their concentrations in the bulk solution. In
the case of one ion being multivalent, the valence effect given in that relation is relevant.
Most importantly, if one or both of the ions exhibit specific adsorption, then the magnitude
of the specific adsorption free energy must be taken into account. The concentration of
DAA is 0.1 mM, which is just below that at which hemimicelles begin to form. However,
barium ions can reverse the zeta potential at 0.03 M, indicating specific adsorption and
meaning that both specific adsorption and valency enter into these depression phenomena.
Onoda and Fuerstenau presented a tabulation of the experimental and calculated concentrations of sodium and barium ions that would lead to depression. As an example, to reduce the
recovery to 70%, the concentration of barium chloride needed was 0.15 mM, whereas the
concentration was 6 mM for sodium chloride. The calculated concentration ratio of Na+/Ba2+
was 62 and the experimental ratio 40, which is reasonably close for such simplified calculations.
N AT U R A L F L O TA B I L I T Y
Most minerals are hydrophilic and require collectors for flotation. Because sulfide minerals
readily oxidize when exposed to air, they are hydrophilic under the usual conditions encountered in processing. In the case of oxide and silicate minerals, all but two are hydrophilic. As
for the sparingly-soluble salt minerals, all are hydrophilic because of broken ionic bonds that
form their surface. However, the silver halides exhibit some natural hydrophobicity.
The Bessel brothers in 1877 were the first to utilize natural floatability in their process
for upgrading graphite ores. It was A.M. Gaudin who, in his texts of 1932 and 1957, postulated that the natural floatability or nonpolar character of certain minerals was the result of
not breaking primary bonds upon forming their surfaces. This condition would be met with
crystals held together by dispersion forces (van der Waals bonds). Examples would be
molecular crystals such as sulfur, which consists of S8 rings held in a crystal by dispersion or
van der Waals forces, as well as paraffin. Most of the nonpolar minerals are sheet crystals in
which their crystal chemistry results in individual layers that are electrically neutral, with
dispersion forces acting between the sheets to hold them together. The faces of such crystals
are nonpolar, but the edges would be polar given that primary ionic or covalent bonds are
broken in forming edge surfaces. Examples of such minerals are graphite and two of the layer
silicates, talc and pyrophyllite, which on their cleavage plane present uncharged siloxane
rings. Two sulfide minerals exhibit natural hydrophobicity, namely stibnite and molybdenite, also a layer mineral. As pointed out by Gaudin, boric acid (H3BO3) has a layer structure
in which all potential hydrogen-bonding OH are internally satisfied and not available for
hydrogen bonding with water molecules. It is the strong tendency of water molecules to
hydrogen-bond with each other that provides the energy for water to be displaced from
nonpolar surfaces by an air bubble or oil droplet.
Again, Taggart was at odds with Gaudin over the concept of natural floatability. In
1934, Taggart, del Guidice, and Ziehl wrote, “It may seem odd, at this date, to resurrect so
old a friend as the inherent floatability of minerals, and would be so had not a recent writer
unearthed the ancient fossil for us and dressed it up in modern appearing clothes.… Consequently, we dissent vigorously and finally from any idea of inherent natural floatability.” The
fact that natural hydrophobicity occurs when the cohesive energy of water is greater than
the dispersion forces interacting between water and a solid was not known at that time. Taggart went to great lengths to prove that natural floatability did not exist. It is interesting that
Fowkes and Harkins in 1940 (Harkins being an extremely meticulous surface chemist)
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
51
published their measurements of contact angles using a carefully designed tilting plate apparatus rather than the captive-bubble technique. They reported contact angles for water on
the following hydrophobic solids: Ceylon graphite, 85.7°; talc, 87.8°; stibnite, 84.2°; and
paraffin, 108° to 111°. They obtained the same value of the contact angle whether the liquid
was advancing or receding before the measurement was made, and commented that, “The
‘advancing’ and ‘receding’ angles obtained by nearly all investigators other than those in this
laboratory are due to improper preparation of the surface and poor techniques in making
measurements.” The contact angles with paraffin oil were found to be zero and these solids
considered to be oleophilic by Fowkes and Harkins. The addition of such compounds as
butyl alcohol, butyl amine, butyric acid, propionic acid, and acetic acid lowered the contact
angles on paraffin and graphite.
Laskowski (1986) has reviewed fundamental aspects of the relation between natural
hydrophobicity and floatability. Using an analysis involving concepts of the work of adhesion, Laskowski and Kitchener (1969) concluded that all solids would be hydrophobic if
they did not carry polar or ionic groups on their surface. It is the high cohesive energy of
water due to its hydrogen bonding that gives rise to hydrophobicity (and to the formation of
micelles in solution and hemimicelles at a mineral surface). The role of the flotation collector is to cover the polar sites on mineral surfaces that are formed by the breakage of primary
bonds to prevent hydrogen bonding of water to surface sites.
In the flotation of talc, graphite, or molybdenite, the addition of a neutral oil is used to
enhance the hydrophobicity of the mineral. In many instances, depression of these minerals
is desired, and the standard depressants are hydrophilic polymers that adsorb and inhibit
bubble attachment. The flotation of coal has become important over the last few decades,
but it is a naturally floatable material whose surface is very susceptible to oxidation that can
severely reduce its hydrophobicity.
Figure 27 is presented to illustrate the flotation response of a naturally floatable mineral, talc, without the addition of a collector as a function of pH. Although not shown, the
isoelectric point (IEP) of this talc sample occurs at pH 2. At pH 1, the zeta potential is
+20 mV, and in the pH range of 4 to 8, the zeta potential is about –30 mV and then becomes
more negative, to about –50 mV at pH 10 and above. The greater magnitude of the negative
zeta potential is responsible for the decrease in flotation observed above pH 10. The induction time correlates well with flotation response. Naturally floating minerals, such as talc,
graphite, and molybdenite, are depressed industrially by the addition of hydrophilic polymers. Such polymers are adsorbed at the surface of a hydrophobic mineral by hydrophobic
bonding phenomena (by adsorbing, they effectively increase the hydrogen bonding of water
molecules near the interface). The adsorbed hydrophilic polymer prevents bubble attachment because water molecules now can hydrogen-bond to the polymer. Figure 27 shows the
effect of 8.1 mg/L of dextrin on the flotation response of talc. Independent of pH, this small
amount of added dextrin reduces the flotation recovery to 40%.
Talc can also be depressed by hydrolyzing trivalent cations, as shown by M.C. Fuerstenau, Lopez-Valdivieso, and Fuerstenau (1988) in a detailed electrokinetic and flotation
study with Fe(III), Al(III), and Cr(III). As the pH is increased, the cations hydrolyze and
sharply change to reverse the sign of the zeta potential but do not affect flotation, apparently
because the hydroxo complex species adsorb onto the polar edges of the talc particles.
However, upon further increase of the pH, the metal hydroxide precipitates, the zeta potential becomes positive, and flotation ceases. As the pH is increased further, the zeta potential
of the precipitated hydroxide becomes negative again, and talc once more responds to flotation.
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52
HISTORICAL ASPECTS OF FLOTATION
100
0
Talc
0.002 M KNO3
40
80
60
120
40
160
Induction Time, μsec
Flotation Recovery, %
80
20
Flotation Without Collector
Induction Time Without Collector
Flotation with Dextrin
200
0
0
2
4
6
8
10
12
240
14
pH
Adapted from D.W. Fuerstenau and Huang 2003.
FIGURE 27 Influence of pH on the flotation of talc and induction time in 2 mM potassium
nitrate without the addition of a collector, and the depression of talc by the addition of 8.1 mg/L
dextrin
Depression is due to heterocoagulation of the hydroxide onto the nonpolar face of the talc
particles. The driving force for this must be the displacement of water molecules from the
talc face, thereby allowing water molecules that were at the interface to resume hydrogen
bonding with each other and with the hydroxide coating on the talc. Once the zeta potential
of the hydroxide particles becomes negative again, the hydroxide particles redisperse, once
more providing a nonpolar talc face.
S PA R I N G LY - S O L U B L E S A LT M I N E R A L S
The first systematic research on the flotation of salt-type minerals was conducted by Gaudin
and Martin (1928) on a wide range of carbonates, namely, calcite, magnesite, rhodochrosite,
siderite, malachite, and azurite. They found that aliphatic fatty acids are effective collectors
for these minerals and that there is a pronounced systematic chain-length effect. In general,
the carboxylic acid needed to have at least 7 carbon atoms (heptylic acid), although chains
as short as propionic collected azurite and malachite. The industrial workhorse collector for
salt-type minerals is oleic acid, which interacts with the mineral surface by chemical
exchange. Gaudin and Martin (1928) conducted experiments at higher temperatures with
longer-chained fatty acids and found marked increase in flotation by raising the temperature
from 25°C to 70°C. An increase with rising temperature is a direct indication of activated
chemical reaction taking place—the chemical exchange reaction of carboxylate with carbonate ions in the crystal lattice. More recent measurements of flotation and zeta potentials by
Somasundaran and Agar (1967) showed that DAC and SDS are physically adsorbed by calcite, at least until solubility products are exceeded, as evidenced by plots similar to those
shown in Figure 16.
The flotation of sparingly-soluble salt minerals such as apatite, barite, calcite, and fluorite
appears to be controlled by chemical interaction of the carboxylate collector with mineral
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
53
cations, and collector interaction appears to be controlled by solubility criteria. Many
researchers have conducted IR spectroscopy studies on all of these salt-type minerals, and
the results generally show the expected metal–carboxylate bond from chemical exchange,
chemical reaction, or chemisorption when oleate is used as the collector. Some physical
adsorption has been observed at lower pH values. However, with these types of systems, various researchers have found that collector uptake may go far beyond a monolayer and that a
new soap phase may form at the surface. Thus, the collection process perhaps should be considered as one of surface reaction rather than one of adsorption. In fact, Kitchener (1984)
wrote the following about the flotation of salt-type minerals:
The main problem over soaps is to identify the form of the product, which, in this
case, seems very unlikely to conform to the naïve monolayer model. There is no
doubt that, given a chance, calcium minerals, for example, would go on reacting
with sodium oleate almost indefinitely. This is not reversible physical adsorption;
Atademin has shown that supposed ‘adsorption isotherms’ for such systems are
almost certainly abstraction-by-precipitation curves.
Industrial processes for the recovery of salt-type minerals from oxide and silicate
gangue minerals are quite straightforward. However, separating salt-type minerals from
each other is complex and difficult. For example, several are calcium salts that interact quite
similarly with the collector, or they have slightly different solubilities such that dissolved
anions (or cations) can react with the surface of the less soluble mineral, causing a surface
transformation that leads to reduced selectivity. Flotation separations of these minerals are
effected by utilizing a number of modifying agents that make insoluble inorganic compounds with the alkaline-earth cations in the minerals, including silicate, fluoride, phosphate, and dichromate, or by the addition of organic molecules such as tannins and starches
that coat the surface with a hydrophilic layer of material.
Pugh and Stenius (1985) presented results of a detailed study of the electrokinetic
behavior, solubility, and flotation of fluorapatite, calcite, and fluorite with sodium oleate.
Figure 28 presents their results for the flotation of these three minerals as a function of
sodium oleate concentration at pH 10. This figure shows that the amount of oleate required
as collector follows the order fluorite < apatite < calcite. Fa et al. (2003) also determined the
flotation response of fluorite and calcite as a function of sodium oleate concentration and
obtained fairly similar results, namely, to obtain 50% flotation recovery of fluorite, 3 × 10–6 M
oleate was required and 4 × 10–5 M oleate for calcite. Using molecular modeling, Pradip and
Rai (2002) carried out computations to model the interactions of oleic acid with calcium
minerals and calculated the interaction energies for oleic acid with these calcium mineral
surfaces to be –52.6, –46.8, and –40.2 kcal/mol for fluorite, fluorapatite, and calcite, respectively. They also calculated the interaction energies for water with these minerals, which is
lower in each case, indicating that oleic acid will replace water at the mineral surfaces. Their
calculated interaction energies give the same order as the observed flotation response. Fa and
colleagues suggest that the lower floatability of calcite is due to the low density of calcium
sites at the carbonate surface.
Aqueous solutions of these three minerals are complex, because all of the ions involved
are subject to hydrolysis, depending on pH. Fa et al. (2003) listed the solubility products of
these three minerals as follows: fluorapatite [Ca10(PO4)6F2], 6.3 × 10–137; fluorite (CaF2),
5.0 × 10–11; and calcite (CaCO3), 4.6 × 01–9. Their measured solubilities of calcium ions in
solution after 15 minutes were 2.5 × 10–5, 1.3 × 10–4, and 1.5 × 10–4 M for fluorapatite,
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54
HISTORICAL ASPECTS OF FLOTATION
100
Flotation Recovery, %
80
Interaction Energy Calculation
by Molecular Modeling:
Fluorite
–52.6 kcal/mol
Apatite
–46.8 kcal/mol
Calcite
–40.2 kcal/mol
Fluorite
Apatite
Calcite
pH 10
60
40
20
0
10–6
10–5
10–4
10–3
Sodium Oleate Concentration, mol/L
Adapted from Pugh and Stenius 1985; Pradip and Rai 2002, 2003.
FIGURE 28 Flotation recovery of fluorite, apatite, and calcite as a function of the
concentration of sodium oleate at pH 10
fluorite, and calcite, respectively. Fa and colleagues report high adsorption densities of oleate on these minerals, the amount being related to their rate of dissolution and solubility,
namely 11, 100, and 300 μmol/m2 for apatite, fluorite, and calcite, respectively. Because
monolayer coverage is 6 μmol/m2, any oleate uptake above that amount cannot be chemisorption but must be the calcium oleate soap resulting from surface reaction.
The solubility product of calcium oleate is 3 × 10–16, which indicates that calcium soap
will be precipitated on addition of sodium oleate in alkaline solutions. Free and Miller
(1996) investigated the precipitation and transport of precipitated calcium oleate soap to
the fluorite surface. Because the fluorite surface would have a coating of chemisorbed oleate,
this process is one of coagulation and not really heterocoagulation, with hydrophobic bonding phenomena playing a significant role. In their paper, Fa et al. (2003) showed that colloidal particles of calcium oleate soap coagulate onto the surface of fluorite and make it readily
floatable. A higher concentration of calcium oleate colloids was required to initiate calcite
flotation.
In flotation systems involving slightly soluble salt minerals, a major complication is that
of the conversion of the surface of a mineral to that of another mineral or compound. As an
illustration of surface conversion, consider the use of soda ash on the surface properties of
barite. Equilibrium is controlled by the following reaction:
BaSO 4 + HCO 3–( aq ) = BaCO 3 ( s ) + H (+aq ) + SO 42–( aq )
(EQ 14)
There are several ways to demonstrate that the surface of barite behaves as barium carbonate (BaCO3) rather than barium sulfate (BaSO4) in the presence of sodium carbonate
(Pradip and Fuerstenau 1991). Figure 29 shows that the addition of sodium carbonate for
pH regulation in a flotation separation involving barite and calcite causes barite to behave as
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
55
3
Electrophoretic Mobility, μm/sec per V/cm
2
Reaction Onset
BaSO4 to BaCO3
BaSO4
(Barite)
1
BaCO3
(Powder)
0
–1
–2
–3
24 Hours Equilibration
11
Equilibrium pH
10
BaCO3
9
Ba
SO
Na2CO3
Solution
4
8
+[
HC
O
–
3]
=〈
Ba
CO
3〉
+[
H+
]+
7
BaSO4
6
10–6
10–5
10–4
10–3
[SO
=
4]
10–2
10–1
Initial Sodium Carbonate Concentration, mol/L
Adapted from Pradip and Fuerstenau 1991.
FIGURE 29 Surface transformation of barite to barium carbonate by the addition of sodium
carbonate, as shown by zeta potential measurements and solution equilibrium pH
though it were BaCO3, as indicated by the electrokinetic behavior of the mineral. This surface transformation is controlled by bulk thermodynamics, as would be expected when any
of these processes are a surface chemical reaction. Finally, in this system, the flotation
response of the barite under these conditions must be that of a carbonate.
When working with single minerals of calcite, azurite, and malachite, the results of
Gaudin and Martin (1928) suggest that it should be possible to separate the copper minerals
from calcite with heptylic acid as the collector, but they found that no separation from calcite could be achieved. Sutherland and Wark (1955) stated that this is probably one of the
first examples of cross-activation to be found in nonsulfides. Gaudin and Martin aptly commented: “It is indeed very remarkable that azurite and malachite, two minerals which are
very similar in chemical composition, and crystallographically, can be separated by flotation
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HISTORICAL ASPECTS OF FLOTATION
with one of the lower fatty acids, while the carbonates of copper cannot, by any means now
known, be separated from calcite, by this same reagent.”
Similar to the foregoing statement, in a complex mixture of salt-type minerals such as
calcite and apatite, dissolution of the minerals can put carbonate, phosphate, and fluoride
ions into solution—ions that can all chemisorb onto the different minerals and compete
with the collector. Likewise, calcium ions can precipitate the collector if the solubility of the
calcium soap is exceeded (the KSP of calcium oleate is 10–15.6). Using the known equilibria of
the various species that would be involved with apatite and calcite in water, Anathapadmanabhan and Somasundaran (1984) constructed a diagram showing the amount of Ca2+
in solution from apatite, calcite (closed and open to the atmosphere), apatite, and calcite
supernatants, and also calcium oleate. Any condition where the concentration of Ca2+
exceeds that of calcium oleate can cause collector precipitation and hence depression, particularly when the mineral is in equilibrium with the solution before the collector is added.
The researchers also conducted detailed experiments on the flotation of apatite and calcite
in water, in their supernatants, and with added nitrate, carbonate, and phosphate salts. Figure 30 presents their results for the oleate flotation of calcite in water and in supernatants of
calcite and apatite. In the case of calcite flotation in water, conditions were such that little calcite would have dissolved during the experiments. This figure shows that supernatants of
apatite and even that of calcite depressed the flotation of calcite in the pH region of about 6
to 13. Turbidity measurements after the addition of oleate to the supernatants, but at
slightly lower oleate concentration without added potassium nitrate, are also given in Figure
30 and show the precipitation of calcium oleate from these solutions. Added calcium nitrate
depressed calcite flotation similar to that shown in Figure 30 for the effect of supernatants.
For this system, depression results from the bulk precipitation of the collector as calcium
100
100
80
Turbidity
80
0.1 mM Potassium Oleate
60
Apatite Supernatant
Calcite Supernatant
40
Calcite Flotation
70
0.2 mM Potassium Oleate
20 mM Potassium Nitrate
60
Water
Apatite Supernatant
Calcite Supernatant
20
0
0
2
4
Turbidity, % transmittance
Flotation Recovery, %
90
6
8
10
12
50
14
pH
Adapted from Ananthapadmanabhan and Somasundaran 1984.
FIGURE 30 Effect of apatite and calcite supernatants on the flotation of calcite with potassium
oleate as collector and also the turbidity of the supernatants upon the addition of oleate at
various pH values showing collector depletion by bulk calcium oleate precipitation
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
57
oleate, probably before any appreciable amount adsorbed onto the mineral surface. By
increasing the oleate concentration, flotation was fully restored.
The standard, yet rather inefficient, process for the flotation of bastnaesite ore involves
six stages of conditioning with steam, soda ash, sodium fluosilicate, sodium lignin sulfonate,
and tall oil from carbonate and sulfate gangue, producing a low-grade concentrate. In an
attempt to improve the separation of bastnaesite, (La,Ce)CO3F, from barite and calcite,
conditions were found under which bastnaesite can be floated quite effectively with hydroxamate as the collector (Pradip and Fuerstenau 1983, 1991). This appears to result because
hydroxamate chelates more strongly with rare-earth ions than with Ba2+ and Ca2+. From
their results, the collector interaction with barite is probably one of adsorption, reaching
only a close-packed monolayer. In the case of calcite, uptake initially appears to be that of
adsorption but eventually changes to multilayer uptake. On the other hand, the rare-earth
chelates with hydroxamate are so stable that a surface layer which is equivalent to 5 or 6
monolayers is formed. Again, this can no longer be considered an adsorption process, but
must be one of surface reaction, where ions are pulled from lattice sites and form multilayers
of a metal hydroxamate compound at the surface. If the rate of metal dissolution and diffusion through the boundary layer is faster than diffusion of the collector to the surface, bulk
precipitation may occur. As previously discussed, hydroxylation of the cations in the mineral
surface may assist surface reaction phenomena by first providing some surface atom movement. Readsorption of hydrolyzed species may participate in surface reactions. Flotation
experiments conducted with potassium octyl hydroxamate as collector at pH 9–9.5 showed
that 50% flotation recovery for bastnaesite, calcite, and barite is achieved at the respective
initial collector concentrations of 0.12 mM, 0.30 mM, and 0.80 mM (Pradip and Fuerstenau 1991). Recent computations by Pradip and Rai (2003) show that the interaction
energies for hydroxamate with bastnaesite and barite are –66 and –33 kcal/mol, respectively,
in accordance with the strong uptake of hydroxamate by the rare-earth mineral.
S O M E E X A M P L E S O F P R A C T I C A L M I N E R A L S E PA R AT I O N S
The recovery of copper, lead, and zinc from a complex sulfide ore can be achieved in various
ways. For an ore that might contain galena, sphalerite, and chalcopyrite with such gangue
minerals as pyrite, carbonates, and quartz/silicates, the first step involves the joint flotation
of chalcopyrite and galena at pH 6–7 with xanthate collector and a small amount of sodium
cyanide to depress pyrite and zinc sulfate to depress any sphalerite activated by heavy-metal
ions in solution. Copper sulfate is then added to the tailings from this first step to activate
the sphalerite, and sodium cyanide and lime are added to bring the pH to 10.5 to ensure
depression of the pyrite. With the addition of more xanthate, sphalerite is then floated. If
the pyrite contains gold, for example, it could subsequently be recovered from the tailings.
Separation of galena and chalcopyrite in the bulk concentrate can be achieved by depressing
galena with sulfur dioxide (SO2) or with sodium dichromate at weakly acidic pH values.
Another procedure is to float the galena after depressing the copper sulfides with sodium
cyanide at pH 8–9.
As an example of oxide mineral flotation, with iron ores, the usual problem is separation
of hematite (PZC, pH 7) from quartz (PZC, pH 2). Hematite can be floated away from
quartz with a sulfonate at pH 2–4, or with sodium oleate at pH 6–8. Quartz can be floated
away from hematite with an amine at pH 6–7, or at pH 11–12 with sodium oleate as collector after activating the quartz with calcium ions and depressing the hematite with starch.
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HISTORICAL ASPECTS OF FLOTATION
Although not used industrially in iron ore processing, hematite can be floated away from
quartz with hydroxamate at pH 8.5, or with an amine at pH 1.5 in the presence of hydrochloric or sulfuric acid as a hematite activator.
In the case of a pegmatite containing spodumene (PZC, pH 2), muscovite (IEP, pH 1),
iron-bearing oxide/silicates (PZC, pH 6), feldspar (PZC, pH 2), and quartz (PZC, pH 2),
separations can be made sequentially by utilizing differences in silicate mineral surface
chemistry and crystal chemistry. First, spodumene is floated at pH 7.6 with oleic acid as collector through chemisorption on aluminum surface sites. Then by adding an amine collector at pH 2–2.5, the layer-silicate mineral muscovite is floated by making use of its fixed
negative surface charge. After that, the iron silicate impurities are floated at pH 3 with a sulfonate as collector. Finally, with hydrofluoric acid (HF) as an activator, the feldspar can be
floated with an amine at pH 3, leaving a marketable pure quartz as the tailing.
The simplest large-tonnage separation of a sparingly-soluble salt mineral is exemplified
by the flotation of apatite from quartz. In a typical Florida phosphate plant, both anionic
and cationic flotation are used to produce an acceptable product. After desliming, the phosphate mineral is floated at pH 9–9.5 with fatty acid and fuel oil extender. Subsequently, the
concentrate is acid-blunged to remove the collector coating and then refloated with an
amine at pH 7–8 to remove silica impurities. A major challenge is the increased dolomite
content of phosphate ores, namely, to effectively prevent the dolomite (Ca,Mg)CO3 from
floating with the apatite.
S U M M A RY
Basic flotation research conducted over the last several decades has answered questions
posed by Rickard in 1916 as to why minerals float. By simultaneously using more than one
technique to study the surface chemistry and flotation response of pure minerals with purified chemical reagent systems, the fundamental mechanisms by which sulfide, oxide, silicate,
naturally-floatable, and even sparingly-soluble salt minerals respond to flotation is now
fairly well understood. As outlined in this review, some systems are better understood than
others. Because collector–mineral interactions appear to be more interesting, more research
has been directed toward the behavior of collectors than depressants. How activators and
inorganic depressants function is fairly well understood, but fundamental knowledge of
how and why such organic depressants as quebracho, starch, gum guars, and so forth, attach
to mineral surfaces is lacking. Systems involving mixtures of sparingly-soluble salt minerals
are subject to complex solution chemistry where species from one mineral may dissolve and
adsorb/precipitate onto the surface of another mineral. Furthermore, minerals made up of
ions, such as carbonates, phosphates, sulfates, and sulfides, appear to react with collectors,
consuming reagent and forming precipitates that may adsorb (coat) more than one mineral,
lowering grade. The chemistry of frothers and the role of frothers in determining selectivity
have not received adequate attention.
Real ores do not behave as pure minerals. Mineral grains may have different chemical
compositions (trace elements and locked particles), surfaces smeared with coatings of a
softer mineral in the ore, and highly active surfaces (that may change with time) due to flaws
produced during comminution. More research should be directed toward the study of mineral
mixtures and the behavior of actual ores—but conducted with an aim toward quantifying
what is going on and not just ore testing.
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DEVELOPMENTS IN THE CHEMISTRY OF FLOTATION PROCESSING
59
As speculated by Kitchener (1984), hypothetically, there should be a systematic way to
plan the flotation processing of an ore by establishing the surface chemistry of all the constituent minerals along with their responses to the reagents relevant to the process, then by
determining the likely interferences between the various species, and, finally, by planning
the best conditions for securing a large difference in hydrophobicity between the reagenttreated minerals. Many of the reagents and reagent schemes that have been successfully used
before and after the invention of chemical collectors were found by trial and error, and often
with great ingenuity. However, with decreasing grade, decreasing grain size, and increasing
complexity of ores with the passage of time, there is an ongoing need for more selective and
effective flotation reagents and reagent schemes. In some cases, finding reagents that adsorb
rather than react with a mineral may lead to reduced reagent consumption. The rational
design of new reagents may result from understanding the selectivity of interaction of flotation reagents with interfaces in terms of identifying the underlying molecular recognition
mechanisms.
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Raghavan, S., and D.W. Fuerstenau. 1975. The adsorption of aqueous octylhydroxamate on ferric
oxide. J. Colloid Interface Sci. 50:319.
Ralston, O.C. 1916. Why do minerals float? Page 175 in The Flotation Process. Edited by T.A.
Rickard. San Francisco: Mining and Scientific Press.
Richardson, P.E., and G.W. Walker. 1985. The flotation of chalcocite, bornite, chalcopyrite and pyrite
in an electrochemical-flotation cell. Page 198 in XVth International Mineral Processing Congress,
Tome II. Orleans, France: Bureau de Recherches Geologiques et Minieres.
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Salamy, S.G., and J.C. Nixon. 1953. The application of electrochemical methods to flotation research.
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Shergold, H.L., A.P. Prosser, and O. Mellgren. 1968. New region of floatability in the hematitedodecylamine system. Trans. IMM 77:C166.
Schuhmann, R., and B. Prakash. 1950. Effect of BaCl2 and other activators on soap flotation of
quartz. Trans. AIME 187:591.
Schulman, J.H., and T.D. Smith. 1953. Selective flotation of metals and minerals. Pages 393–413 in
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Simmons, G.L., J.N. Orlich, J.C. Lenz, and J.A. Cole. 1999. Implementation and start-up of N2TEC
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Wakamatsu, T., and D.W. Fuerstenau. 1968. The effect of chain length on the adsorption of sulfonates
at the solid-water interface. Advances in Chemistry Series 79. Columbus, OH: American
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Woods, R. 1984. Electrochemistry of sulfide flotation. Pages 91–115 in Principles of Flotation—the
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The Australasian Institute of Mining and Metallurgy.
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History of Flotation Technology
A.J. Lynch, J.S. Watt, J.A. Finch, and G.E. Harbort
A B S T R AC T
The development of flotation as a major industrial process occurred during three main time
periods. During 1860 to 1900, small-scale attempts were made in industry to float or agglomerate the valuable minerals in ores and wash away the waste minerals. From 1900 to 1925, the
economic necessity to concentrate fine sulfide particles led to immense research efforts for floating
zinc and lead minerals at Broken Hill, Australia (1901 to 1915), and copper minerals at the
huge mines in the western United States (1911 to 1925). During this time, flotation became an
industrial technology and provided much of the copper that made the widespread distribution of
electricity possible. Two major advances occured after 1960. X-ray and radioisotope on-stream
analysis systems were developed, which gave rapid information about assays of process streams
and made accurate process control possible, and new flotation machines were introduced. There
were high-volume columns in which the pulp and air bubbles moved in countercurrent flow,
and high-energy cells in which the pulp was aerated with very small bubbles prior to separation.
The high-energy cells provide much higher flotation rates than the columns. This chapter presents developments in flotation technology that occurred during these periods.
THE NEED FOR A NEW PROCESS
“A new metallurgical process never springs fully developed from the brain of one person, but
is the result of patient investigation, application, and improvement by many minds, during
many years” (Hoover 1914, p. 2). Flotation did not happen in isolation; it was one of many
inventions in the second half of the 19th century that brought a seminal change to mining
and mineral processing technology and greatly increased mineral production. This was an
exciting period in the mineral industry as the Industrial Revolution was gaining momentum
and was causing rapid increases in the consumption of minerals and metals (see Table 1).
In 1850 the mineral industry had been technically stagnant for more than 200 years,
the last major innovations being the use of water power to drive crushing and grinding
machinery in the 16th century and the amalgamation process and blasting by black powder in the 17th century. The industry was ill-equipped then to handle the problems presented by the rising demands for minerals and metals, but it was transformed by new
technology during 1850–1900 and moved from the era of black powder, hand carts, stamp
TABLE 1
World production of copper, lead, zinc, and coal, 1850–1900
Commodity
Copper, kt
Lead, kt
Zinc, kt
Coal, Mt
1850
55
130
65
75
1875
130
320
165
233
Source: Habashi 1994.
65
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1900
525
850
480
660
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TABLE 2
HISTORICAL ASPECTS OF FLOTATION
Production rates and profits at Broken Hill Proprietary Company, 1886–1902
Mining, profits
Kilotons mined
Dividends and bonuses,
$A × 1,000
1886
10
1888
80
1890
170
1892
300
1894
590
1896
440
1898
400
1900
520
1902
660
50
370
1,000
800
580
420
280
180
110
Pb, %
69.2
30.8
100.0
Proportion
Ag, %
Zn, %
49.8
12.8
50.2
87.2
100.0
100.0
Source: Lynch 1987.
TABLE 3
Typical milling statement, Broken Hill, 1900
Product
Lead concentrate
Tailing
Crude ore
Ton
11,141
37,936
49,077
Pb, %
60.6
7.8
19.9
Assay
Ag, oz
19.6
5.8
8.9
Zn, %
10.4
20.8
18.5
Source: Woodward 1965.
TABLE 4 Recoveries of copper by gravity concentration in mills processing porphyry
copper ores
Milling
Year
Tons milled per day
Average copper in ore, %
Average copper in concentrate, %
Copper recovery, %
Utah*
1913
25,000
1.25
17.31
63.95
Chino†
1915
7,357
2.16
21.55
66.59
Ray‡
1915
7,805
1.67
19.29
64.11
Nevada§
1915
8,442
1.54
7.77
70.18
Source: Hines and Vincent 1962.
*Utah Copper Company, Utah.
†Chino Copper Corporation, New Mexico.
‡Ray Consolidated Copper Company, Arizona.
§Nevada Consolidation Copper Company, Nevada.
mills, and sluices to the era of dynamite, steam shovels, ball mills, and Wilfley tables. Even
with all the improvements, there was still a serious problem—fine particles could not be
concentrated efficiently by gravity machines, and fine-grained ores were replacing coarse-grained
ores as the source of many metals. The problem can be illustrated by referring to what happened at the Broken Hill Proprietary Company (BHP) in Australia. Table 2 shows how dividends and bonuses at BHP declined per ton of mined ore from 1896.
BHP started operating at Broken Hill in 1886, and, initially, profits were very high
because miners extracted the surface ore that was rich in coarse-grained silver and lead minerals, but earnings plummeted when this ore was exhausted, and the fine-grained primary
sulfides had to be mined. This was a problem because there were high losses of silver and
lead when the fine particles from the mills were concentrated in gravity machines, and zinc
was lost almost entirely. Table 3 shows a typical milling statement in 1900.
By 1900 the early years of prosperity had given way to pessimism, and employment had
fallen by 30%. The economics were simple—find a new process or abandon the mines.
Heavy investments were made in magnetic separation and in the unproven flotation process,
and it was flotation that provided the answer. The growth of flotation from ideas described
in patents into a remarkable industrial process was described as “…one of the outstanding
achievements in twentieth century technology…” (Klassen and Mokrousov 1963, p. xiv).
The same problem occurred with porphyry copper ores some years later in the western
United States. There were high metal losses using gravity concentration, as shown in Table 4,
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HISTORY OF FLOTATION TECHNOLOGY
67
and a process had to be found to reduce these losses and increase copper recovery by 20%.
Again, flotation was the answer.
The research and development programs that established flotation as the major concentration process at Broken Hill and in the western United States overlapped, the main activity at Broken Hill occurring from about 1900 to 1915 and in the western United States
from about 1911 to 1920. Minerals Separation Company (MSC), which was active in both
areas, ensured that there was transfer of technology, although at the cost of bitter patent lawsuits.
At Broken Hill the major results of the program were as follows:
• Froth flotation, which was developed as an industrial process for concentrating sulfides and was used to extract zinc from millions of tons of slime tailings. By 1907, the
annual production of zinc concentrate had risen to 236,251 tons.
• Differential froth flotation, which was developed during 1910–1915 for making
separate lead and zinc sulfides
Progress must have seemed slow to investors in the Broken Hill mines, but metallurgists
had to find out how to float minerals on water and how to control the many variables that
made the process successful. It was not an easy task. “The propensity of minerals to float was
a valuable discovery, but scores of more elusive and more important discoveries had to be
made before the process could earn a profit. ‘It is to manipulation, learned empirically in the
laboratory and mill, that the flotation process owes its metallurgic success,’ wrote T.A. Rickard” (Blainey 1968, p. 70; Rickard 1932).
In the United States, froth flotation was first used in a zinc mill in Montana in 1911,
and its success gave companies the incentive to investigate the process for the concentration
of copper sulfides. Its potential to improve the economics of copper milling was realized in
1915 when a 15,000-tons-per-day flotation plant was built at the Inspiration Company, and
the recovery of copper was increased to 80%. Not surprisingly, flotation circuits swept the
copper industry within a few years. How flotation developed as a great industrial process
will be discussed in this chapter. It is necessarily brief, and more information about the early
years of flotation is given in the bibliography, in particular, Hoover (1914), members of
the Broken Hill Branch of the Australasian Institute of Mining and Metallurgy (1930),
Hines and Vincent (1962), Crabtree and Vincent (1962), Fuerstenau (1999), and Megraw
(1918).
E A R LY I D E A S , 1 8 6 0 – 1 9 0 0
The first hint that differences in surface properties could be used to separate minerals
appeared in a patent awarded to William Haynes (Haynes 1860). The process claimed that
sulfides in a powdered ore could be agglomerated by oil and the nonsulfides could be
removed by washing. There is no evidence that the idea was tested in a plant.
The first commercial flotation plant was built by the Bessel brothers in Dresden, Germany, in 1877 to clean graphite ore (Graichen et al. 1977). Adolph Bessel graduated from
the University of Gottingen in 1855 and joined a factory that made refractories and crucibles in Grobalmerode. In 1864 this factory was moved to Dresden, close to the Polytecnic
and the Bergakademie Freiberg. Adolph and his brother became its owners in 1866. Because
the quality of the graphite used in the crucibles was poor, they developed a process for cleaning it that involved mixing graphite ore with a small amount of oil, adding water, and boiling
the mixture to float the graphite to the surface of the pulp. Their process yielded a concentrate containing 90% graphite from 40% graphite in the feed. Bessel patents of July 2, 1877,
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HISTORICAL ASPECTS OF FLOTATION
and May 12, 1886, contained all the essential features of froth flotation including the use of
nonpolar oils to enhance process kinetics (Bessel 1877, 1886). The first patent referred to
bubbles being generated by boiling and the second to CO2 bubbles being generated by the
reaction of lime with acid. In 1878 the Wohler Gold Medal was awarded to Adolph Bessel
for the invention patented in 1877.
The first flotation plant to process sulfide ores was based on Carrie Everson’s work,
although it was not a commercial success as was Bessel’s plant. The circumstances were
unusual. Everson was born in 1842 in Massachusetts and studied medicine before marrying
Dr. W. Everson in 1864. He invested in mining shares which failed, and this led to Carrie
taking an interest in mineralogy so she could understand the reason for the failure. About
1878 she started to experiment with ways to concentrate sulfide minerals and eventually
patented a process in 1885 for separating sulfides from gangue by mixing powdered ore with
a small amount of oil in an acid solution and floating the sulfides in a scum (Everson 1885).
Flotation must have been due to entrained air. The myth is that Everson’s patent originated
in observations made while washing geologists’ sample bags; the less romantic reality is that
she was a good scientist who would carry out experiments in a laboratory and was prepared
to test the results in practice. Everson’s process was successful in small plants but not on a
larger scale (Megraw 1918), perhaps because the ores were unsuitable—sulfide flotation did
not reveal its secrets easily. Unfortunately, she did not have the financial resources to continue her research, and she became a teacher to earn a living. Later assessments of Everson’s
work were “…if the invention had been a less startling innovation, it would probably have
received more attention from engineers and metallurgists, and the application of the idea
would probably in that case have taken place many years before it did,” (Hoover 1914 p. 6)
and “as a metallurgist she was a quarter of a century in advance of her profession” (Megraw
1918 p. 7).
The year 1885 was important in the history of flotation because of the patents by the
Bessel brothers and Carrie Everson. The same year, a patent was awarded to Hezekiah
Bradford in the United States for a film flotation process in which powdered sulfide ore was
placed gently onto the surface of water and the sulfides adhered to the surface while other
minerals sank (Bradford 1885). It is likely that these inventors made their discoveries independently, and although their efforts had little immediate technical impact, their patents
showed that the potential significance of flotation-type processes was becoming recognized.
In 1898 Francis Elmore patented a process for concentrating sulfide minerals by adding
oil to pulverized ore in water, agglomerating the sulfides and buoying them to the surface of
the water, and washing away the gangue particles (Elmore 1898). He proved its value at
the Glasdir mine in Wales, and his work was discussed at an Institution of Mining and Metallurgy meeting in London in 1900 (Hoover 1914). The process was widely applied and can
be regarded as the first successful process for floating sulfides, although it was not froth flotation as it is currently known. Entrained air was an unrecognized but important factor. By
1900 it seems that only the Bessel brothers had deliberately used gas bubbles to accelerate
flotation rates, but in 1901 an engineer in Italy, Alcide Froment, patented the use of gas bubbles
to float sulfide particles (Froment 1902). Froment also used sulfuric acid and limestone to generate bubbles although he recognized that gas of any kind would be suitable (Hoover 1914).
This is the background to the events that occurred at Broken Hill during 1901–1915.
Engineers there would have known of the process in which sulfides could be agglomerated
by oil and gangue washed away, so it is not surprising that they became interested in the
technique.
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HISTORY OF FLOTATION TECHNOLOGY
69
F R O M A S TA R T L I N G I N N OVAT I O N T O A N I N D U S T R I A L
PROCESS, 1901–1925
Zinc and Lead at Broken Hill, 1901–1915
In 1901 the immediate problem at Broken Hill was how to extract zinc from the dumps. As
a measure of their size and potential, in 1904 the dumps contained more than 7 million tons
of tailings that assayed at about 4% lead, 140 g/ton silver, and 15% zinc. A process to extract
zinc would revive the mines and restore prosperity. The zinc problem at Broken Hill was,
therefore, the question of the hour. From 1901 engineers at Broken Hill worked on several
flotation processes and machines (Woodward 1965).
• Froth flotation was independently investigated by Guillaume Delprat, general manager of BHP, and by Charles Potter, who was a brewer in Melbourne. They each patented a process in which the zinc mineral in gravity plant tailings was floated by
carbon dioxide generated by adding acid to hot pulps that contained carbonate minerals (Potter 1902; Delprat 1902). Neither used oil; probably the tailings contained a
sufficient amount to make flotation occur. Potter’s process had a short life, but the
BHP process worked on gravity plant tailings from 1902 to 1923, the acid consumption being 26 lb per ton and the pulp temperature being 82°–88°C (180°–190°F).
About 90,000 tons of zinc concentrate was made annually from 300,000 tons of tailings.
• A film flotation process that was similar to Bradford’s 1886 invention was patented
by Auguste de Bavay in 1904. In this process a pulp that had been deslimed, acidified,
and oiled flowed down a corrugated cone dipped at an angle into water. The hydrophobic sulfides floated on the water, whereas other particles were wetted and sank. It
worked at Broken Hill from 1905 to 1917, and at its peak produced 80,000 tons of
zinc concentrate annually from 300,000 tons of tailings.
• Vacuum flotation, a form of froth flotation, was patented by Francis Elmore in 1904
and was used in the Zinc Corporation plant for 6 years. In this process, a small
amount of oil was added to an acidified tailings pulp, and the sulfide particles were
floated with bubbles generated by applying a vacuum of 600 mm of mercury to the
pulp and precipitating the dissolved air. At its peak, it produced 80,000 tons of concentrates annually from 250,000 tons of tailings.
• The Minerals Separation Company, which had been formed in England in 1903 to
specialize in ore dressing problems, came to Broken Hill in 1904 to test a process patented by Arthur Cattermole (1902) for which it had purchased the rights. In this
process, a small amount of oil—insufficient to give a buoyant effect—caused the sulfides to agglomerate and sink, and other minerals were washed away. This process
was a failure because the granules of sulfides that still required further concentration
tended to break on the concentrating tables. Another test proved successful, however, wherein an even smaller amount of oil was added and the pulp was violently agitated to entrain air because the sulfides were carried into a froth and removed in a
spitzkasten. Staff at MSC and the Central Mine at Broken Hill developed this concept into stirred flotation cells that were used in series.
• Sketches of early flotation cells are shown in Figures 1 and 2. All the concentrates
made by flotation at Broken Hill contained 47%–49% Zn; recoveries were more
than 80%, and working costs varied only by 10%. A measure of the success of flotation
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HISTORICAL ASPECTS OF FLOTATION
To Vacuum Pump
Separating
Cone
Acid
9 ft
Worm-wheel
Feed
Oil Acid
Froth
Launder
15 ft
Feed
Mixer
Baffles
Tailing
Tailing
Potter-Delprat Cell
Concentrate
Elmore Vacuum Cell
Feed
Regulating Wheel
for Discharge Valve
Corrugated Cone
Water Level
Overflow
Launder
Froth
Launder
Agitation
Box
Water Level
Feed
Frothing
Chamber
Discharge
Tailing
Minerals Separation Cell
De Bavay Cone
Source: Truscott 1923.
FIGURE 1
Flotation cells at Broken Hill, 1902–1910
Feed
Concentrate
a
f
e
a — Thickener
b — Feed tank
c — Pump
d — Separating cone
e, f, g — Pipe and valves for air and frother
k, l — Concentrate, tailings outlets
Lyster Cell
c
g
h
k
d
Impellor
l
Air Inlet
Owen Cell
Source: Hoover 1914.
FIGURE 2
Cells built for differential flotation at Broken Hill, 1911–1913
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HISTORY OF FLOTATION TECHNOLOGY
71
was that in 1913, 10 years after the start of investigations, more than 11 Mt of material had been floated, and 3 Mt of zinc concentrate produced (Woodward 1965;
Hoover 1914).
While these cells were being developed and used at Broken Hill, a flotation cell was patented in 1903 by H.L. Sulman and H.F.K. Picard of London, in which particles were floated
by air bubbles formed in the cell by compressed air flowing through holes in an immersed,
perforated tube (Sulman and Picard 1903). It would be some years before pneumatic cells
were used in plants. By 1908 bulk flotation of zinc concentrates was working well, and it was
time to develop a differential flotation process for primary sulfides. Three approaches were
investigated (Woodward 1965):
1. In 1910, E.J. Horwood of BHP roasted to 400°–500°C a concentrate made by bulk
flotation to “deaden” galena by oxidizing its surface to lead sulfate. The blende was
unaltered and was refloated to make salable concentrate.
2. In 1912, F.J. Lyster, mill superintendent at the Zinc Corporation, observed that the
natural flotation rates of galena and blende were different and devised a process to
make separate concentrates. Galena was collected during gentle flotation of an alkaline pulp to which eucalyptus oil had been added as a frother; blende was then
floated from the deleaded pulp. Lyster recognized the importance of air control and
devised a cell in which air was added to a pulp, and the mixture was passed through
a pump before entering a tank in which the froth separated. Differential flotation
was achieved by controlling the air flow rate. A subaeration cell was developed by
T.M. Owen at Broken Hill South in 1913 for the same purpose, and in an improved
form, it became the standard cell.
3. Leslie Bradford at BHP activated and depressed minerals selectively by adding
chemicals to the pulp. In 1913 he patented his discoveries: that copper sulfate activated sulfide minerals and that sulfur dioxide depressed blende during galena flotation. Some months later, John Myers in the United States independently discovered
that copper sulfate was an effective activator.
These discoveries, in particular the selective activation and depression of minerals,
changed flotation from an inflexible bulk process into a process that could be used for the
production of individual mineral concentrates. By 1916 the urgent problem of finding a
new concentration process for the Broken Hill ore had been solved; the prosperity of the
town was ensured, and the turbulence and excitement associated with testing new ideas
diminished, at least for a time.
Copper in the United States, 1911–1920
Film flotation was the first flotation process used successfully in the United States.
Machines designed by Arthur Macquisten in Scotland were used in Nevada in 1906 and in
Idaho in 1911. The principle of operation was that deslimed and oiled sands flowed through
a rotating drum designed to continually lift the particles and gently present them to the surface of the pulp (Figure 3). Sulfides adhered to the pulp surface and were collected. Macquisten cells worked well on sands; the Nevada plant produced a 20% copper concentrate
from a 2.5% copper feed, and the Idaho plant, which operated for 10 years, produced a 45%
zinc concentrate and a high-grade lead concentrate (Truscott 1923; Crabtree and Vincent
1962).
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HISTORICAL ASPECTS OF FLOTATION
6 ft
Water Level
Feed
Roller
Concentrate
Tailing
Source: Truscott 1923.
FIGURE 3
Macquisten film flotation cell
In 1911 froth flotation was used in the United States for the first time (Hines and Vincent 1962). “James M. Hyde installed the first froth flotation plant in the U.S., about August
1, 1911, at the Basin Reduction Company plant, Basin, Montana, which was then under
lease to the Butte and Superior Copper Company (Hines and Vincent 1962, p. 11). …It
would have been difficult however to pick an ore as suitable as the Black Rock ore for treatment by froth flotation with the knowledge available in 1911. The ore contained only 1%
pyrite, 1.1% Pb as galena with which the silver was associated, 17–20% Zn, and 0.25% Cu
(p. 19).” The events relating to this plant have been fully described by Hines and Vincent
(1962); suffice it to say that an important technical innovation was the use of rougher and
cleaner cells in closed circuit with cleaner tailing returning to the rougher. James Hyde
clearly understood the importance of cleaning concentrates to remove entrained gangue
particles although the 50-ton mill at Basin was his first experience with froth flotation. During the next year, the company built mills with capacities of 200 and then 1,200 tons per day
to verify that flotation would work, and with their success, flotation was poised to take off.
The Basin ore was a good ore to start with, but the real prize would be the copper ores.
Growth in the use of copper in the United States during the late 1800s and early 1900s was
5.8% annually, and even the rich deposits in Upper Michigan, Montana, and Arizona could
not support this indefinitely. Lower-grade deposits had to be mined eventually. Daniel Jackling
set the pattern in 1903 when he was given the task to build a 300-metric-tons-per-day mill
to process 2% copper ore and persuaded the owners to mine 5,000 metric tons per day,
which required a total change in the mining and milling systems. The new mill was ready for
operation in 1907, and even at 60% recovery, it made large profits on 1% ore. Its capacity
was soon doubled and redoubled. Others followed his lead, and large mills were built in
Nevada and Arizona to process low-grade ores that were also very profitable at high copper
prices although the recoveries were low.
The Inspiration Company was a leader in developing flotation for porphyry copper
ores. In 1911 it owned part of a huge deposit in Arizona, but this ore gave poor results with
simple gravity concentration. The investment required to make it profitable could not be
made by its backer, W.B. Thompson, and it was purchased by the Anaconda Company. Dr.
Louis Ricketts became consulting engineer. “He did not think the mill Thompson’s engineers had planned would recover enough of the copper. To the horror of the stockholders,
he threw away one million dollars’ worth of mill construction and spent a year and another
million dollars experimenting. Then, he built the first mill that used the new flotation process. The result of Dr. Rickett’s delay was that this company caught the high copper price of
1915 with the most successful mill that had ever been built” ( Joralemon 1973, p. 243). In
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HISTORY OF FLOTATION TECHNOLOGY
Feed
9 ft
Water
Level
73
Overflow Lip
Float
Porous
Canvas Mat
Concentrate
Tailing
Feed
Air Box
Air at 4.5 lb
Cross Section
Inclination 1:20
Air Mat
Tailing
Complete Cell
Source: Truscott 1923.
FIGURE 4
Pneumatic Callow cell modified by Inspiration Copper
late 1912, Dr. Ricketts carried out flotation tests in the MSC laboratory and obtained 87%
recovery of copper in a 15% concentrate from a 2% ore. Then he built a 50-tons-per-day mill
in early 1913 and a 600-tons-per-day mill in early 1914 to verify the results. His testwork
was comprehensive and included studies of different flowsheets and flotation machines,
new reagents, and the effect of fine grinding. In 1915 a 15,000-tons-per-day mill was built,
and 80% of the copper in the feed was recovered. It was the first mill in which flotation was
applied to ore instead of gravity tailings.
Two of the innovations in the new mill were ball mill–classifier circuits for grinding the
ore directly to flotation feed size and flotation machines that used compressed air. Both of
these had an effect on the control of the process: the closed grinding circuits controlled
particle size and minimized the production of coarse, composite particles; and the compressed air cells controlled bubble size. The cells built by J.M. Callow (see Figure 4) were
chosen by the Inspiration Company in preference to subaeration cells, and their success led
to their wide use in concentrators for many years. “Flotation spread at a rapid rate; by 1914,
42 mining companies were operating or experimenting with the flotation process. The list
increased in 1915 to include most of the principal copper and lead mines” (Hines and Vincent
1962, p. 29), and in 1918, 25 million short tons of copper ore were concentrated by flotation.
A spin-off from the new technology was a growing business in flotation chemicals for
enhancing or retarding the flotation rates of specific minerals, and by 1916 many companies
were making reagents. By 1925 thousands of patents had been awarded for flotation chemicals, the most important being to Cornelius Keller and Carle Lewis in 1923 of MSC for the
use of xanthates as collectors for sulfides. Xanthates took much of the guesswork out of sulfide flotation because they increased flotation rates of sulfides considerably; they were soluble in water, and their addition rates could be controlled. It is not known how Keller and
Lewis came to discover the collecting properties for sulfides; perhaps it was because sodium
ethyl xanthate was used in making rubber and as a defoliant and herbicide, or the initial tests
might have been conducted because a bottle of xanthate was readily available on the shelf.
The result of their discovery was that xanthate-lime-pine oil circuits were soon in common
use, and by 1925 xanthate had transformed flotation into a process that was stable and reliable because it could be added in controllable amounts.
With the success of copper flotation, the “startling innovation” proposed by Carrie
Everson had become a reality. Because she lived until 1914, she saw the start of large-scale
flotation and would have known that her remarkable, but unrewarded, efforts during 1885–
1892 had not been in vain.
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HISTORICAL ASPECTS OF FLOTATION
Zinc and Lead at Cominco in Canada, 1917–1922*
Flotation started in Canada in 1917 at the Sullivan mine of Consolidated Mining and
Smelting Company (Cominco). The metallurgist responsible was Ralph Diamond who had
joined the Anaconda Company in Anaconda upon graduation from the University of Toronto in 1913. In May 1914 he was invited to lead a project to study a new “secret” process.
The first flotation plant in North America had been built at the Superior Mill in Butte,
Montana, and arrangements were made for Diamond to learn laboratory flotation testing
there. The Superior Mill used the Hyde process on ore from the Black Rock mine. Following this assignment, Diamond commenced testwork on Anaconda slimes under the direction of George Chapman of MSC. Chapman had been one of the pioneers in froth flotation
in Australia between 1904 and 1906. Later, in 1914, Diamond worked under J.M. Callow at
the Inspiration Test Mill in Miami, Arizona, on MS cells in which ore was being treated at a
rate of about 16,000 tons per day. It is interesting that the Anaconda plant used the experimental Callow pneumatic cell. The Anaconda slimes flotation plant started operation using
standard Hardinge mills—likely an association that resulted in the use of Hardinge mills at
Sullivan in 1922. The Anaconda Company had taken out a license under MSC.
The processes used for copper slimes and copper sands and zinc ore had been developed
by MSC, largely under the direction of George Chapman. Diamond remained in charge of
flotation research for Anaconda until February 1917. At that time, Diamond went to Utah
to install and operate a flotation plant for the Ohio Copper Company near Bingham, Utah,
for the treatment of a partially oxidized copper ore. While in Utah, Diamond was contacted
by Selwyn Blaylock, assistant general manager of the Cominco, regarding work in the development of the electrolytic treatment of zinc ores. This proposal was declined but resulted in
a proposition to instigate testwork on the application of the froth flotation process to the
refractory ore from the Sullivan deposit. The Sullivan deposit presented two new challenges
to existing froth flotation practice. These were, firstly, a remarkably fine association between
the valuable galena and blende minerals and the gangue iron sulfide present. Secondly, the
ratio of iron sulfide, mainly pyrrhotite together with a small amount of pyrite, to the lead
and zinc sulfide minerals was significantly higher than in ores previously studied. A sample
of ore was sent to Diamond in Utah for preliminary testing. This work was sufficiently
encouraging that Diamond joined the Consolidated company in June of 1917 and continued flotation work on the Sullivan ore at Trail. By the end of 1918, a successful three-stage
differential flotation process had been demonstrated on a 600-tons-per-day test mill. The
1918 annual report of the company includes the following statement: “Important improvements in metallurgical practice in regard to treatment of the complex ores of the Sullivan
Mine have added many years to the life of that property and have made it one of the most
valuable mineral deposits in America if not in the whole world.”
Cominco had taken out an MSC license in 1917. This was imperative despite the high
royalty rates as MSC controlled most of the important patents. Furthermore, much of the
general flotation knowledge that prevailed among millmen in those early days stemmed
from the MSC reservoir of knowledge although MSC’s work on Sullivan ore was negative
throughout. MSC never had an employee stationed in Trail to work on the Sullivan deposit
problems, not even for a few days. MSC had a laboratory in San Francisco, but trained men
were at a premium, and at that time, and for the next few years, they were simply deluged
with ore samples for testing, requests for information, and help in the field. They had just
*This section was written by Mike Fairweather.
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HISTORY OF FLOTATION TECHNOLOGY
75
three or four and occasionally five men at any time who were available for all laboratory and
fieldwork. Most of the ores submitted were simple and reacted favorably to the existing acid
or neutral circuit flotation configurations. Some of these represented great reserves, potential large operations, and large royalty producers. With many, major construction was soon
under way. All ores as received were simply subjected to certain standard tests, and those
presenting difficulties were set aside. There were plenty of simple ores to fully occupy their
time. The Sullivan was a very complex ore and did not react to standard tests. Moreover,
MSC knew little at that time of the potential of the Sullivan ore body, and Cominco, as a
company, did not occupy a prominent place in the mining world. In those early years, MSC
was not impressed with the possibilities of copper sulfate, and this reagent was not included
in testwork in 1916. It was around 1920 that they became aware of its importance. Sullivan
ore was a special problem, and copper sulfate meant more to its successful treatment than
was the case with 99 out of 100 ores.
Coal in Europe
“It was not until 1920 that the experimental work of Bury, Broadbridge and Hutchinson
directed attention to the possibilities of the use of froth flotation for coal cleaning. The first
froth flotation plants for coal cleaning were erected the same year in Spain and in France.
The first British plant was erected in 1922, and during the following year, plants were
erected in Germany and in Belgium” (Chapman and Mott 1928, p. 385). Flotation was the
only process available in 1920 to clean the fine fraction (–0.5 mm) of coking coal. Laboratory studies of coal flotation had started in the United States in 1915, and it was found to be
very effective, but there was little incentive to use the process because coal was mined from
thick seams, and relatively few fines were made during mining and transport. The high cost
of dewatering coal concentrate was also a deterrent.
In Europe, coal mining was mainly carried out in thin seams underground, and mechanized mining was used, which generated a high proportion of fines, so there was a higher
incentive to use flotation. Consequently, coal flotation was first used there in 1920, and by
1925, flotation was cleaning about 1 Mt of coal annually. In 1927, 36 plants were using
MSC cells, 14 being in Spain and 12 in Germany. Coal flotation took much longer to
become popular than sulfide flotation because it could be used to clean only a small part of
the mined coal, and its product was low in value. It did grow, however, and in 1933, there
were about 60 coal flotation plants in Europe and 1 in the United States (Aplan 1999). It is
interesting that the Elmore vacuum process, which had only limited success on sulfide ores,
was widely used in England for many years to float minus-1/8-in. coal because the power
required was low, the froth was easy to dewater, and coarse particles were floated.
Technology Transfer and Litigation
The growing interest in flotation was reflected by the fact that the Engineering and Mining
Journal published 3 articles about the process in 1901, 10 articles in 1902, and 32 in 1903
(Hoover 1914). There were many inventions and patents, and inevitably, there was litigation
between inventors who developed virtually identical processes at different mines and who
wished to claim damages for breach of patents. An important issue was license fees. These
encouraged the sale and transfer of technology if they were reasonable but often led to
piracy and secrecy if they were excessive. At Broken Hill the royalty cost in flotation was up
to 30% of the operating cost, so it is not surprising that companies sought to make changes
in the process to avoid paying royalties. Litigation often continued for years and must have
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HISTORICAL ASPECTS OF FLOTATION
rewarded many lawyers with high incomes. Whether litigation helped or hindered flotation
is debatable, but there is no doubt that the expensive court cases are a reminder that legal
action should be a last resort in metallurgical disputes.
To close this section on the formative years of flotation, a comment should be made on
the contribution of MSC. It was formed in London to apply oil agglomeration processes to
the Broken Hill ore to concentrate sulfides, but by 1904, it had found that floating sulfides
in a scum with a little oil gave better results than agglomerating them with excess oil. So the
company turned to flotation. Then came stirred cells, froth flotation, and eventually xanthates—the company was involved in all three processes. The company employed skilled
and experienced engineers who successfully promoted the flotation process and sold MSC
flotation cells in many countries. Not surprisingly, MSC became involved in many lawsuits
as staff left the company and worked on flotation as consultants or employees for other companies. There is no doubt that staff of MSC contributed much to the growth of flotation
and that the rather aggressive tactics of the company helped the transfer of the new flotation
technology considerably.
Y E A R S O F C O N S O L I D AT I O N , 1 9 2 5 – 1 9 6 0
By 1925 efficient subaeration cells were available, and chemicals were being used for the
selective activation and depression of minerals. This was just as well, given that technical
progress was slow during the next 25 years, which were dominated by economic depression
and war. But the demand for minerals produced by flotation continued to increase. The
scope of flotation also expanded during this time, two examples being an oil flotation process that was patented in 1928 to recover phosphates from previously discarded fines and a
process that was developed in 1937 to recover potash from salt-saturated brines.
To meet the increased demand for flotation, plant sizes increased. But caution was the
order of the day in plant design, and large plant capacities were obtained by use of many
small units that were known to be reliable rather than by use of new large units that promised
economy of scale but were untested. One example was the Morenci concentrator built in
1942, which had 432 cells, each with 2.2 m3 of volume to float 45,000 tons per day of ore.
A by-product of the early years of flotation was the interest of many senior engineers in
research because they had seen the immense rewards that it could bring. The result was that
companies established research units in several universities in the 1920s to work on flotation
fundamentals, and these contributed much to the understanding of flotation. Particular
mention is made of the groups led by A.F. Taggart at Columbia University (New York),
A.M. Gaudin at the Massachusetts Institute of Technology (MIT), and I.W. Wark at the
University of Melbourne (Victoria).
A D VA N C E S I N F L O TAT I O N T E C H N O L O G Y, 1 9 6 0 – 2 0 0 0
The production of minerals by flotation increased rapidly from 1950, as shown in Table 5
for copper and zinc, and during this time there were many improvements in the process. The
focus was on reagents, flotation machines, and the control of circuits. Discussion in this section will be limited to flotation machines and to on-line analysis of pulp streams, which is
necessary for accurate control.
The most commonly used flotation machines can be divided into three groups based on
flotation rates:
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HISTORY OF FLOTATION TECHNOLOGY
TABLE 5
77
World mine production of copper and zinc, 1950–2000
Metal
Copper, Mt
Zinc, Mt
1950
2.3
2.0
1960
4.1
2.8
1970
6.3
5.7
1980
7.6
6.1
1990
9.0
7.0
2000
11.0
8.9
Source: International Zinc Association 2003; International Copper Study Group 2004.
1. Flotation columns in which feed pulp enters the collection zone at the top of a column, flows downward, and contacts air bubbles that have been generated by spargers at the base of the column. The column tank itself is both the primary collection
zone and the disengagement zone. Columns are considered to be low-intensity
machines, and the flotation rate constants are low.
2. Mechanically agitated flotation machines in which a rotating mechanism is used to
keep solids in suspension and to create bubbles by shearing air that is either applied
to the machine under pressure or induced. They are considered to be mediumintensity devices with flotation rate constants 1.2 to 1.5 times that of a column.
3. High-intensity flotation machines that consist of an external aeration/contacting
mechanism by which pulp is brought into intense contact with fine bubbles. The
external contactor can use either pressurized air or air entrained into a fluid jet. The
contactor is the primary collection zone, and the tank is the disengagement zone.
These machines have flotation rate constants 2 to 4 times greater than that of the
mechanical flotation cells.
Only mechanically agitated machines were used until the 1960s, after which columns
were introduced, followed by high-intensity machines. These type of machines will be discussed in the following sections.
Mechanical Cells
Mechanical flotation cells have changed little in their principle of operation since they were
invented in 1912, although many aspects of their operation have been improved. Significant
literature is available on their design and development, with excellent recent reviews by
Arbiter (1999), Weber et al. (1999), and Yianatos (2003). Their maximum size has been
increased greatly during the last 35 years to reduce the capital cost of equipment and the
floor area required per metric ton of ore floated in high-capacity plants. Increase in maximum cell volume from 1960 to 2004 is shown in Table 6.
Pneumatic Flotation
These machines introduce air through diffusers or aerators (or, as a general term, spargers)
rather than by mechanical dispersion. Flotation columns are the current principal representatives of this class. Their history contains the colorful characters, claims, and counterclaims
that have marked flotation from the earliest days.
The first machine with obvious column geometry, tall relative to its side dimensions,
appears to have been invented by Norris (1907; Figure 5a). The novelty, however, seemed
less to do with geometry and more with introducing air using pressurized water to overcome
the limitations of vacuum release flotation at high altitude. Towne and Flinn (1919) patented a column depicting a countercurrent-flow slurry descending against a rising swarm of
bubbles (Figure 5b). Towne and Flinn made some prescient observations: that the addition
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HISTORICAL ASPECTS OF FLOTATION
Maximum size of Outokumpu mechanical flotation cells (m3), 1950–2000
TABLE 6
1970
16
1976
38
1980
60
1995
100
1997
160
2002
200
2004
300
Source: S. Ronkainen, personal communication.
Wash Water
1
2
1 – Feed
2 – Concentrate
3 – Tailing
4 – Air
2
2
1
1
4
3
3
4
A. 1907 Norris column
(Norris 1907). Air
injected into pulp.
Not tested in plants.
FIGURE 5
4
3
C. 1963 Boutin and Tremblay
B. 1919 Towne and Flinn
column (Taggart 1927;
column (Boutin and Tremblay
1963; Rubenstein 1995). An
Rubenstein 1995). Air
air sparger and wash water
entered through porous
medium. Unsuccessful in
were used—a simple and
successful column.
plants because of sanding.
Evolution of flotation columns
of oil increased “bubble-forming capacity” and that attached “oil-jacketed” particles stabilized froth. They also noted that particles not attached to bubbles in the froth “settle back
towards the water or pulp column”; that is, they were describing entrainment and dropback.
Depending on the perspective, either of these candidates represents the “first” flotation column (Figure 5c; Rubenstein 1995; Jameson 2002).
Initially, interest in pneumatic flotation was short-lived and not revived until the second half of the 20th century. The problems that developed, which were articulated in the
patent issued to Hollingsworth (1968), included sanding of coarse particles in the absence
of mechanical agitation, plugging of porous diffusers, and channeling in scaled-up versions.
It was not so much solving these problems (many still exist, in fact) that prompted the rapid
commercial expansion after about 1980 but rather the use and control of wash water into
the froth—the “key feature which permits high upgrading” (Finch and Dobby 1990, p. 3).
If wash water is the “key,” it is of interest as to who first claimed this innovation. Searching column patents has identified four that incorporated wash water: Bennett and Dell
(1963), Boutin and Tremblay (1964), Hukki (1967), and Hollingsworth (1968). For
wash-water use, the patents reference Bennett and Dell back to 1958, Boutin and Tremblay
to 1963, Hukki to 1965, and Hollingsworth to 1965. Therefore, the honor of first mention
appears to go to Bennett and Dell.
The use of water sprays was already known to increase grade by washing away of gangue
particles, which are mechanically entrained into the froth (Klassen and Mokrousov 1963).
But application was fitful. Why was it more successfully exploited in columns? It was certainly not because wash water “can kill the froth” in mechanical cells (Fuerstenau and Han
2003, p. 297). Column geometry, small cross-section to volume compared to most mechanical cells, does economize on wash water. But, most importantly, column operators learned
how much to add. Column Flotation of Canada Ltd. (the first column flotation company,
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HISTORY OF FLOTATION TECHNOLOGY
79
founded to develop and market the invention of Boutin and Tremblay) promoted the addition of sufficient wash water to generate a so-called “positive bias,” that is, a net flow of water
downward across the froth–pulp interface, which could be controlled by maintaining a tails
flow higher than the feed flow (ideally, this refers to flow of water, but in reality, it is usually
flow of slurry). Operating with a positive bias helped ensure that entrainment was countered.
Wheeler (1983), president of the company after 1963, openly discussed this strategy during
a column seminar at McGill University in Montreal and within days, one of the attendees,
Roger Amelunxen, was back at Gibraltar Mines successfully operating a “homemade” column.
Many column designs have been proposed in the West and in Russia (Rubenstein
1995), but the one that survived is that of Boutin and Tremblay. The deciding factors were
probably the simple design, basically an open vessel, coupled with the positive bias strategy.
The invention germinated in trials using columns for a solvent-in-pulp process (P. Boutin,
personal communication). Entrainment of slurry in the rising droplet wake was solved by
the addition of clear aqueous phase just below the interface. The inventors were quick to
realize that flotation offered more potential. Given the role of “oil” in the early history of flotation, it is interesting to see it as a source of inspiration here. The inspiration continues with
a recent proposal to combine flotation and solvent extraction (Chen et al. 2003).
The commercial road was not smooth. Some 17 years after founding Column Flotation
of Canada Ltd., the first industrial installation was recorded at Les Mines Gaspé (Cienski
and Coffin 1981). This may seem a long gestation period but is about the norm for the minerals business (Napier-Munn 1997). After 1981, progress was rapid. A scale-up methodology was developed (Dobby and Finch 1986) that was first used to design the columns at
Mount Isa Mines (Espinosa-Gomez et al. 1989). By the early 1990s, three additional companies were marketing flotation columns based on the Boutin–Tremblay design, which was
becoming known as the conventional or Canadian column. The main suppliers today (in no
particular order) are MinnovEX, CPT, Dorr-Oliver-Eimco, Cisa, Control International,
Multotech, RIF, Metso, and Dual Extraction. One estimate is that there are some 3,000 columns installed worldwide, about 30%–40% being homemade, ranging from the minerals
industry to the offshore oil industry (the flotation capacity in the oil industry may actually
exceed that in the minerals industry).
The success of columns helped usher in other novel developments in flotation machines
(Finch 1995). Some were aimed at overcoming the well-known problem of pneumatic
cells—reliable gas injection. Jet-type spargers is one outcome. Dispersing air into slurry and
injecting the mix into the column is another. The patent of Hollingsworth (1968) ventured
this possibility, the Microcel being a commercial version (Yoon, Adel, and Luttrell 1992).
The high retention time in some columns (tens of minutes on occasion) is eliminated in the
Jameson cell, which exploits air dispersion/slurry contacting in a downcomer where retention time is just seconds ( Jameson 1988). The Voith Sulzer cell developed for de-inking
recycled paper shares some design features (Finch and Hardie 1999). Research focused on
understanding mechanisms has also led to a range of sensors finding application in mechanical cells (Gomez and Finch 2002), such as optimizing gas distribution to banks of cells
(Cooper et al. 2004).
The conventional cell manufacturers were not idle during this period. From the “virtual
columns” of the early 1990s (Finch 1995), a range of competing tank cell designs is now
available, which, when viewed from a certain angle, reveal their inspiration.
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HISTORICAL ASPECTS OF FLOTATION
Tank
Air
Feed
Flotation
Tank
Jet Chamber
Nozzle Body
Baffle
Tailing
Deaeration Zone
Source: Cusack 1968.
FIGURE 6
Davcra flotation cell
FIGURE 7
XPM-series flotation machine
Flotation columns have taken their place among the options in cell selection. Overused
in the initial flurry perhaps, they have retained a place in many circuits for most mineral
commodities.
High-Intensity Flotation Cells
High-intensity cells use very small bubbles produced by forced air (e.g., Davcra, Bahr, and
Microcel cells) or induced air (e.g., Jameson and XPM cells). The Bahr cell, which was
invented in 1974, was a column fed by a high-pressure air–pulp mixture, and this cell started
a trend to combine the best ideas of columns and high-intensity cells into single machines.
The first reported high-intensity cell was the Davcra cell (shown in Figure 6) devised by
Bill Davis at the Zinc Corporation in Broken Hill and tested there in 1966. The principle
was that most particle–bubble interactions are time independent, and recovery depended
mainly on the characteristics of the intensely mixed zone (Davis 1966). The cell worked by
air and feed slurry being injected into the tank through a dispersion nozzle, with energy
being dissipated via collision with a vertical baffle, as shown in Figure 6 (Cusack 1968). The
Davcra cell was used for some years in plants for floating sulfide minerals and coal.
In China, Professor Daiwei Wu and colleagues developed jet flotation cells and started
using them in an industrial plant in 1967. The XPM flotation machine is similar to a
mechanical flotation machine, but the rotating mechanisms are replaced by jets of pulp and
air that form “aeration-agitation” zones. Part of the pulp within each cell is drawn along with
air into a circulating pump, and the mixture is pressurized and squirted from a conical jet, as
shown in Figure 7. Froth forms rapidly and is removed by scrapers. These cells are used in 14
Chinese coal-preparation plants (Wu and Ma 1998). The largest cells are now 23 m3.
For the past 30 years, extensive development of high-intensity flotation machines has
been conducted in Germany with much of the original work undertaken by Professor Bahr
at the Technical University of Clausthal. The Bahr cell shown in Figure 8 was one of the initial
flotation devices developed from this work. The Bahr cell uses aerator units in which compressed
air flows through small openings via channels into the pulp. The aerator units were located
beneath the main flotation tank and entered the tank vertically (Cordes 1997). Since its
development, many derivations and variations of the Bahr cell have been produced under a
variety of names including Ekoflot, Pneuflot, Allflot, and Imhoflot.
The Jameson cell, shown in Figure 9, was developed by Graeme Jameson of the University
of Newcastle (Australia) and the technology was developed for commercial application by
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HISTORY OF FLOTATION TECHNOLOGY
81
Air
Air
Feed
Tailing
Concentrate
Feed
General arrangement. The
tank contains several feed
pipes with their aeration zones.
Typical feed pipe showing
the aeration zone.
Source: Truscott 1923.
FIGURE 8
Bahr cell
Pulp Feed
Induced Air
Downcomer
Nozzle
Free Jet
Plunging Jet
Recirculating Eddy
Mixing Zone
Submerged Jet
Pipe Flow Zone
Concentrate
Disengagement Zone
Pulp Mixture
Cell
FIGURE 9
Tailings
Jameson cell
Mount Isa Mines Limited. Its principles of operation have been discussed by numerous authors
including Jameson (1988); Jameson et al. (1988); and Evans, Atkinson, and Jameson (1995).
The high-intensity contacting zone is the downcomer. Feed pulp is pumped into the
downcomer through an orifice plate, creating a high-pressure jet. The plunging jet of liquid
shears and then entrains air, which has been naturally aspirated. Because of a high mixing
velocity and a large interfacial area, there is rapid contact and collection of particles. One
unique feature of the Jameson cell is the operation of the downcomer under a vacuum,
which results in a high-intensity contact residence time that varies from 1 to 10 seconds.
Since its invention in 1986, there have been 225 Jameson cells installed in variety of coal,
metalliferous, and industrial mineral applications.
The Microcel was developed by Yoon, Adel, and Luttrell (1992) at the Virginia Polytechnic Institute. In the Microcel flotation machine, the slurry is mixed with small bubbles in
the microbubble generator, which is outside the column, and separation occurs inside the column.
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HISTORICAL ASPECTS OF FLOTATION
Wash Water
Distributor
Froth Product
Slurry
Air
Feed Inlet
Slurry Manifold
Wash Water Inlet
Pressure Transducer
Frother Inlet
Inlet from Pump
Bubble
Generators
Tailing Outlet
Control Valve
Outlet to Pump
Microbubble
Suspension
Source: Yoon, Adel, and Luttrell 1992.
FIGURE 10
Microcel
The microbubble generator is a high-intensity bubble-contacting zone formed by a
static in-line mixer; slurry is drawn from the base of the column and pumped through the
generator; pressurized air is injected into the mixers, and the high resultant shear forces create
fine bubbles (see Figure 10). The operation of the Microcel has been described in several
publications, for example, Phillips et al. (1997) and Brake (1998). There are more than 100
installations in mineral and coal plants worldwide.
O N - S T R E A M A N A LY S I S
A seminal change in flotation technology occurred when on-stream analysis (OSA) systems
were developed. These enabled the metal contents of streams to be measured on-line and
circuit grades and recoveries to be calculated every few minutes. In the days before OSA, flotation was an art, and results depended on the operator’s skills in observing the froth, using
the panning dish, and controlling air and reagents manually. After OSA became available,
flotation could be monitored accurately and controlled. Its development in the 1960s was
timely; cell size was about to grow rapidly, and there was no alternative sensor for process
control. OSA by itself was not enough. In the early 1960s, digital computers became available, and it was the OSA–computer link that changed flotation into an advanced technology.
Setting the Scene
The problem with flotation for many years was that there was a long delay between taking
samples from circuits and obtaining the assays, so the results were of historical value only
and were not useful for circuit control. Better control could only be achieved if the delay was
reduced to minutes, but rapid analysis of elements such as lead and copper in circuit streams
was not possible until particles could be analyzed in the pulps rather than by sampling, drying, and use of conventional wet methods.
In the 1950s, it was realized that X-ray techniques offered the most promising
approach, and there seemed to be only one likely candidate: X-ray fluorescence (XRF) analysis by wavelength dispersive methods based on X-ray tube and (Bragg) crystal spectrometer. In the 1960s, an alternative approach was developed of using radioisotope X-ray
techniques based on gamma-ray preferential absorption, or XRA, and energy dispersive
(XRF) analysis.
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HISTORY OF FLOTATION TECHNOLOGY
83
Developments Required
In 1960 the X-ray and radioisotope approaches required completely different developments
before a practical OSA system could be realized.
• The X-ray tube and crystal spectrometer system was relatively complex and expensive
but was known to give accurate results because the wavelength dispersive system
could resolve fluorescent X-rays from adjacent atomic number elements. The equipment
had to be mounted in a central location of the plant, where it would sequentially
analyze continuous samples taken from each process stream. This involved accurate
sampling of each process stream, long runs of pipelines to the central analyzer, pumping, sample splitting, constant head tanks, and flow cells.
• The radioisotope techniques of analysis involved the use of relatively inexpensive
source-detector systems. Hence, one or more head units could be placed near each
plant process stream, with only a short sample by-line between process stream and
analysis system being required so that construction problems were simpler. The two
radioisotope techniques were at different stages of development:
1. XRA techniques: These were suitable for high atomic number elements such as
lead, uranium, tungsten, and bismuth. The problem was that there were only a few
radioisotope sources emitting gamma rays of suitable energy, and even fewer were
commercially available.
2. XRF techniques: These were essential for medium atomic number elements such
as iron, nickel, copper, zinc, and tin. The problem in this case was that the detectors that were available could not resolve fluorescent X-rays from adjacent atomic
number elements.
X-ray Tube Systems
Plant tests were carried out in the late 1950s using laboratory X-ray tube systems for off-line
measurement of dry plant samples and on-line measurement of plant pulps. The systems
worked well. The first on-line system was tested during 1959–1960 in the 36,000-tons-perday concentrator of the Anaconda Copper Corporation in Butte, Montana. The X-ray system was satisfactory, but problems occurred with sample handling and presentation, which
took months to solve. For example, “Sufficient wood is present to completely stop all flow
through the X-ray head. Various types of screens were tried before finding a satisfactory
solution. This was typical of the type of mechanical problem that plagued and delayed the
final process control by X-ray analysis” (Lucy, Fulmore, and Holderreed 1963, p. 682). The
system presented assays of 13 streams every 20 minutes. Another system was built in 1962
by the Research and Instrumentation Division of Rhoanglo Mine Services Ltd. in Northern
Rhodesia (now Zambia). It was installed in the Bancroft concentrator and presented copper
assays of 6 streams every 8 minutes (Barlin and Keys 1963).
In 1962 the Outokumpu Group in Finland started what became a successful program
of X-ray tube on-stream analyzers. In that year, the company established an Institute of
Physics and brought the Pyhasalmi multimetal mine into operation. The mine became a
large-scale laboratory in which instruments developed by the institute were tested, and this
was an important factor in the success of the program. The objectives of the institute were to
develop electromagnetic and XRF technology, and Professor Pekka Rautala became its
director. Ore from the Pyhasalmi mine was floated to produce lead, copper, zinc, and pyrite
concentrates, so there was a broad scope for experimentation.
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HISTORICAL ASPECTS OF FLOTATION
The institute developed a 14-stream, wavelength dispersive X-ray analyzer, which was
installed in the Pyhasalmi plant in 1968, and its success led to the installation of a second
analyzer in 1970 (Lahteenmaki, Miettunen, and Saloheimo 1999). Assays of each stream
were available to the operators every 6–7 minutes; the cost per assay was about 2 cents, and
the analyzers had a high availability that was usually more than 99%. Using this analyzer,
Outokumpu developed its own process control and management system and installed the
first at the Kotalahti nickel concentrator in Finland in 1973. The analyzers have been continually improved, and current models can be installed in the process area instead of in a separate room. By 2004 more than 400 Outokumpu X-ray tube analyzers had been sold for use
in mineral concentrators worldwide (M. Kongas, personal communication).
Radioisotope Systems
Radioisotope systems offered the possibility of simpler and less expensive OSA compared
with the X-ray tube systems. The plant system could be built up in stages as needs arose.
Short sample by-lines would reduce the problems of pipe blockage caused by coarse particles
and wood chips, and wear on detector windows would also be reduced. The critical link
between the mineral industry and expertise in radioisotope X-ray techniques was established in 1962 when North Broken Hill Ltd. (NBH) approached the Australian Atomic
Energy Commission (AAEC) with their requirements for OSA. The approach by Conzinc
Riotinto of Australia (CRA) to the AAEC in 1965 resulted in the highly productive collaboration between physicists at the AAEC led by John Watt and metallurgists at CRA led by
Bruce Rawling.
Gamma-ray Transmission for High Atomic Number Elements
Gamma-ray transmission was the first technique developed for OSA using radioisotopes. In
1957 the AAEC was investigating a reactor system based on uranium powder suspended in
liquid sodium, and the suspension was simulated by tungsten powder in water which was
pumped around a 25-mm cross-sectional loop. The concentrations of tungsten over a cross
section of pipe were determined by scanning the gamma-ray beam (thulium 170, 84 keV
gamma rays) over the pipe cross-section. This was the first radioisotope OSA system of slurries (Watt and Lawther 1958).
In 1962 NBH asked the AAEC whether it was possible to continuously determine the
lead concentration of their flotation feed slurry on-line in a 150-mm-diameter steel pipe.
Calculations showed that this should be possible by combining measurements of gamma-ray
transmission at two different gamma-ray energies, about 200 and 662 keV. The AAEC overcame the lack of a suitable 200-keV radioisotope by developing a novel source based on
Compton scattering of higher-energy gamma rays yielding an output of about 225-keV. This
dual-energy gamma-ray transmission technique was successfully tested in the NBH plant in
1964 and 1966 (Ellis et al. 1967). The radioisotope system was installed on-line at the NBH
concentrator in 1968. This was the world’s first permanent installation of a radioisotope
OSA system in a mineral concentrator.
In December 1965, Bruce Rawling asked John Watt about measuring both lead and
zinc in sample by-lines from various process streams in a CRA concentrator at Broken Hill.
Watt’s response was that lead in flotation feed would be accurately determined by dualenergy gamma-ray transmission based on radioisotopes Gd-153 (100 keV) and Cs-137
(662 keV). For tailings, a correction would have to be made for matrix variations by a further
transmission measurement with gamma rays of suitable energy. These predictions were later
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HISTORY OF FLOTATION TECHNOLOGY
85
confirmed by calculation and experiments on samples of solids taken from plant streams
(Watt 1967; Ellis et al. 1969). Philips Industries Pty. Ltd. provided laboratory equipment to
CRA for sample by-line trials, and in 1968 CRA staff proved that lead could be determined
accurately in a sample by-line from the main flotation feed stream (Hinckfuss and Rawling 1968).
XRF for Medium Atomic Number Elements
The discussions between Rawling and Watt in 1965 crystallized thoughts on the urgency of
developing radioisotope XRF techniques for OSA for medium atomic number elements
such as copper and zinc. Australian mineral companies had to be contacted to obtain comprehensive information about their requirements, and samples had to be collected over
extended periods of time from various process streams in several concentrators so that the
AAEC could determine whether the techniques provided sufficient accuracy to meet these
requirements. The AAEC contracted Australian Mineral Development Laboratories
(Amdel) in 1966 to undertake the survey of mineral company requirements for OSA, and
this was completed in mid-1967. Over the period from 1966 to 1968, mineral companies
supplied the AAEC with suites of about 25 samples from each of several process streams in
their plants that were taken over a period of at least 6 weeks.
During 1966–1968 the AAEC undertook extensive development of radioisotope XRF
assemblies and techniques (Watt and Gravitis 1973; Watt 1983). In 1966, XRF measurements were made on samples of lead/zinc ore, taken from widespread locations throughout
the CRA mine, with excellent results of zinc in the range 0–34 wt % being determined to
0.6 wt % (1σ) (Watt 1967). Measurements on the samples taken from several streams in
each of six concentrators also gave promising results (Ellis et al. 1969) with one exception:
copper in the iron-rich tailings from Tennant Creek, which was later solved with the development of the detector-radiator assembly (Watt 1972). The success of laboratory measurements on samples was followed up with on-stream trials of radioisotope XRF systems at five
mineral concentrators, undertaken during 1968–1971 by AAEC, Amdel, and plant staff
(Fookes et al. 1971). Overall, these trials were very successful and led to improvements in
radioisotope X-ray assemblies and techniques.
During 1967 Douglas Hinckfuss of CRA proposed replacing XRF measurements on a
sample by-line with measurements by probes directly immersed into the plant process
stream (Hinckfuss 1972). The probe was a casing containing the radioisotope source and
detector assembly. The immersion probe, shown in Figure 11, was a key development to the
radioisotope OSA system because it overcame the need for use of sample by-lines and dramatically reduced window wear (Fookes et al. 1973). There was now a complete contrast in
approaches to OSA by the radioisotope and the X-ray tube systems. Joint CRA–AAEC trials
Density Probe
Pb Probe
Pulp In
Zn Probe
To Adjacent
Flotation Cells
Existing Slurry
Vessel
Source: Cutmore et al. 1993.
FIGURE 11
Radioisotope probes immersed in a plant mineral slurry
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HISTORICAL ASPECTS OF FLOTATION
at Broken Hill using immersion probes demonstrated excellent results for lead and zinc
determined in plant process streams (Fookes et al. 1973). Stump and Roberts (1974) demonstrated better control of grinding and flotation at the New Broken Hill Consolidated
concentrator based on using radioisotope probes and computer control, as well as excellent
accuracies of the OSA for lead and zinc in actual plant installations.
From 1967 the AAEC also used silicon solid-state detectors for laboratory XRF measurements on mineral samples from plants. These detectors had good X-ray energy resolution and proved to be very suitable for the analysis of samples from process streams
(Gravitis, Greig, and Watt 1974). These detectors were then not sufficiently stable for use in
industrial plants. In the early 1980s, Amdel incorporated these detectors into immersion
probes for plant use in tailing streams.
Commercial System
The radioisotope XRF and XRA assemblies developed by the AAEC, incorporated into the
casing of the immersion probe developed by CRA, became the basis of the commercial OSA
system. Philips Industries Ltd. was selected as the licensee to design and manufacture the
commercial system hardware, and Amdel was selected to undertake the feasibility studies,
installation, and calibration. Amdel installed the first three plant-analysis systems in concentrators in 1973. The development of the radioisotope on-stream analysis system had been a
productive 10-year project that contributed much to flotation technology.
Amdel took over the system manufacture in 1978. Thermo Electron Corporation took
over the Amdel Instrumentation Division, including staff, in 1999 and continues to manufacture the radioisotope OSA system in Australia and market it worldwide. By 2003 about 170
radioisotope X-ray systems had been sold worldwide for mineral processing operations.
Automatic Control of Flotation Circuits
By 1975 the problem of long delays between taking samples in plants and receiving assays
had been solved; flotation circuit performance could be assessed every few minutes, and
automatic control systems could be developed for flotation circuits that would take into
account changes in mineral contents and floatabilities. Multistream X-ray tube systems and
radioisotope systems were being used to monitor and control flotation circuits in several
concentrators.
The objective of automatic control was and still is to operate each flotation circuit at
the point on its optimum grade-recovery curve that gave the best economic results. Some of
the approaches tried in early control systems were
• Controlling reagents by the feed grade and/or the concentrate grade
• Automatic raising and lowering of concentrate diverter trays on rougher banks to
maintain stable performance of cleaner banks
• Controlling reagents by making incremental changes and searching for optimum circuit performance
These early systems were designed for specific cases, because no two ores are identical,
and they usually improved circuit performance by reducing variations in concentrate grades
and increasing recoveries of valuable minerals. They also improved the skills of operators
who were given much more information about the circuit by computer-based data logging
systems than was available through observation.
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HISTORY OF FLOTATION TECHNOLOGY
87
The problem with control systems in those early days was that if the ore changed significantly, it was difficult to determine on-line the new grade-recovery curve and the best operating point, so the control target was difficult to define. There has been some progress on
this problem, mainly through the use of various mathematical techniques. Software to control flotation circuits is commercially available.
O T H E R U S E S O F F L O TAT I O N
Over the decades, flotation has found uses in areas far removed from mine sites. It is used for
the removal of solids in wineries, breweries, butter and cheese factories, dairies, and sewerage
plants, among others. It is commonly used as a means for removing oil and contaminants
from water in smelters and refineries, as well as for removing algae and other organic contaminants from water. Other areas in which flotation can be found are the de-inking of recycled
paper, treatment of abattoir and sawmill effluent, sugar milling and refining, wool scouring, the
production of vegetable oil and margarine, paint manufacturing, and paper processing.
POSTSCRIPT
The Centenary of Flotation Symposium celebrated the life of a remarkable process and the
achievements of scientists and engineers in many countries who overcame early problems
and made flotation an advanced technology. Theodore Hoover wrote of the difficulties of
developing “concentration upside down,” as flotation was described (Ingalls 1907), into an
acceptable and standard process: “The only previous authentic case where substances
heavier than water have been made to float was the occasion of Elisha’s miracle with the axe
(2 Kings 6) and mining and metallurgical engineers are not great believers in miracles”
(Hoover 1914). But the engineers were never daunted. Flotation was very difficult to operate in its early years because of the variable nature of the ores, yet from its inception, engineers have continued to extend its limits of application. The variety and extent of its uses
today could not have been imagined 100 years ago when thin scums of zinc concentrates
were being floated on top of hot, acid pulps. Progress since then has been due to the cumulative achievements of a legion of engineers working on flotation processes in many countries.
There will be many advances yet in flotation technology as its capabilities for separating
valuable and waste materials are used more extensively. It is hoped that this brief history
gives incoming flotation engineers and scientists some of the fascinating background of the
process with which they will be working.
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———. 2002. The froth phase in column flotation. Pages 161–168 in Flotation and Flocculation,
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Jameson, G.J., M. Belk, N.W. Johnson, R. Espinosa-Gomez, and J.P. Andreaditis. 1988. Mineral
flotation in a high intensity column. Pages 507–510 in Chemeca 88, 16th Australian Conference
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Klassen, V.I., and V.A. Mokrousov. 1963. An Introduction to the Theory of Flotation. London:
Butterworths.
Lahteenmaki, S., J. Miettunen, and K. Saloheimo. 1999. 30 years of on-stream analysis at the
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Napier-Munn, T. 1997. Invention and innovation in mineral processing. Miner. Eng. 10:757–774.
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thereof. U.S. Patent 873,586. December 10.
Phillips, D.I., R. Yoon, G.H. Luttrell, L. Fish, and T.A. Toney. 1997. Installation of 4-meter diameter
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Potter, C.V. 1902. U.S. Patent 776,145. January 14.
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PART 2
Flotation Fundamentals
93
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Some Aspects of Flotation
Thermodynamics
D.W. Fuerstenau and S. Raghavan
A B S T R AC T
This chapter presents a brief summary of some of the thermodynamic aspects of flotation processes.
Thermodynamic considerations that control interfacial and wetting behavior in mineral–
water–air systems are discussed. Particular attention is given to the thermodynamics of collector
adsorption.
INTRODUCTION
The application of the principles of thermodynamics to flotation systems has contributed
significantly toward understanding some of the underlying foundations of flotation processes. When applied to physical and chemical systems, such as flotation, thermodynamics
includes several topics:
1. The conditions under which chemical substances, or different physical states of the
same substance, exist in equilibrium
2. Whether, under certain specified conditions, a chemical reaction or a phase change
will take place spontaneously.
3. The relation between the interchange of heat and other forms of energy when a
chemical reaction or phase displacement occurs
4. The effect of temperature on chemical reactions and phase equilibria
5. The principles underlying the methods of measurement of those properties whose
values are required for quantifying the foregoing
Thermodynamics is used to predict whether or not a change will tend to occur and yet
reveals nothing about the rate at which the change will take place. Though thermodynamics
cannot actually report to a flotation engineer about what the mineral recovery will be at a
given temperature or under given solution conditions, it can help the engineer make some
predictions as to how the flotation response may change with temperature, type of collector,
type of mineral, and so on. One criticism that has been leveled against the practical need for
studying the thermodynamics of flotation is that thermodynamics is concerned mainly with
equilibrium processes and the fact that, in the time span during which flotation takes place,
the system may not be in equilibrium. Interestingly, Wada (1960) actually defined flotation
as a thermodynamic process in which gas–liquid–solid interfaces participate in separating
finely divided solids from one another. In this chapter, an attempt is made to briefly review
various thermodynamic investigations and approaches undertaken to gain an insight into
those interfacial processes that play a dominant role in the flotation process. This review is
not considered to be exhaustive, but only summarizes some of the highlights.
95
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96
FLOTATION FUNDAMENTALS
T H E R M O DY N A M I C S O F S U R FA C E S
The enthalpy of the surface on a unit area basis, HS, is defined by
(EQ 1)
HS = ES + (PV)0 = GS + T · SS
where ES is the total surface energy, PV is the pressure volume, GS is the Gibbs surface free
energy, T is the temperature, and SS is the surface entropy per square centimeter of surface.
Because the PV term is negligible for a surface, the surface energy and surface enthalpy are
equivalent. The surface tension, γ, and surface free energy, GS, are defined by
∂G
G S = γ = ⎛ -------⎞
⎝ ∂A⎠ T, P, n
(EQ 2)
where A is the interfacial area, and n is the number of moles present in the system. These
quantities are usually given as ergs per square centimeter or dynes per centimeter, which are
identical in magnitude. Actually, only in a one-component system are GS and γ identical,
but for the purpose of this review, no distinction will be made between them. Because the
surface entropy at constant pressure is given by
∂G
dγ
S S = – ⎛ --------S-⎞ = – ------⎝ ∂t ⎠ p
dT
(EQ 3)
the relation between total surface energy and surface tension is
dγ
E S = γ – T ------dT
(EQ 4)
For most liquids, the surface tension decreases linearly with temperature (Adamson
1967). In the case of water at 20°C, γ = 72.75 ergs/cm2 and dγ/dT = –0.16, from which one
evaluates the total surface energy of water to be 120 ergs/cm2. In the case of octane, a typical
liquid hydrocarbon, γ = 21.80 ergs/cm2 at 20°C and dγ/dT = –0.10, from which ES is calculated to be 51.1 ergs/cm2.
In the case of solids, the evaluation of surface energies is less straightforward, both
experimentally and theoretically. Some typical values of the surface tension (and surface
energies) of selected materials are given in Table 1. These can only be considered approximate, but they enable one to see how widely solids do differ. Clearly, the type of chemical
TABLE 1
Approximate surface energies of some materials at room temperature
γ, ergs/cm2
25
110
230
450
1,000
1,800
1,900
5,600
Material
Paraffin
Graphite
Halite (NaCl)
Fluorite (CaF2)
Magnesia (MgO)
Gold
Alumina (Al2O3)
Diamond
ES, ergs/cm2
280
1,090
5,600
Source: Adamson 1967.
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
97
bonds that hold a crystal together have a marked effect on the magnitude of the surface
energy of solids. Finally, it should be added that mineral–water interfacial tensions are generally much lower than the surface energies, particularly for oxides.
T H E R M O DY N A M I C S O F A D S O R P T I O N
What happens in the case of a system with several components when one or more of the
components accumulates, that is, adsorbs, at an interface? The basis for the thermodynamics
of the adsorption of dissolved substances was unambiguously laid down by Gibbs. The
Gibbs adsorption equation relates the interfacial tension between two phases to the temperature, T, of the system; the chemical potentials of various species, μ1, μ2…μi, in the bulk; and
the surface excess or adsorption density of the various species, Γ1, Γ2…Γi, at the interface;
and has the following form (Defay and Prigogine 1966):
dγ = – S S dT – ∑ Γ i dμ i
(EQ 5)
i
There are several ways to define Γi, but the simplest way mathematically is to use the
convention proposed by Gibbs, namely, that the adsorption of the solvent is zero and that Γi
is the excess surface concentration. Excess surface concentration means the excess at the
interface over that which would be expected if the solution phase were uniform up to a
hypothetical plane surface that divides the two bulk phases in a heterogeneous system. If one
measures adsorption densities from differences in bulk concentrations in dilute systems (as
are usually encountered in flotation systems), then the Gibbs convention is most applicable.
Overall, the Gibbs adsorption equation is highly relevant to flotation through its application to phenomena involving frother systems, wettability, the development of surface charge
at mineral–water interfaces, and so on, because it quantitatively expresses the change in surface tension due to the adsorption of surface-active materials. By defining the adsorption of
the solvent (component 1) as zero, Equation 5 can be modified as
i
dγ = – S S dT –
∑ Γi( 1 ) dμi
(EQ 6)
i=2
where Γi(1) refers to the relative adsorption of component i at the dividing surface, such that
Γ1 = 0. Because flotation processes are often carried out at constant temperature, Equation 6
can be simplified to
i
dγ = – ∑ Γ i( 1 ) dμ i
(EQ 7)
i=2
Recalling that dμi = RT d ln ai, where ai is the activity of species i in the bulk aqueous
solution, R is the gas constant, and T is the temperature, Equation 7 can be rearranged to give
∂γ
1
Γ i( 1 ) = – -------- ⎛ ------------⎞
⎝
RT ∂ ln a i⎠ T, μj
;i ≠ j
(EQ 8)
The adsorption density, Γ(1), will be positive if ( ∂γ ⁄ ∂ ln a i ) T, μ ;i ≠ j is negative, and
j
vice versa.
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98
FLOTATION FUNDAMENTALS
The adsorption density can be calculated from the slope of the surface tension-versuslog activity (or concentration) curve if the surface tension of the particular interface under
consideration can be determined experimentally. To illustrate the use of the Gibbs equation
in flotation, first consider aqueous solutions of frothers, which are surface-active agents that
have a tendency to concentrate at the air–water interface. The Gibbs equation, when
applied to a dilute aqueous solution of a frother such as an alcohol (ROH) in water, has the
following form (de Bruyn and Agar 1962):
dγ = ( – RT )Γ ROH d ln a ROH
(EQ 9)
≈ ( – RT )Γ ROH d ln C ROH
The dependence of surface tension on the frother concentration and the adsorption
density of the frother at the air–water interface are shown in Figure 1 for the water–butyl
alcohol system. The lowering of the surface tension of water due to the addition of alcohol
shows that the alcohol is positively adsorbed at the water–air interface. For example, at an activity of 0.712 (0.854 molality), the adsorption density of butyl alcohol is 6.03 × 10–10 mol/cm2.
At this concentration, the area per molecule is 27.4 sq Å. Close-packed films of longchained carboxylic acids and alcohols exhibit areas per molecule of 21.6 sq Å, suggesting
that the adsorbed butyl alcohol on water is monomolecular. For proper frothing, there must
80
Surface Tension, dynes/cm
72.8
60
Butanol/Water
20°C
40
A
20
120
6
Γ, mol/cm 2 × 10 10
100
80
4
60
2
40
20
Area per Adsorbed Molecule, sq. Å
140
B
0
–3
0
–2
–1
0
log a2
Source: de Bruyn and Agar 1962.
FIGURE 1 Adsorption of butanol at the aqueous solution/gas interface: (a) Surface tension
and (b) adsorption density and area per adsorbed molecule versus log butanol activity
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
99
be a surface tension–concentration gradient such as that shown in Figure 1 (i.e., dγ/dC ≠ 0),
so that momentarily deformed films can withstand the shock.
The surface tension lowerings produced by the positive adsorption of a homologous
series of n-aliphatic alcohols at the water–air interface are most conveniently compared by
plotting the surface tension of the solution as a function of the logarithm of the activity (or
concentration) of the alcohol (Defay and Prigogine 1966). In the concentration range corresponding to the lower parts of a series of such curves, the slope is independent of the alcohol
considered, showing that the same number of molecules are adsorbed per unit area of the
surface. Further, the concentration ratio of two neighboring homologs is constant and
approximately equal to 3.2 for the same Δγ. In other words, the surface activity is approximately
tripled for each additional –CH2– group in the molecule. This is called Traube’s rule.
Collectors are heteropolar organic compounds whose main function is to adsorb at the
mineral–water interface, but they also tend to adsorb at the mineral–air and water–air interfaces. They differ from the frothers in the sense that they are generally electrolytes. Taking
into consideration the aqueous dodecyl sulfonate system, the change in the surface tension
of the solution–air interface is given by
dγ = – Γ R – dμ R – – Γ Na + dμ Na +
(EQ 10)
where R– refers to the dodecyl sulfonate anion (C12H25SO3–). In the absence of any other
electrolytes where Na+ and R– would be the only species that are adsorbing, it can easily be
shown that
dγ = – 2RT Γ R – d ln C NaR
(EQ 11)
If the organic electrolyte is a weak electrolyte, such as the salt of a fatty acid, then
depending on the pH of the solution, the adsorption of neutral collector molecules must
also be considered.
Probably the most marked lowering of surface tensions due to adsorption is that exhibited in solid–vapor systems. For example, water vapor adsorption can lower the surface tension of oxides by several hundred ergs per square centimeter. Adsorption phenomena at the
bubble–mineral interface appear to have a significant role in flotation.
T H E R M O DY N A M I C S O F W E T T I N G
Bubble–particle contact is one of the key factors controlling the process of froth flotation.
In this section, the thermodynamic aspects of the bubble–particle contact will be reviewed
and critically analyzed (Adamson 1967; Gaudin 1957).
The general thermodynamic condition for three-phase contact is defined by Young’s
equation for the system depicted schematically in Figure 2.
γ SG = γ SL + γ LG cos θ
(EQ 12)
where γSG, γSL , and γLG are the tensions of the solid–gas, solid–liquid, and liquid–gas interfaces, respectively, and θ is the contact angle. The change in the free energy accompanying
the replacement of a unit area of the solid–liquid interface by solid–gas interface is given by
Dupre’s equation, namely:
ΔG = γ SG – ( γ SL + γ LG )
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(EQ 13)
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100
FLOTATION FUNDAMENTALS
Liquid
γLG
Gas
θ
Solid
γSG
γSL
FIGURE 2 Schematic representation of the equilibrium contact between an air bubble and a
solid immersed in a liquid. The contact angle θ is the angle between the liquid–gas and the
liquid–solid interfaces, measured through the liquid.
Combination of Dupre’s equation with Young’s equation yields the following expression for the free energy change, namely:
ΔG = γ LG ( cos θ – 1 )
(EQ 14)
Thus, for any finite value of the contact angle, there will be a free energy decrease upon
attachment of a mineral particle to an air bubble.
The theoretical relationship of Dupre’s equation expresses the maximum possible
decrease in the free energy of the system resulting from the bubble–particle contact, which
can be realized only when there are no other energy-consuming effects, such as deformation
of the bubble. Thus, the geometry of the system is not taken into account in Dupre’s equation. Further, Young’s equation (Equation 12) is valid in an ideal system where all gravitational effects are absent and the system is at equilibrium ( Johnson 1959).
Often, considerable hysteresis exists in measured contact angles because of surface
roughness, contamination, nonequilibrium adsorption effects, etc. (Adamson 1967). If
there is hysteresis, that is, if a liquid-advancing contact has a different value than a liquidreceding angle, equilibrium is not attained in the system. Under these conditions, the use of
Young’s equation is not valid. Figure 3 presents equilibrium, advancing and receding contact
angles on alumina in sodium dodecyl sulfonate solutions at pH 7.2 as measured by Wakamatsu and Fuerstenau (1973). As can be seen in this figure, a pronounced hysteresis effect
exists in this system under the conditions of this investigation.
Leja and Poling (1960) conducted an interesting theoretical study on the attachment of
air bubbles to flat solids, both in the presence and absence of gravitational effects, and their
results will be summarized at this stage. Typically, because the solid surface is not of the same
contour as that of the air bubble, work must be expended by the system in deforming the
air–liquid interface during attachment, and the actual free energy of adhesion per unit area,
WP, is smaller than the theoretically available value, WA (equal to ΔG given by Equation
13). In the absence of other forms of energy (gravitational and kinetic), the work of deformation must be performed solely by the interfacial energy. The deformation is then governed by the shape of the solid surface and, for a perfectly flat surface, the magnitude of the
contact angle, θ, determines the extent of deformation that the interfacial energy pool, WA ,
is capable of performing.
Leja and Poling (1960) suggest that gravitational and kinetic energies (due to the
motion of particles and air bubbles) affect the energy expended in deformation of the air–
liquid interface, either decreasing or increasing the amount of deformation to be performed
by the interfacial energy pool. So when part of the deformation work is performed by gravitational or kinetic energy, hysteresis of the practical contact angle is observed, even when
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
100
Contact Angle, degrees
80
S
A
S
A
S
A
Liquid
Receding
Equilibrium
Liquid
Advancing
101
60
Advancing
40
Equilibrium
Alumina
Ionic Strength 2 × 10–3N
pH 7.2
24±1°C
20
Receding
0
10–6
10–5
10–4
10–3
Concentration of Dodecyl Sulfonate, mol/L
Source: Wakamatsu and Fuerstenau 1973.
FIGURE 3 Equilibrium, and receding and advancing contact angles on alumina in aqueous
dodecyl sulfonate solutions at pH 7.2
other effects, such as surface roughness, surface contamination, and so forth, are excluded.
According to Leja and Poling, an advancing contact angle indicates that a greater portion of
WA is being converted to the actual adhesion, WP, than under the condition of equilibrium.
On the other hand, a finite receding contact angle indicates a lower WP. Their conclusion
was that the contact angles, which were determined experimentally with fairly large bubbles
or drops, should not be used in Young’s equation unless a suitable correction in their magnitude is made to account for gravitational effects, a correction that becomes particularly significant at the threshold of hydrophobic character (i.e., with small contact angles).
A limited amount of work has been carried out on the temperature dependence of contact angles. The temperature coefficient of the contact angle can provide thermodynamic
information about wetting processes. For example, knowledge of the temperature coefficient
of contact angle provides a means of calculating the heat of immersion, which is the enthalpy
change upon immersing a clean, dry solid into a liquid. It can be shown (Adamson 1967) that
d cos θ
– ΔH imm = E LG cos θ – Tγ LG -------------dT
(EQ 15)
where ELG is the total surface energy of the liquid–gas interface. For an excellent account of
the temperature dependence of contact angles on low-energy surfaces, refer to an article by
Neumann (1974).
T H I N F I L M P H E N O M E N A A N D F L O TAT I O N
The well-known classical condition for the possibility of a bubble–particle contact in a
given liquid medium has been specified by Equation 13 where
ΔG = γ SG – γ LG – γ SL < 0
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102
FLOTATION FUNDAMENTALS
This relation assumes that there is no liquid layer between the bubble and the particle
after contact. This equation ignores the possibility of the existence of a thin liquid layer
between the bubble and the solid even after attachment. The thinning of this layer is
believed to control flotation (Laskowski 1974). Laskowski has made an attempt to modify
this equation for the presence of a liquid layer by making the following substitutions:
1. GS(h) for γSG where GS is the free energy of a column of liquid layer (of height h)
and solid of unit cross section, and
2. GS(h) = γSL + γLG(h). γLG(h) is the surface tension of the liquid–gas interface
expressed as a function of the thickness of a liquid layer h.
These substitutions lead to
ΔG ( h ) = γ LG ( h ) – γ LG
(EQ 16)
where ΔG(h) expresses the energy barrier in the transition from no contact to bubble contact, which must be overcome for attachment to occur (Laskowski 1974).
Derjaguin (1932) suggested that the thermodynamic properties of thin films are different from those of the bulk phase and introduced the parameter “disjoining pressure,” Π, as a
measure of the corresponding change in the thermodynamic properties. In effect, it is the
change of free energy with thickness and is given by the expression
∞
γ = γ o + ∫ Π dh
(EQ 17)
h
where γ is the specific-surface free energy of the thin liquid film, γo is the specific-surface
free energy of an infinitely thick liquid film, h is the thickness of the film, and Π is the disjoining pressure. If p is the vapor pressure in equilibrium with a flat thin film, and po is the vapor
pressure in equilibrium with the bulk liquid, then Derjaguin and Shcherbakov (1961) showed
p
∂G
RT ln ----- = V m ⎛ -------⎞
⎝ ∂h ⎠ A, T
po
(EQ 18)
where Vm is the molar volume of the liquid and the term ∂G/∂h is the disjoining pressure, Π.
If Π is positive, the thin film is stable. (For curved liquid–air interfaces, the above expression
has been suitably modified by Padday [1970]). Depending on the liquid and the surface, Π
can be positive or negative and can change sign with film thickness (Clifford 1975). For
liquid–solid systems with a finite contact angle, Π must be negative for certain film thicknesses (Clifford 1975).
A complete understanding of the mechanism of attachment of particles to air bubbles
in flotation should be based on the analysis of changes in the surface free energies of thin
films of water between the solid and the air bubbles. Rehbinder (1949) made some theoretical calculations to determine the excess free energy of thin films. In the thinning of the water
layer on the approach of a solid surface to an air bubble, the excess free energy of the layer,
according to Rehbinder, changes in relation to the initial hydration of the surface. When the
surface is highly hydrated (threshold hydrophilicity), the free energy of the liquid film
increases continuously during the thinning of the liquid film, thus preventing spontaneous
attachment of the air bubble to the surface. When the hydration of the surface is low, the
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
103
free energy of the thin film does not increase on initial thinning of the film, thus favoring
spontaneous attachment of a bubble to a mineral. The previous two cases are represented
schematically in Figure 4. Curve 2 in this figure is most typical of practical flotation systems.
In the case of strongly hydrophobic solids where water vapor adsorption is far lower than a
monolayer, the film would certainly be unstable and would rupture spontaneously.
The disjoining pressure can be due to many effects. It is usual to consider that these
effects are additive (Sheludko 1967) and that the disjoining pressure can be broken down
into a number of components (Padday 1970):
Π = Π LL + Π SL + Π el + Π h + Π p + Π i
(EQ 19)
Free Energy of a Thin Hydrated Layer, γ
where ΠLL is the disjoining pressure of the liquid film in the absence of a solid, ΠSL is the
disjoining pressure due to the effect of the solid on the liquid film, Πel is an electrostatic
term to account for double-layer effects, Πh is the contribution from hydrogen bonding, ΠP
is from polar interactions, and Πi is the change in free energy due to any other specific interactions between the liquid layer and the surface. Chander and Fuerstenau (1972) have made
an attempt to explain the natural floatability of molybdenite using this approach.
In 1960, Eigeles and Volova commented that “the experimental data on collector action
suggest that there may exist forces which are not taken into account by the present-day theory of coagulation of hydrophobic colloidal sols but exert a major influence on the kinetics
of flotation adhesion.” They added that “this unaccounted factor is what causes sharp acceleration of the thinning of the boundary layer and a pronounced increase of its fluidity.” This
led to their conducting a detailed investigation of the effect of temperature on induction
time, in which they observed a regular decrease in induction time with increasing temperature. From their experimental observations, they evaluated an apparent activation energy for
film thinning and attachment processes. The additional force suggested for bubble–particle
1
2
3
Δγ
Distance from the Surface
Source: Rehbinder 1949.
FIGURE 4 Change in the free energy of a water film between a gas bubble and solid surfaces
with differing hydrations: (1) maximum surface hydration, (2) moderate surface hydration, and
(3) very weak surface hydration
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FLOTATION FUNDAMENTALS
attachment processes in flotation by Eigeles and Volova was first observed and measured by
Israelachvili and Pashley in 1982 between mica surfaces coated with cetyltrimethylammonium bromide, using the surface force apparatus (Israelachvili and Pashley 1988). This
force, which is caused by the disruption of normal hydrogen bonding in water near a hydrophobic surface, is now called the hydrophobic force and would be an additional component
in Equation 19. Yoon (2000) has made extensive measurements of hydrophobic forces and
applied the concept to bubble–particle interaction in flotation.
T H E E N E R G E T I C S O F F L OATA B I L I T Y
Consider the conditions for floatability, namely, a finite contact angle of an air bubble on
the solid to be floated, in terms of work of adhesion of water to the solid (Wad ) and the work
of cohesion of the liquid (Wco). Because Wco = 2γLG (the surface tension of the liquid,
which is usually only slightly smaller than that of pure water) and Wad = γSG + γLG – γSL ,
assuming that the contact angle is given by Young’s equation (cos θ = [γSG – γSL]/γLG), it can
be shown that Wad – Wco = γLG (cos θ – 1). Hence, flotation is possible if Wad < Wco
(≤144 dynes/cm). It should be remembered that Wad is correctly defined as the work
required to separate liquid water from a solid–water interface, leaving behind an adsorbed
layer in equilibrium with saturated water vapor (Mellgren et al. 1974). After detailed experimentation with methylated silica, Laskowski and Kitchener (1968) analyzed hydrophobicity (floatability) in terms of these relations and concluded that all solids would be
hydrophobic if they did not carry polar or ionic groups on their surface. They also emphasized that hydrophobicity arises from the exceptionally large cohesive energy of water that is
due to hydrogen bonding between water molecules.
The magnitude of Wad depends on γSL which reflects the extent of interaction between
the solid and water. Only a few solids are naturally hydrophobic and therefore respond to
flotation without adding a collector. These solids include such materials as graphite, sulfur,
talc, molybdenite, and stibnite (Gaudin 1957). Hydrophobic materials such as graphite, in
principle, do not exhibit significant polar interactions with water, and the energetics of the
solid–liquid interface are mainly controlled by dispersion forces. Less hydrophobic surfaces
have some polar interactions with water, and hydrophilic solids such as oxides exhibit strong
polar interactions with water and are covered with hydroxyl groups (e.g., silica). Because the
function of the flotation collector is to render the surface more hydrophobic by eliminating
or shielding the polar sites of the solid, investigation of the energetics of the solid–water
interface, both in the presence and absence of a collector, should help one to comprehend
the concomitant phenomena of flotation. Wad = (γLG(cos θ + 1)) can readily be obtained
for hydrophobic solids given that θ > 0. But for hydrophilic solids, θ = 0 (Young’s equation
is no longer valid under these conditions), and consequently, evaluation of Wad is not easy.
However, two methods (Mellgren et al. 1974) for measuring solid–water interactions have
been widely used: (1) adsorption of water vapor on solids and (2) calorimetric heats of
immersion in liquid water.
In the first method, the adsorption density of water vapor is determined as a function of
water vapor pressure from the dry state up to (or close to) saturation. The free energy of saturation of the solid with water (ΔGsat) is given by the expression
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
105
Po
ΔG sat = RT ∫ Γ d ln p ⁄ p o
(EQ 20)
o
where Γ is the adsorption density (expressed in moles per area), p is the water vapor pressure,
and po is the saturation pressure. With hydrophilic solids, the major contribution to ΔGsat
comes from the most strongly bound water (i.e., at low p/po) and the net value is much
greater than 144 ergs/cm2. Some reported values are –316 ergs/cm2 for quartz, –264 ergs/cm2
for calcite, and –240 ergs/cm2 for barite (Mellgren et al. 1974).
The water-vapor adsorption method has also been used to study changes in the wettability of surfaces induced by the adsorption of collectors. Hall, Lovell, and Finkelstein
(1970) determined the effect of adsorbed oleate on calcite and fluorite on the ΔGsat value.
Their results gave a ΔGsat value of –300 ergs/cm2 for clean fluorite and –50 ergs/cm2 for
oleate-covered fluorite.
Heats of immersion can provide some information on the energetics for systems in cases
of zero contact angles. On a unit area basis, the heat of immersion of a clean solid, ΔHimm, is
given by
ΔH simm = H SL – H S ≈ E SL – E S
(EQ 21)
The hydrophilic character of a surface is clearly revealed by a large exothermic heat of
immersion, which indicates a strong interaction with water molecules. Specifically, calorimetric heat measurements suggest that the first layer of water on polar solids is very strongly
bound. When approximately three monolayers of water vapor are preadsorbed on the solid
before immersion, the heat value falls to about 120 ergs/cm2, which is the surface enthalpy
of normal liquid water (Mellgren et al. 1974). Therefore, water molecules beyond the first
few layers behave similar to bulk water.
Griffiths (1973) analyzed the results of Wade and Hackerman (1964) on the effect of
the outgasing temperature on the heat of immersion of alumina (Al2O3) in water and concluded that about one-half of the heat of immersion of alumina is due to the physically
adsorbed water with the remainder being due to the reformation of surface hydroxyl groups
lost during heat treatment. To gain more insight into the nature of physically adsorbed water
layers, Griffiths studied the temperature coefficient of the heat of immersion of Al2O3 in
water and found it to have a near-zero value. He concluded that the water structure outward
from the surface does not gradually grade into that of bulk water. In his words, what exists
“is an inner film of strongly bound water that discontinuously ends with near normal water
structure (at least energetically) continuing on outward from the surface” (Griffiths 1973).
Numerous heats of immersion measurements of minerals have been reported in the literature (Healy and Fuerstenau 1965; Zettlemoyer 1968), and Table 2 presents a brief summary of some typical values.
Although values of ΔHimm are high for hydrophilic and low for hydrophobic surfaces,
there is no well-defined “critical heat of immersion” for development of contact angles.
Hence, the two approaches discussed in this section are still of limited value in flotation
research. One study that relates the heat of immersion of oxides to the point of zero charge
(which plays an important role in selective flotation and collector choice) is worth mentioning as a final point. Healy and Fuerstenau (1965) found a quantitative linear relation between
the heat of immersion and the pH of the point of zero charge for a number of inorganic oxides.
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106
TABLE 2
FLOTATION FUNDAMENTALS
Heats of immersion and contact angles of some materials in water*
ΔHimm, ergs/cm2
–Δ
6
26
260–370
400–570
550
530
650–900
Solid
Teflon
Graphon
Quartz
SnO2 (cassiterite)
TiO2 (rutile)
Fe2O3 (hematite)
Al2O3 (alumina)
Contact Angle, degrees
108
82
0
0
0
0
0
*The oxides were activated at 200°C.
C R I T E R I O N F O R F L O TAT I O N I N T E R M S O F C O L L E C T O R
ADSORPTION
Starting with the work of de Bruyn, Overbeek, and Schuhmann (1954), there has been a
growing interest in whether collector adsorption at the S–L, S–G, or L–G-interface is the
main factor controlling flotation behavior. Consider a collector x that is present at all three
interfaces. On the assumptions that the collector adsorption at each interface follows the
Gibbs equation and the contact angle of an air bubble on the solid is given by Young’s equation, the following relations can be written:
dγ LG = – RT Γ xLG d ln a x
(EQ 22a)
dγ SG = – RT Γ xSG d ln a x
(EQ 22b)
dγ SL = – RT Γ xSL d ln a x
(EQ 22c)
γ SG – γ SL = γ LG cos θ
(EQ 22d)
where Γxij represents the adsorption density of x at interface ij, ax is the activity of the collector in the bulk solution, γij is the tension of the interface ij, and θ is the contact angle of the
air bubble on the solid.
Assuming ax = Cx, which is the concentration x, by differentiating Equation 12 with
respect to ln Cx at constant P and T and applying relations 22a–c, the following result is
obtained (Smolders 1961).
dθ
γ LG sin θ --------------- = RT ( Γ xSG – Γ xSL – Γ xLG cos θ )
d ln C x
(EQ 23)
From Equation 23, the following results can be deduced:
1. If the angle of contact increases with the increase of solute concentration (dθ/dln ax
is positive), ΓxSG must be greater than ΓxSL + ΓxLG cos θ.
2. In cases where the contact angle does not vary with concentration (dθ/dln ax = 0),
ΓxSG must be equal to ΓxSL + ΓxLG cos θ.
3. If the contact angle decreases with increasing concentration (which may happen
because of the reverse orientation of the second layer of ionic surfactants at the
solid–liquid interface), ΓxSG must be less than ΓxSL + ΓxLG cos θ.
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
107
As early as 1954, de Bruyn, Overbeek, and Schuhmann stressed the first result obtained
from Equation 23, namely, that ΓxSG > ΓxSL (for θ < 90°). In an attempt to apply the analysis
of de Bruyn, Overbeek, and Schuhmann to a simple flotation system, Aplan and de Bruyn
(1963) undertook an investigation of the adsorption of hexyl mercaptan onto gold and the
flotation of gold particles from an aqueous solution of hexyl mercaptan using nitrogen saturated with both water and mercaptan vapor as the gaseous phase. They found that excellent
flotation occurs at 3% to 30% surface coverage (low ΓSL), substantially below an equilibrium concentration where an abrupt change in the adsorption density from solution takes
place. In addition, they observed ΓSG to be greater than ΓSL in the flotation range.
In 1968, Somasundaran, in an attempt to explain the effect of collector chain length on
flotation at concentrations below those necessary for hemimicelle association of the hydrocarbon chains at the S–L interface, stressed the need for looking at adsorption conditions on
the bubble and at the S–G interface, in addition to the S–L interface. From the results of his
investigation of the quartz-dodecylamine solution–gas system, he found that the amount of
collector adsorption at the S–G interface was approximately equal to the adsorption at the
S–L interface. Sandvik and Digre (1968), without directly measuring ΓSG and ΓSL , showed
that the adsorption of a collector on a solid surface (silica) increased in the presence of gas
bubbles. They proposed a bubble-transfer hypothesis, which envisages the collector as being
transferred from the bubble surface to the solid. Their hypothesis in fact stresses the importance of adsorption at the L–G interface. Pope and Sutton’s (1972) work did not confirm
this hypothesis given that they observed a decrease in collector adsorption density at the
solid after flotation. Finch and Smith (1972) found that the flotation recovery of quartz and
magnetite (in the presence of dodecylamine) as a function of pH can be correlated with the
variation in the surface pressure [γsolvent – γsolution], of the L–G interface with pH. They concluded that this correlation implies that the greater the adsorption of the collector at the
bubble surface (higher the surface pressure), the better the flotation. In a later investigation
on the quartz–dodecylammonium acetate and the magnetite–dodecylammonium acetate
systems, Finch and Smith (1975) found a decrease in the tenacity of the bubble–solid
attachment with decreasing surface tension of the liquid–air interface, thus casting a doubt
on their earlier conclusion. Bleier, Goddard, and Kulkarni (1976) later reported good correlation between the flotation recovery of quartz in the presence of amines and the decrease in
the surface tension of the water–air interface.
It should be stressed that the interface of primary importance to flotation is the solid–
liquid interface. The collector simply must adsorb at this interface in order to reduce molecular attractions between the solid and the liquid.
T H E E L E C T R I C A L D O U B L E L AY E R AT M I N E R A L – WAT E R
I N T E R FA C E S
Because adsorption phenomena at mineral–water interfaces are controlled in most cases by
the electrical double layer, one must be concerned with factors responsible for the charge on
the solid surface and with the behavior of ions that adsorb as counterions to maintain electroneutrality (Kruyt 1952). Figure 5 presents a schematic representation of the electrical
double layer of counterions extending out into the liquid phase. This figure also shows the
drop in potential across the double layer, neglecting the potential because of dipole effects.
The closest distance of approach of counterions to the surface (δ) is the Stern plane.
Depending on whether ions remain hydrated or are dehydrated upon adsorption, there can
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108
FLOTATION FUNDAMENTALS
Surface
Charge, σo
Counterions, σd
δ
ψo
Solid
Potential
ψδ
δ
Distance
FIGURE 5 Schematic representation of the electrical double layer and the potential drop
across the double layer at a solid–water interface
exist an inner and an outer Stern plane (Grahame 1947); for purposes of this review,
though, no distinction shall be made between the inner and outer Stern plane. The surface
potential is ψo, at the Stern plane it is ψδ ; and from the Stern plane into the bulk of the solution, the potential drops to zero.
In the case of the ionic solids such as barite (BaSO4), fluorite (CaF2), silver iodide
(AgI), and silver sulfide (Ag2S), the surface charge arises from a preference of one of the lattice ions for sites at the solid surface as compared with the aqueous phase. Equilibrium is
attained when the electrochemical potential ( μ = μ + vF φ where φ is the Galvani potential in the phase) of these ions is constant throughout the system. Those particular ions that
are free to pass between both phases and therefore establish the electrical double layer are
called potential-determining ions. In the case of AgI, the potential determining ions are Ag+
and I–. For a solid such as calcite, CaCO3, the potential-determining ions are Ca2+ and
CO32–, and also H+, OH–, and HCO3– because of the equilibria between these latter ions
and CO32–. Similarly for apatite, the potential-determining ions are Ca2+, PO43–, and OH–,
with the other hydrolysis products also functioning in this role because of the complex equilibria involved in this system.
For oxide minerals, which will be discussed in more detail in the next section, hydrogen
and hydroxyl ions have long been considered to be potential-determining (Fuerstenau and
Healy 1972). In the layer silicate minerals such as clays and micas, because of the substitution of Al3+ for Si4+ in the silica tetrahedra or Mg2+ for Al3+ in the octahedral layer of the
crystal lattice, the surfaces of these crystal faces carry a negative charge that is independent of
solution conditions.
The surface charge, σs, on a solid in water is determined by the adsorption density of
potential-determining ions on the solid surface. In the case of a 1–1 valent salt, σs is given by
σS = F ( ΓM+ – ΓA– )
(EQ 24)
where F is the Faraday constant, ΓM+ is the adsorption density in moles per square centimeter of the potential-determining cation, and ΓA– is that of the potential-determining anion.
For an oxide, M+ and A– can be considered as H and OH–, respectively, and for AgI, M+
and A– are simply Ag+ and I–. By means of a simple titration procedure (Kruyt 1952; Parks
and de Bruyn 1962), the magnitude of the surface charge can be determined if the surface
area of the solid–liquid interface is known. The single most important parameter describing
the surface is the condition under which the surface charge, σs, is zero. The activity of potentialdetermining ions at which this occurs is called the point of zero charge, or PZC.
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
TABLE 3
109
PZC for some ionic solids
Material
Barite, BaSO4
Calcite, CaCO3
Fluorapatite, Ca5(PO4)3(F,OH)
Fluorite, CaF2
Hydroxyapatite, Ca5(PO4)3(OH)
Scheelite, CaWO4
Silver chloride, AgCl
Silver iodide, AgI
Silver sulfide Ag2S
PZC
pBa 6.7
pH 9.5*
pH 6*
pCa 3
pH 7*
pCa 4.8
pAg 4
pAg 5.6
pAg 10.2
Source: Fuerstenau 1971.
*The activities of the other potential-determining ions can be calculated from the hydrolysis equilibria and solubility data.
Assuming that potential differences due to dipoles, and so forth, remain constant, the
total double-layer potential or the surface potential, ψo, is considered to be zero at the PZC.
The value of the surface potential at any activity of 1–1 valent potential-determining
electrolyte is given by
aM+
RT
ψ o = -------- ln ---------------------F ( a M + ) PZC
(EQ 25)
where R is the gas constant, T is the temperature in degrees Kelvin, and ( a M + ) PZC is the
activity of the potential-determining cation at the PZC. Table 3 presents PZC values for a
number of ionic (salt-type) solids that have been investigated. Calcite, fluorite, and barite
are positively charged in their saturated solution at neutral pH, whereas the other materials
listed are negative, except for hydroxyapatite, which is uncharged at the pH shown.
The importance of the PZC is that the sign of the surface charge has a major effect on
the adsorption of all other ions and particularly those ions charged oppositely to the surface
because these ions function as the counterions to maintain electroneutrality. In contrast to
the situation in which the potential-determining ions are special for each system, any ions
present in the solution can function as the counterions. If the counterions are adsorbed only
by electrostatic attraction, they are called indifferent electrolytes. As has been well established
(Kruyt 1952), the counterions occur in a diffuse layer that extends from the surface into the
bulk solution. The “thickness” of the diffuse double layer is 1/κ, where κ is given by
κ = ( 8πF 2 v 2 C ⁄ εRT ) 1 ⁄ 2
(EQ 26)
where v is the valence of the ions (for a symmetrical electrolyte), and ε is the dielectric constant of the liquid. For a 1–1 valent electrolyte, for example, 1/κ is 1,000 Å in 10–5 M, 100
Å in 10–3 M, 10 Å in 10–1 M solutions.
The charge in the diffuse double layer σd given by the Gouy–Chapman relation (Kruyt
1952) as modified by Stern (for a symmetrical electrolyte) is
σ d = – σ S = – [ ( 2εRT ⁄ π )C ] 1 ⁄ 2 sinh ( vFψ δ ⁄ RT )
(EQ 27)
Further, if the potential does not change appreciably, this equation shows that the
adsorption density of counterions should vary as the square root of the concentration of
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110
FLOTATION FUNDAMENTALS
added electrolyte, as de Bruyn (1955) has found for the adsorption of dodecylammonium
acetate on quartz at low concentrations.
On the other hand, some ions exhibit surface activity in addition to electrostatic attraction and adsorb strongly in the Stern plane because of such phenomena as covalent bond
formation, hydrophobic bonding, hydrogen bonding, solvation effects, and so forth.
Because of their surface activity, the charge of such surface-active counterions adsorbed in
the Stern plane can exceed the surface charge. In this case
σS = –( σδ + σd )
(EQ 28)
where σδ is the charge due to adsorption in the Stern plane, δ. Flotation collectors generally
function as surface-active counterions.
Electrokinetic phenomena, which involve the interrelation between mechanical and
electrical effects at a moving interface, have found widespread use in colloid and surface
chemistry. The two electrokinetic effects that have been most widely used are electrophoresis and streaming potential measurements. Electrokinetic results are generally expressed in
terms of the ζ-potential, which is the potential at the slipping plane when liquid is forced to
move relative to the solid; only those ions in the diffuse layer outside of the slipping plane
are involved in the electrokinetic process. Thus, while knowledge of the ζ-potential at some
single condition may be of certain value, determination of the change in ζ-potentials as solution conditions are varied is extremely useful. From these changes, modes of adsorption of
various kinds of ions can be ascertained if one makes the useful assumption that the slipping
plane and the Stern plane coincide (Kruyt 1952). This approximation seems permissible
because the potential differences between the plane δ and the slipping plane are small compared with the total potential differences across the double layer. It should be pointed out
that the case in which there is no ambiguity is when ψδ = 0, because ζ must be zero.
T H E E L E C T R I C A L D O U B L E L AY E R O N OX I D E M I N E R A L S
Because oxides constitute the most important class of nonmetallic minerals, they will be
given considerable detail in this section. Given that oxide minerals form hydroxylated surfaces when in contact with water vapor, a hydroxylated surface should be expected when the
solid is in equilibrium with an aqueous solution. Adsorption-dissociation of H+ from the
surface hydroxyls can account for the surface charge on the oxide (Yopps and Fuerstenau
1964; Zettlemoyer 1968):
MOH ( surf ) ↔ MO (–surf ) + H (+aq )
(EQ 29)
MOH ( surf ) + H (+aq ) ↔ MOH 2+( surf )
(EQ 30)
Parks and de Bruyn (Parks and de Bruyn 1962; Parks 1967) have postulated a different
mechanism for the charging of oxide surfaces, involving partial dissolution of the oxide and
formation of hydroxyl complexes in solution, followed by adsorption of these complexes:
M 2 O 3 ( solid ) + 3H 2 O ↔ 2M ( OH ) 3 ( aq )
(EQ 31)
3 – m + ( 3 – m )OH –
M ( OH ) 3 ( aq ) ↔ M ( OH ) m
( aq )
( aq )
(EQ 32)
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
111
3 – m ↔ M ( OH ) 3 – m
M ( OH ) m
( aq )
m ( surf )
(EQ 33)
The formation of a surface charge by either of these mechanisms, or even by direct adsorption of H+ and OH–, would result in an equivalent change in the pH of the solution.
Titration of a suspension of an oxide in water at high ionic strength (with an indifferent
electrolyte such as sodium chloride, or NaCl) can yield the surface charge by assuming that
all added H+ or OH– change the solution pH or are adsorbed at the surface. Under these
conditions, given that the counterions are the indifferent ions,
σ S ≡ F ( Γ H + – Γ ( OH ) – )
(EQ 34)
Figure 6 presents the results of such titration of synthetic ferric oxide (hematite) with
hydrogen and hydroxyl ions in the presence of nitrate (KNO3) as supporting electrolyte
(Parks and de Bruyn 1962). This figure clearly shows that the surface charge on ferric oxide
reverses its sign at pH 8.6 and that it increases in absolute magnitude with increasing ionic
strength and increasing concentration of potential-determining ion. The intersection of the
curves yields the PZC because the adsorption density is zero at this point. Interestingly, the
adsorption isotherms given in Figure 6 are linear with pH at high ionic strengths and can be
represented by an equation of the following form:
( Γ H + – Γ ( OH ) – ) = – F ( pH – pH PZC )
(EQ 35)
where pHPZC is the pH of the solution at the PZC of ferric oxide. (This results as a consequence of the capacitance of the Stern layer being constant, but this will not be discussed
further here.)
For an oxide, the surface potential would ideally be given by
( aH+ )
RT
ψ o = -------- ln ---------------------- = 0.059 ( pH PZC – pH )
F ( a H – ) PZC
volts
(EQ 36)
The potential drop across the double layer at the alumina–water interface is shown
schematically in Figure 7. As the ionic strength is increased, the double layer is reduced in
thickness, and the potential at the Stern plane (δ) is reduced.
Table 4 presents a tabulation of typical PZC values of several oxides. This table shows
that the surfaces of oxides range from being acidic in nature to quite basic. Examples of how
pH strongly affects adsorption at the surface of oxides will be shown later.
At this point, consider the Gibbs equation for oxide minerals in aqueous media in the
presence of an electrolyte, such as NaC1. The change in tension of the mineral–water interface due to the adsorption of H+, OH–, Na+, and Cl– is given by
dγ = – Γ H + dμ H + – Γ OH – dμ OH – – Γ Na + dμ Na + – Γ Cl – dμ Cl –
(EQ 37)
At constant ionic strength, this equation simplifies to
dγ = – ( Γ H + – Γ ( OH ) – )dμ H + = 2.3RT ( Γ H + – Γ OH – ) ⋅ d ( pH )
(EQ 38)
Thus, it can be seen that the surface tension of the mineral–water interface can be controlled by changing the pH at constant ionic strength. Figure 7 also shows schematically the
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112
FLOTATION FUNDAMENTALS
–40
1M
Ferric Oxide
KNO3
21°C
10–1
10–2
Adsorption Density (ΓH+ – ΓOH–), μmol/g
–20
10–3
10–4
0
+20
+40
PZC
+60
5
6
7
8
9
10
11
pH
Source: Parks and de Bruyn 1962.
FIGURE 6 Adsorption density of potential-determining ions on ferric oxide as a function of pH
and ionic strength using KNO3 as the indifferent electrolyte
δ
+59
+118
10–1
10–3 M
Potential, mV
Potential, mV
+118
0
Distance
–59
+59
0
–59
Distance
10–1
10–2 M
–118
–118
pH 7
pH 11
0
10–3 M
10–2
Δγ
10–1
7
9
11
pH
FIGURE 7 Electrical double layer at the alumina–water interface, showing the potential
distance curves for 10–3 and 10–1 molar indifferent electrolyte at pH 7 and pH 11 (the PZC
occurs at pH 9). The change in the solid–liquid interfacial tension with pH and ionic strength is
also shown schematically.
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
TABLE 4
113
PZC of some oxides*
Material
SiO2, silica gel
SiO2, α-quartz
SnO2, cassiterite
ZrO2, zirconia
TiO2, rutile
Fe2O3, hematite (natural)
Fe2O3, hematite (synthetic)
FeOOH, goethite
Al2O3, corundum
AlOOH, boehmite
MgO, magnesia
PZC, pH
1–2
2–3
4.5
4
5.8–6.7
4.8–6.7
8.6
6.8
9.1
7.8–8.8
12
Source: Fuerstenau 1970.
*These are typical results. The sources of oxide—its trace impurities, method of pretreatment, etc.—cause variations in
observed values.
change in interfacial tension of alumina with pH for several salt concentrations. The maximum in the interfacial tension occurs at the PZC (i.e., at pH 9).
In view of this discussion on surface charge, the Gibbs adsorption equation for an oxide
can be rewritten as
dγ = – σ S dψ o – Γ Na + dμ Na + – Γ Cl – dμ Cl –
(EQ 39)
or at constant ionic strength as
dγ = – σ S dψ o
(EQ 40)
The greater effect of increased ionic strength on the interfacial tension lowering (Figure 7)
simply results from the greater magnitude of σs.
Effect of Temperature on the Electrical Double Layer
In order to understand temperature effects in the adsorption of ionic collectors, the effects
that temperature has on the double layer at interfaces must be considered.
Experimentally, the PZC of oxides has been found to decrease with increasing temperature (Parks 1960; Lai 1970; Tewari and McLean 1972). An example of this is shown in Figure 8, which presents electrophoretic mobility data for alumina and magnesia at three
temperatures. The overall reaction and its equilibrium constant, K, for the charging of an
oxide surface given by Equations 29 and 30 can be written as
MO (–surf ) + 2H (+aq ) ↔ MOH 2+( surf )
(EQ 41)
( a MOH + )
2
K = -------------------------------( a MO – ) ( a H + ) 2
(EQ 42)
From thermodynamics, it is known that
ΔG o = ΔH o – T ΔS o = – RT ln K
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(EQ 43)
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114
FLOTATION FUNDAMENTALS
or
( a MOH + )
H o ΔS o
2
- = Δ
– ln ------------------------------------------ – -------2
RT
R
( a MO – ) ( a H + )
(EQ 44)
Given that the number of positive and negative sites must be equal at the PZC, it can be seen
that
Δ H o ΔS o
ln ( a H + ) = ----------- – -------2RT 2R
(EQ 45)
– ΔH o ΔS o
( pH ) PZC = -------------- + ----------4.6RT 4.6R
(EQ 46)
Assuming that the enthalpy does not vary over the temperature range considered, the
enthalpy and entropy for the protonization of the surface sites on oxides can then be evaluated from the shift in the PZC with temperature. Table 5 gives the results of this calculation
for several oxides (Lai 1970). These calculations show that the enthalpy for protonization of
surface sites varies with the nature of the metal oxide, but the entropy is approximately constant.
Because the process is exothermic, the PZC of oxides decreases with increasing temperature.
+6
MgO
25°C
15°C
5°C
Electrophoretic Mobility, μm/sec per V/cm
+4
+2
0
–2
Al2O3
45°C
25°C
10°C
Calculated
–4
–6
4
6
8
10
12
14
pH
Source: Lai 1970.
FIGURE 8 Electrophoretic mobilities of alumina and magnesia at different temperatures as a
function of pH. The pH at which the mobility reverses sign under these conditions is the PZC.
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
115
TABLE 5 Enthalpy and entropy of hydrogen ion adsorption on oxides, expressed for the
adsorption/desorption of a single H+
Oxide
SiO2, cassiterite
TiO2, rutile
Al2O3, alumina
MgO, magnesia
ΔH° kJ/mol H+
–0.65
–6.05
–10.16
–13.59
(pH)PZC at 25°C
2.2
6.3
9.1
11.8
ΔS° entropy units/mol H+
7.9
8.7
7.8
8.7
Source: Lai 1970.
–6
5°
25°
Surface Charge, δs, μC/cm 2
–4
45°
65°
85°
–2
0
+2
85°
Silver Iodide
0.1 M Ionic Strength
5°
+4
+200
+100
0
–100
–200
–300
-400
Surface Potential, Ψ0, mV
Source: Lyklema 1966.
FIGURE 9 Temperature dependence of the surface charge on AgI particles, showing the
decrease in surface charge with increasing temperature. All curves are normalized to the PZC
at reference temperature.
In addition to shifting the PZC, temperature also affects the charge on mineral surfaces. In general, raising the temperature results in a decrease of σs. Figure 9 presents
Lyklema’s (1966) results for the charge on silver iodide at temperatures ranging from 5°C to
85°C. The curves are all normalized to the PZC at the temperature of reference. Lyklema
suggests that the decrease in surface charge with increasing temperature might be due to a
gradual desorption of counterions from the double layer, with the double layer having a
more diffuse character. Because the surface charge also is given by the expression:
ε dψ
σ S = – ------ ⎛ -------⎞
4π ⎝ dx ⎠ x = 0
(EQ 47)
where ε is the dielectric constant of the medium, the decrease in ε with increasing temperature would account for some of the decrease. In the case of the double layer on oxides, based
on Equation 47 one can calculate the charge at various temperatures and pH’s, and such calculations show that the charge decreases with increasing temperature.
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116
FLOTATION FUNDAMENTALS
F R E E E N E R G Y O F C O L L E C TO R A D S O R P T I O N
Information on the thermodynamic aspects of collector adsorption can be obtained from
experimentally measured adsorption isotherms. Analysis of adsorption isotherms to obtain
thermodynamic data for collector adsorption has been completed using two different
approaches. Fuerstenau, Harper, and Miller (1970) have successfully used the Stern–Grahame
model of the electrical double layer to explain the adsorption of alkylammonium ions at the
quartz–water interface and alkyl sulfonates at the alumina–water interface. These collector
ions are considered to be adsorbed as counterions in the double layer. Cases (1970), on the
other hand, has interpreted the results of his adsorption studies on biotite in terms of the
theoretical isotherms established by Frumkin, Fowler, Hill, and Halsey. Both approaches
will be briefly discussed in the next few paragraphs.
First, equilibrium in heterogeneous systems is attained when the chemical potential of
species i is equal in all phases. For collector species i, its chemical potential in bulk solution is
given by
μ i = μ io + RT ln a i
(EQ 48)
where μio is its standard chemical potential and ai is its activity in solution. The chemical
potential of the same species at the surface, μiS, is
μ iS = ( μ io ) S + RT ln a iS
(EQ 49)
where (μio)S is the standard chemical potential of this species at the surface and aiS is its
activity in the surface. At equilibrium, given that μi = μiS,
μ io – ( μ io ) S
a iS ⁄ a i = exp ------------------------RT
(EQ 50)
This relation can be transformed into the well-known Stern–Grahame equation by making
the following assumptions:
a i = C(the bulk concentration) and a iS = Γ δ ⁄ 2r
(EQ 51)
where Γδ is the adsorption density in the Stern plane, and r is the effective radius of the
adsorbed ion. The standard free energy of adsorption, ΔGadso, is defined as
o
ΔG ads
= ( μ io ) S – μ io
(EQ 52)
Substituting these three relations (from Equations 51 and 52) into Equation 50 yields the
Stern–Grahame equation (Grahame 1947; Fuerstenau 1970)
o ⁄ RT )
Γ δ = 2rC exp ( – ΔG ads
(EQ 53)
Another approach to developing the Stern–Grahame equation can be made by applying the Law of Mass Action to a binary mixture of similarly sized molecules showing ideal
behavior in both the liquid phase and in the adsorbed layer. This leads to
x 1S x 2 ⁄ x 2S x 1 = K
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(EQ 54)
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
117
where x1 and x2 refer to the mole fraction of solute and solvent, respectively, in the liquid
phase, and x1S and x2S are mole fractions in the adsorbed state. Given that x2S = 1 – x1S,
Equation 54 can be rewritten as
o
– ΔG ads
x
x
x kS
⎞
------------- = K ----1 = ----1 exp ⎛ ----------------⎝
⎠
S
RT
x
x
1 – x1
2
2
(EQ 55)
At small values of x1, Equation 55 becomes essentially identical to the usual form of the
Stern–Grahame equation (and to the Langmuir equation).
The Stern–Grahame approach is useful in that it allows one to take electrical effects
into account when dealing with ionic collectors. If ions adsorb only through electrostatic
interactions, then the standard free energy of adsorption is given by
o
o
ΔG ads
= ΔG elec
= vFψ δ
(EQ 56)
where v is the valence of the adsorbed ion including sign. On oxide minerals, alkali cations
appear to be surface-inactive, that is, only electrostatic interactions appear to be operative.
Nitrate anions are surface-inactive on cassiterite (SnO2), rutile (TiO2), Al2O3, and hematite
(Fe2O3). Although chloride ions are not surface-active on Al2O3, they appear to be specifically adsorbed on Fe2O3 (Fuerstenau 1970).
When an ion exhibits surface affinity, it is considered to be specifically adsorbed, and
the free energy of adsorption has additional terms:
o
o
ΔG ads
= vFψ δ + ΔG spec
(EQ 57)
o
where ΔG spec
represents specific interaction terms. Certain inorganic ions exhibit surface
activity: Ba2+ on quartz, Ba2+ and SO4= on alumina, Ca2+ and SO4= on rutile, and hydrolyzed metal cations on a wide variety of solids (Fuerstenau 1970). Equation 57 shows that if
o
ΔG spec
is finite, then the ion will be positively adsorbed even if ψδ is zero or has the same
o
sign as the adsorbing ion. An estimation of ΔG spec
can be made for conditions when ψδ is
zero (e.g., when the electrophoretic mobility of the particles is zero).
For the adsorption of organic collectors of interest to flotation scientists, various
o as follows:
attempts have been made to split ΔG ads
o
o
o
o
o
ΔG ads
= ΔG elec
+ ΔG chem
+ ΔG CH
+ ΔG solv
+…
2
(EQ 58)
o
o
where ΔG elec
is the electrostatic contribution to the total free energy, ΔG chem
represents
o
the free energy due to the formation of covalent bonds with the surface, ΔG CH
represents
2
the interaction due to association of hydrocarbon chain of adsorbed collector ions at the
o
interface (sometimes called hydrophobic bonding), and ΔG solv
is the contribution of solvation
effects on the polar head of the adsorbate (collector) and adsorbent (mineral) to adsorption.
o
It is often customary to lump the ΔG o terms other than ΔG elec
together and call it
o
o
o
o ).
o
ΔG spec
= ΔG chem
+ ΔG CH
+
Δ
G
(i.e., ΔG spec
solv
2
Specific adsorption can be either physical or chemical in nature. If the ions are adsorbed
only through such forces as electrostatic attraction and through hydrophobic bonding (van
der Waals interaction between hydrocarbon chains), the process should be termed physical
adsorption or physisorption. If the collector forms a covalent bond with metal ions in the surface of the mineral, then the process should be termed chemisorption. Much of the work that
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118
FLOTATION FUNDAMENTALS
has been carried out with collector adsorption at interfaces has been concerned with physical adsorption, such as in the quartz–alkylammonium (de Bruyn 1955; Gaudin and Fuerstenau 1955; Somasundaran, Healy, and Fuerstenau 1964) and the alumina–alkyl sulfonate
systems (Somasundaran and Fuerstenau 1966; Wakamatsu and Fuerstenau 1968). Adsorption isotherms and electrokinetic measurements strongly suggest that at low concentrations,
these collector ions are adsorbed as individual ions in the Stern plane but that at higher concentrations, they associate at the interface into two-dimensional aggregates, which have
been termed hemimicelles (Gaudin and Fuerstenau 1955).
Adsorption in the alumina–alkyl sulfonate system has been studied in considerable
detail. Figure 10 presents the effect of the concentration of dodecyl sulfonate on the electrophoretic mobility (zeta potential), adsorption density, and contact angle on alumina at pH 7.2
and at an ionic strength of 2 × 10–3 M controlled by NaCl. From this figure, it can be seen
that the adsorption isotherm can be divided into three distinct regions. At low concentrations, adsorption of sulfonate ions occurs by exchange with chloride ions in the double layer;
during the exchange, the zeta potential (and ψδ) remains constant. In this region, only electrostatic adsorption potential is active. In the second region, the adsorbed ions begin to associate, with adsorption increasing markedly because of the enhanced adsorption potential as
o
ΔG CH
becomes effective. The third region is reached when the zeta potential reverses. At
2
concentrations higher than this, the electrostatic interaction opposes the specific adsorption
effects, resulting in a decrease in the slope of the adsorption isotherm. Using data published
o for the adsorption
by Somasundaran and Fuerstenau (1972), Dick (1972) evaluated ΔG ads
of sodium dodecyl sulfonate on alumina at pH 6.9 and his results are plotted in Figure 11.
The onset of hemimicelle formation is accompanied by the standard free energy of adsorpo again has
tion becoming sharply more negative. When the zeta potential reverses, ΔG ads
much less dependence on the bulk concentration of sulfonate.
By means of the Stern–Grahame model of the double layer, the contribution of the
cohesive energy per mole of CH2 groups to the adsorption potential can be quantitatively
evaluated. If the standard free energy for removing 1 mol of CH2 groups from water is
through association φ, then the total contribution is nφ if n is the number of CH2 groups in the
chain. Thus, the contribution from hydrocarbon chain association to the adsorption process is
o
ΔG CH
= nφ
2
(EQ 59)
The adsorption density, Γδ, of collector ions in the Stern plane in the absence of chemisorption will be given by
o ⁄ RT ) = 2rC exp [ ( – vFψ – n φ ) ⁄ RT ]
Γ δ = 2rC exp ( – ΔG ads
δ
(EQ 60)
From the results of Wakamatsu and Fuerstenau (1968), φ has been evaluated to be
about –1.0 RT (about 0.6 kcal/mol of CH2 groups) in agreement with values obtained from
solubility data and micelle formation. Because of the role that CH2 groups have in controlling the adsorption free energy in these systems, collector chain length will significantly
affect flotation behavior (Wakamatsu and Fuerstenau 1973). Lin and Somasundaran (1971)
have considered the transfer of aqueous surfactants to various types of interfacial states and
found that transfer energies can range from –0.6 RT (for micellization) to –2.0 RT (for
evaporation).
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
Cosine θ
Adsorption Density
Electrophoresis
III
0.2
–40
II
–30
0.4
0.6
10–11
–20
Cosine θ
Adsorption Density, mol/cm 2
10–10
–10
0
I
+10
Zeta Potential, mV
10–9
119
0.8
Point
of Zeta
Reversal
+20
10–12
+30
1.0
Alumina
2 × 10–3 N Ionic Strength
pH 7.2
24±1°C
+40
+50
10–13
10–5
10–4
10–3
10–2
Equilibrium Concentration of
Sodium Dodecyl Sulfonate, mol/L
Source: Wakamatsu and Fuerstenau 1973.
FIGURE 10 Adsorption density of sodium dodecyl sulfonate, the electrophoretic mobility, and
the contact angle of alumina as a function of the equilibrium concentration of sodium dodecyl
sulfonate at pH 7.2 and ionic strength 2 × 10–3 M controlled with NaCl
–24
–22
45°C
25°C
ζ=0
ΔG°ads, kJ/mol
–20
–18
–16
Alumina
2 × 10–3 N Ionic Strength
pH 6.9
–14
–12
10–5
10–4
10–3
Equilibrium Concentration of Sodium Dodecyl Sulfontate, mol/L
Source: Dick 1972.
FIGURE 11 Variation of the standard free energy of adsorption of sodium dodecyl sulfonate
on alumina at 25°C and 45°C for the isotherms given in Figure 13
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120
FLOTATION FUNDAMENTALS
In many mineral–collector systems, chemisorption occurs. Chemisorption is often
arbitrarily referred to as an adsorption process in which the adsorbate attaches to the surface
of the adsorbent with a molar free energy of approximately 42 kJ or greater. In terms of the
o
concepts of Equations 57 and 58, chemisorption occurs when ΔG chem
has some finite value.
In flotation systems, chemisorption is of primary interest because selectivity may be
obtained if there is a specific collector–mineral adsorption reaction rendering a single mineral or group of minerals to be air avid. In nonmetallic flotation systems, several examples of
chemisorption may be cited. Hexanethiol chemisorbs on the surface of zincite (ZnO) and
willemite (Zn2SiO4), forming a strong zinc–mercaptan bond. Oleate chemisorbs on fluorite
(CaF2), forming strongly adsorbed films that are difficult to remove, and on hematite and
numerous other minerals. Hydroxamates strongly chemisorb on chrysocolla and hematite.
A different approach has been taken by Cases and his associates, who consider the
adsorption process as a condensation process on either a homogeneous or a nonhomogeneous surface similar to that of gas adsorption. The general relation for adsorption isotherms when the adsorption is localized without the dissociation of adsorbed molecules
onto a homogeneous surface is (Cases 1970; Predali and Cases 1974)
f -+A
kT ln x = – k ln W a – φ a + kT ln --------1–f
(EQ 61)
where x is the mole fraction of adsorbate in the bulk, k ln Wa is the sum of all entropic terms
except the configurational entropy for a molecule in the adsorbed state (f is the fractional
surface coverage), φa is the differential energy of desorption per molecule, k ln f/ (1 – f ) is
the configurational entropy of an adsorbed molecule, k is the Boltzmann constant, T is the
temperature, and A is a constant. On assuming φa to be independent of surface coverage,
Equation 61 can be rearranged to give the familiar Langmuir isotherm:
x
f = ------------------------------------G
1
x + ------ exp ⎛ ------a-⎞
⎝ kT⎠
A2
(EQ 62)
where A2 = exp (–A/kT), and Ga is the free energy of adsorption per molecule = (–φa – kTlnWa).
Langmuir isotherms have been widely used for estimating the free energy of adsorption
from adsorption isotherms. If φa is assumed to be a function of surface coverage, f, one can
obtain the well-known Frumkin–Fowler isotherm. To derive an equation for thermodynamic equilibrium, Predali and Cases (1974) assumed (1) φa = φao + fω, where φao is the
normal binding energy between one adsorbed molecule and the surface, and fω represents
the lateral interactions between the adsorbed species in the adsorbed layer; (2) in the
adsorbed layer, the number of coordinations and the lateral interactions are the same as in
the plane of the lattice of the micelle, which leads to the expression φo = φoo + ω/2, where φo
is the energy to dissolve a molecule from half-crystal position in the micelle and φoo is the
half of normal binding energies in micelle per molecule from half-crystal position; (3) the
entropic functions are the same in the adsorbed layer and in the micelle; and (4) k T ln xo =
–k T ln Wo – φo + A where k ln Wo is the sum of all entropic terms for a molecule in the
half-crystal position, and xo is the mole fraction of the adsorbate (collector) in solution at
saturation (i.e., in equilibrium with micelles of collector). Predali and Cases obtained the
following relation for thermodynamic equilibrium between an adsorbed layer and a dilute
solution of the same substance:
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
121
f
ω
kT ln ---------- = ( φ ao – φ oo ) – ---- ( 1 – 2f ) + Δμ
1–f
2
(EQ 63)
where Δμ = k T ln x/xo and is a measure of undersaturation of collector in the aqueous solution. A plot of f versus Δμ for two cases, namely, ω < 4 kT (e.g., C8 alkylammonium chlorides) and ω > 4 kT (e.g., C12 surfactants) is given in Figure 12a. From the figure, it can be
noted that when ω < 4 kT, the slope of the isotherms on homogeneous surfaces at f = 0.5 is
finite. It is possible to calculate the lateral binding energy between hydrocarbon chains from
the value of the slope at f = 0.5. On the other hand, when ω > 4 kT, the slope of the isotherms is infinite at f = 0.5. Predali and Cases (1974) consider the portion of the curve
between M and N in Figure 12a to be related to a change of state because of the condensation of the layer. Cases (1970) studied the adsorption of dodecylammonium chloride on
homogeneous surfaces and found the isotherm to exhibit a constant slope up to f = 0.015,
followed by an infinite slope up to f = 0.985. This corresponds to a case where ω > 4 kT.
Cases suitably modified Equation 63 for nonhomogeneous surfaces, and this modified
equation predicts a constant slope for adsorption isotherms (f versus C plot) up to a value of
f close to 1. Predali and Cases (1974) studied the adsorption of long chain (> C10) alkylammonium chlorides on biotite and found the adsorption isotherms to exhibit two welldefined branches. The first has a finite slope, and, according to their developed theory, this
is due to the condensation on a nonhomogeneous surface. The second branch has an infinite
slope, which is considered to be due to condensation on a homogeneous surface—that is,
onto the first adsorbed layer (see Figure 12b).
Though the theoretical isotherm of Frumkin, Fowler, Hill, and Halsey fits the experimental data, as claimed by Cases (1970), there is one serious assumption underlying this
approach, that is, the entropic effects involved in the adsorption from solution are similar
and equal to that involved in the adsorption from gaseous state. This assumption is highly
questionable because water structure at a solid–liquid interface plays a key role in adsorption, and any water molecules released on adsorption of the collector would significantly
increase the entropy of the system. Furthermore, the utility of the Cases approach is somewhat
A
B
Biotite–Alkylammonium
Chlorides
Δμ*
N
II
ω < 4 kT
f
0.5
I
ω > 4 kT
M
Fractional Surface Coverage, f
1.0
2.0
C18
C16
C14 C12
1.5
1.0
C10
0.5
0.0
Δμ
0.0
10–6
10–5
10–4
10–3
Equilibrium Concentration, mol/L
Source: Predali and Cases 1974.
FIGURE 12 (a) Schematic Frumkin–Fowler isotherm on a homogeneous surface, and (b)
adsorption isotherms of alkylammonium chlorides of different chain lengths on biotite at pH
5.5 and 25°C
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122
FLOTATION FUNDAMENTALS
limited in that the introduction of electrical double-layer concepts to adsorption is a more
convenient method of dealing with the interaction of charged collectors with the mineral.
Finally, plotting results in the form of fractional surface coverage versus concentration masks
the very important effects encountered at low surface coverages.
Heat of Adsorption of Collectors
Collectors of interest in flotation can be classified into two types: (1) surfactants that
chemisorb or chemically react at the surface (e.g., hydroxamic acids on hematite or chrysocolla, oleate on fluorite or calcite, xanthates on sulfide), and (2) long-chain ionic surfactants
that adsorb physically as counterions in the electrical double layer. The second class of collectors is characterized by strong hydrocarbon chain–hydrocarbon chain interaction in the
Stern layer (e.g., dodecyl sulfonate on alumina or hematite, dodecylammonium chloride on
quartz). The interaction of long-chain carboxylates with oxides might be considered as an
intermediate category.
Measurements of the heat of adsorption of collectors may be able to shed some light on
the following aspects of the collection process:
1. Nature of the interaction with the substrate (For example, by analogy with the heats
of chemisorption and physisorption of gases, it may be expected that type 1 reactions will be more exothermic than type 2.)
2. Identification of the mode of attachment of flotation collectors
Pioneering work on calorimetric determination of the heat of adsorption of collectors
on sulfides has been conducted at the Royal School of Mines, London. Mellgren (1966)
investigated the heat of xanthate adsorption on galena samples “as ground” or previously
treated with potassium carbonate, sulfate, or thiosulfate solutions. The heat of ethyl xanthate adsorption on untreated galena (–83 kJ/mol of Pb2+) was equivalent to that obtained
when xanthate was reacted with lead thiosulfate at neutral pH values. This was as expected
because thiosulfate was found to be the oxidation product on the galena used. Mellgren
measured rather erratic and large ΔH values at high pH values and concluded that secondary reactions were involved at these high pH values. Heat of adsorption measurements on
samples treated with carbonate, sulfate, and thiosulfate gave values equivalent to those
obtained when ethyl xanthate was reacted with lead carbonate, lead sulfate, or lead thiosulfate. These findings suggested that metathetical reactions between the xanthate and the oxidation products on the surface of galena were involved. The quantities of xanthate adsorbed
on the carbonate- and sulfate-treated galena surfaces were equivalent to the quantities of
thiosulfate released into the solution after the treatment. This led to the conclusion that an
ion exchange mechanism might be involved in the adsorption process.
Mellgren and Rao (1968) performed similar experiments with potassium diethyldithiocarbamate instead of xanthate and obtained similar results. Gochin (1972) carried out an
elaborate thermochemical study of the flotation of sphalerite and pyrite along similar lines.
In addition, he studied the heat of Cu2+ adsorption on sphalerite (ZnS) and iron sulfide
(FeS) with a view toward understanding the activation of sphalerite by Cu2+, which takes
place according to the following reaction:
o
ZnS + Cu 2+ = CuS + Zn 2+ ; ΔH 298
K, theor. = 63kJ ⁄ mol
Gochin’s measured values agreed quite well with the theoretical value stated previously.
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
123
Two systematic studies on the heat of adsorption of type 2 collectors on oxides have
been reported—one by Roy and Fuerstenau (1968) and the other by Mellgren et al. (1974).
The heat of adsorption of type 2 collectors should consist of two parts: (1) heat due to interactions between the charged polar head groups and the charged surface sites and (2) heat
due to the interaction between the hydrocarbon chains. Both types of interactions involve
the removal of water in the vicinity of the surface. Roy and Fuerstenau (1968) studied the
heat of immersion of alumina in aqueous solutions containing sodium dodecyl sulfonate at
various pH values. They found that as the sulfonate ions begin to adsorb (associate) in the
Stern layer, the heat of immersion begins to become more negative than the heat of immersion of Al2O3 in water, reaching a value of –1,100 ergs/cm2 at monolayer coverage. At
higher coverages, the heat of immersion remained at about –1,100 ergs/cm2. They did not
detect any measurable heat effect when the sulfonate ions occurred only in the diffuse part
of the double layer. They have presented a model in which adsorption occurs by a process
where a layer of water molecules exists between the solid surface and the adsorbed layer of
sulfonate ions. By this model, they estimated an integral heat of adsorption of –5.0 kJ/mol.
Mellgren et al. (1974), working with the hematite–dodecyl sulfate system, found that
the heat of adsorption (–5.7 kJ/mol) remained constant from an adsorption density that is
about one-sixth of the monolayer coverage right up to that of the monolayer coverage. They
measured a heat close to the heat of micellization of sodium dodecyl sulfate. (Roy and Fuerstenau [1968] are of the opinion that only at multilayer coverages should the heat of adsorption be close to that of the heat of micelle formation.) Mellgren and colleagues found
evidence for sites of different adsorption energy on powdered natural hematite, the proportions of which depend on the method of grinding (or heat treatment) of hematite.
An indirect method for determining the heat of adsorption is to measure the amount of
collector adsorbed (Γ) as a function of concentration (C) at a given solution pH at two or
more different temperatures. From a Clausius–Clapeyron type of equation:
o
– ΔH ads
ln C⎞
⎛ ∂-----------= ----------------⎝ ∂T ⎠
RT 2
(EQ 64)
This heat of adsorption is the same as the isosteric heat of adsorption, Q st. The heat of
adsorption is calculated from adsorption isotherms obtained at two different temperatures.
For collectors such as amines, sulfonates, and so forth, which do not chemically react with
the surface, Q st is the heat of “adsorption” in the real sense of the word. On the other hand,
collectors such as salicylaldehyde (Rinelli, Marabini, and Alesse 1976) and xanthates
(Mellgren et al. 1974) chemically react with the surface and form metal–collector complexes
both on the surface and in the bulk. In those cases, Q st is not the real heat of adsorption but
represents a heat intermediate between heat of reaction and heat of adsorption. From measurement of isotherms for the adsorption of sodium dodecyl sulfonate at two temperatures
(Figure 13), Somasundaran and Fuerstenau (1972) evaluated the enthalpy and entropy of
the adsorption process. Their results show that the adsorption decreases with increasing
temperature, indicating the exothermic nature of the process.
Ball and Fuerstenau (1971) used a different approach for evaluating the heat of adsorption. Taking the logarithm of the adsorption density given by the Stern–Grahame equation,
they obtained
o ⁄ RT )
ln Γ δ = ln 2r + ln C – ( ΔG ads
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(EQ 65)
flotation0.book Page 124 Tuesday, January 2, 2007 7:36 PM
124
FLOTATION FUNDAMENTALS
Adsorption Density, mol/cm 2
10–10
10–11
Alumina
pH 6.9
2 × 10 M Ionic Strength
(NaCl)
10–12
–3
25°C
45°C
10–13
10–6
10–5
10–4
10–3
10–2
Equilibrium Concentration of Sodium Dodecyl
Sulfonate, mol/L
Source: Somasundaran and Fuerstenau 1972.
FIGURE 13 Adsorption density at 25°C and 45°C of dodecyl sulfonate on alumina as a
function of its equilibrium bulk concentration at pH 6.9 and ionic strength 2 × l0–3 M
Differentiating this equation with respect to temperature yields
o ⁄ RT )
o
d ( ΔG ads
– ΔH ads
∂ ln C
d ln r ∂ ln Γ
---------------------------------- = ⎛ ------------⎞ + ----------- – ⎛ -------------δ-⎞ = ----------------⎝ ∂T ⎠ Γ δ dT ⎝ ∂T ⎠ C
dT
RT 2
(EQ 66)
If the radius of the adsorbing ion is assumed to be independent of temperature,
o
– ΔH ads
ln C⎞ ⎛ ∂ ln Γ δ⎞
⎛ ∂-----------– -------------= ----------------⎝ ∂T ⎠ ⎝ ∂T ⎠ C
RT 2
(EQ 67)
o
Under conditions of constant adsorption density, ΔH ads
becomes the isoteric heat of
adsorption and is given by
Q st
ln C⎞
⎛ ∂-----------= --------⎝ ∂T ⎠ Γ δ
RT 2
(EQ 68)
The values of isoteric heat of adsorption can be calculated from the slope of the plot of
o can be obtained as a function of temperature
(log C)Γ against temperature. Because ΔG ads
o
from the adsorption data using Equation 65, ΔS ads
can be computed easily as
o
o
ΔS ads = – d ( ΔG ads ) ⁄ dT. Ball and Fuerstenau (1971) used this approach to calculate various thermodynamic quantities for the process of adsorption of dodecylammonium acetate
on quartz. Their calculations indicate large entropic effects that have been attributed to the
phenomenon of hydrophobic bonding.
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SOME ASPECTS OF FLOTATION THERMODYNAMICS
125
Effect of Temperature on Collector Adsorption
Investigations of the effect of temperature on the adsorption of collectors on mineral surfaces can be classified into the following broad categories:
1. Systems in which the collector exhibits physisorption characteristics
2. Systems wherein the collector chemisorbs onto the mineral
3. Systems in which the collector seems to chemically react at the mineral surface with
ions derived from the mineral surface. This is actually a borderline category and, for
all practical purposes, could be classified under category 2.
For any adsorption process, the free energy change is given by
ΔG ads = ΔH ads – T ΔS ads
(EQ 69)
If the process occurs spontaneously, then ΔG ads is negative. This can be achieved by
the adsorption process being exothermic or by its being accompanied by a large entropy
increase. Not much is known about entropy changes in collector adsorption processes. On
the one hand, there will be an entropy decrease because of the more ordered structure of collector ions or molecules at the interface, but on the other hand, there may be a significant
increase in entropy because of the release of structured water molecules from the solid–
water interface and/or from the collector molecule/ion. Particularly for longer-chained collectors, this could result in ΔS ads having a large positive value.
Type 1 systems are characterized by an exothermic heat and, as a consequence, adsorption decreases with increase in temperature. A good example of this type is the alumina–
dodecyl sulfonate system investigated by Somasundaran and Fuerstenau (1972). Figure 13,
which demonstrates their results, shows such a decrease in the adsorption density. Although
no one has yet analyzed the results in such a manner, these particular kinds of isotherms are
also complicated by a shift in the PZC and a reduction in surface charge as the temperature
is increased. Another example is the work of Ball and Fuerstenau (1971) on the aqueous
dodecylammonium acetate–quartz system. They determined the effect of temperature on
the electrokinetic potential of quartz in the presence of dodecylammonium acetate (see Figure 14) and found it to decrease in absolute magnitude with increasing temperature, which
points indirectly to a decrease in adsorption with increasing temperature.
Of the chemisorbing systems investigated, the temperature dependence of the adsorption of soaps on mineral surfaces has been investigated in some detail. Falconer (1949),
referring to a 1938 French patent, wrote in 1949 about the value of using high temperature
for the conditioning of nonsulfide ores for flotation with soaps. Specifically, he mentioned
conditioning above 35°C, and preferably above 60°C. In 1950, Cook and Last conducted an
interesting investigation of the effect of conditioning temperature on the flotation of fluorite
with oleic acid. Figure 15 presents their results on the flotation of a fluorite/calcite/barite ore.
Below 40°C, the recovery of fluorite is very low, but raising the conditioning temperature
causes the recovery to rise sharply, to 90%–95% at temperatures of 70°C or greater. Their
explanation is that the oleic acid is physisorbed at room temperature and that chemisorption
does not take place appreciably until the temperature is raised to 45°C–60°C. Because of an
activation energy, raising the temperature causes the chemisorption reaction rate to proceed
reasonably rapidly during the conditioning period.
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flotation0.book Page 126 Tuesday, January 2, 2007 7:36 PM
126
FLOTATION FUNDAMENTALS
–40
pH 7
25°C
15°C
5°C
Zeta Potential, mV
–30
–20
–10
0
+10
1
2
3
4
5
6
7 8 9 10
Dodecylammonium Acetate Concentration, M × 104
Source: Ball and Fuerstenau 1971.
FIGURE 14 Zeta potential of quartz as a function of dodecylammonium acetate concentration
at different temperatures
100
Flotation Recovery or Grade,%
80
60
40
Fluorite
20
Grade
Recovery
0
0
20
40
60
80
100
Conditioning Temperature,°C
Source: Cook and Last 1950.
FIGURE 15 Effect of conditioning temperature on flotation recovery and concentrate grade for
the flotation of a fluorite ore, which had been conditioned for 5 minutes with oleic acid,
quebracho, and sodium carbonate
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Amount Adsorbed
SOME ASPECTS OF FLOTATION THERMODYNAMICS
127
Chemisorption
Physisorption
Temperature
FIGURE 16 Schematic representation of an adsorption isobar, showing the transition from
physisorption to chemisorption. Only in the transition region should the adsorption increase
with increasing temperature.
In conditioning at elevated temperatures, activating the adsorption process for changing it from physisorption to chemisorption is illustrated schematically in Figure 16. The
change from one type of process to the other could begin at about 40°C in these systems.
Kulkarni and Somasundaran (1975) found that the adsorption of oleate on hematite
increases as the temperature is raised from 25°C to 75°C at low ionic strengths, but the converse was observed at very high ionic strengths (due to salting-out effects at lower temperatures). Figure 17, after Somasundaran and Kulkarni (1977), shows that the flotation
recovery at low ionic strength increases as the temperature is increased, with maximum
recovery always occurring at about pH 8. This implies marked increase in collector adsorption with increasing temperature.
Raghavan and Fuerstenau (1975) have investigated the effect of concentration, pH, and
temperature on the adsorption of octylhydroxamate on hematite. Optimum adsorption
always occurred at about pH 8.4; they considered the adsorbing species to be the hydroxamic acid molecule. Using Equation 55 in the following form:
o
– ΔG ads ⎞
Cf - = -----------exp ⎛ ------------------------⎝ RT ⎠
55.55
1–f
(EQ 70)
where f is the fraction of the surface covered by the collector, and C is the collector conceno to be approximately –32 kJ/mol and nearly indepentration in solution, they found ΔG ads
dent of pH and surface coverage (for f < 0.8). For a chemically adsorbing species, the
o
observed value of ΔG ads
is rather small. Their results on the effect of temperature, which
are presented in Figure 18, show a marked increase in adsorption with increasing temperature. However, the isotherms for 41°C and 60°C given in Figure 18 exhibit adsorption densities that exceed monolayer (7 × 10–10 mol/cm2) coverage. This indicates possible
formation of ferric hydroxamate by surface reaction. Thus, one effect of the temperature
might simply be to increase the solubility of the mineral, thereby giving rise to a larger number of Fe ions for reaction at the surface. In an investigation of the flotation of iron oxide
with octylhydroxamate and oleate as collector, Fuerstenau, Harper, and Miller (1970) found
enhanced flotation with increasing temperature. Their conclusions were that the chemisorption process definitely involves increased mineral solubility as the temperature is raised. In
© 2007 by the Society for Mining, Metallurgy, and Exploration.
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flotation0.book Page 128 Tuesday, January 2, 2007 7:36 PM
128
FLOTATION FUNDAMENTALS
100
Hematite
3 × 10–3 M KOI
80
Flotation Recovery, %
100°C
60
94°C
75°C
40
60°C
20
25°C
0
3
4
5
6
7
8
9
10
Flotation pH
Source: Somasundaran and Kulkarni 1977.
FIGURE 17 Effect of temperature on the flotation of hematite with 3 × l0–5 M potassium oleate
as collector
20
Amount Adsorbed, mol/cm 2 × 10 10
16
12
60°C
41°C
8
4
20°C
Hematite–Hydroxamate
pH 5.5
2 × 10–3 M KNO3
0
0
2
4
6
8
10
12
Hydroxamate Equilibrium Concentration × 104, mol/L
Source: Raghavan and Fuerstenau 1975.
FIGURE 18 Effect of temperature on the adsorption density of potassium octylhydroxamate
on hematite
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flotation0.book Page 129 Tuesday, January 2, 2007 7:36 PM
SOME ASPECTS OF FLOTATION THERMODYNAMICS
129
an earlier study, while working with octylhydroxamate and various other chelating agents as
collectors for chrysocolla, Peterson et al. (1965) clearly showed that raising the temperature
can markedly increase flotation response. The optimum pH for chrysocolla flotation always
occurred at about pH 6.
In an interesting study of the flotation characteristics of pyrolusite, Fuerstenau and Rice
(1968) found that flotation decreased slightly with a sulfonate when the temperature was
raised from 23°C to 60°C. On the other hand, with oleate as collector for the same mineral,
flotation increased markedly upon raising the temperature from 23°C to 60°C. However,
the pH for optimum flotation was reduced from 10 at 23°C to about 8 at 60°C. On the
other hand, Yousef, Arafa, and Malati (1971) found the adsorption of oleate on beta-manganese dioxide to be an exothermic process both in the absence and the presence of Ba2+ ions.
Cooke, Iawasaki, and Choi (1960) conducted a detailed investigation of the effect of
temperature on the flotation of hematite. With a series of collectors—stearic, elaidic, oleic,
linoleic, and linolenic acids—they observed only a very slight increase in flotation and in
contact angles by increasing the temperature from 25°C to 70°C, except for stearic acid,
which exhibited a strong increase due to its increased solubility. They also found that contact angles of calcium-activated quartz at pH 11 decreased with increasing temperature,
indicating diminished calcium adsorption. On the other hand, flotation under the same
conditions increased with increasing temperature, but this was probably due to rather
unusual frothing characteristics at pH 11.
As has already been discussed at length, pH has a marked effect in flotation because of
its effect on collector ionization, surface charge, and mineral solubility. Thus, any temperature study of flotation systems should also take into account the temperature dependence of
the dissociation constant of water and pH (Harned and Owen 1958).
In conclusion, the flotation process, as a consequence of its being a multiphase heterogeneous system, is quite complicated to analyze thoroughly by simple thermodynamic
approaches. What this chapter attempted to do is bring out a few salient features of the
application of thermodynamic concepts to the understanding of flotation.
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FLOTATION FUNDAMENTALS
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The Nature of Hydrophobic
Attraction Forces
Jan Christer Eriksson and Roe-Hoan Yoon
INTRODUCTION
To gain an improved understanding of the molecular basis of hydrophobicity and hydrophobic attraction forces is of crucial importance for several scientific and technological
fields, such as surface and colloid science, biochemistry, and mineral processing. The formation of surfactant micelles and biomembranes, for instance, can be traced to the tendency of
hydrocarbon chains to associate when present in a water medium at a high enough (though
still very low) concentration. In addition, many enzymatic reactions take place while the
substrate is locked in a hydrophobic pocket. In the field of flotation, on the other hand,
long-range hydrophobic attraction forces contribute to establishing the mineral–particle/
air–bubble attachment necessary to achieve selective ore enrichment.
Early experiments indicating the existence of long-range hydrophobic attraction forces
were carried out by Blake and Kitchener (1979) on thin water films sandwiched between
hydrophobic surfaces. The instability observed was due to the attractive forces operating
across the film that arise because of the contact of the water with the hydrophobic surfaces.
Derjaguin and Churaev (1974) and Derjaguin, Churaev, and Muller (1987), referring to a
large body of experimental results, discussed the same effect in terms of “the structural component of the disjoining pressure.”
Most of the hydrophobic force measurements conducted during the last several decades
have been based on employing the surface force apparatus (SFA), developed by Tabor and
Winterton (1969) and Israelachvili and Tabor (1972) (Figure 1); the atomic force microscope (AFM) (Figure 2); or some similar direct-surface force measurement device such as
the measurement and analysis of surface interaction forces (MASIF) of Parker (1994) and
the interfacial gauge of Yaminsky, Ninham, and Stewart (1996).
This particular research area, which focuses on thin, aqueous films between hydrophobic surfaces, was inaugurated in 1982 by Israelachvili and Pashley by using mica surfaces and
successively adding the cationic surfactant hexadecyltrimethylammonium bromide (CTAB)
to the water medium (Israelachvili and Pashley 1982, 1984). They established that in addition to the normal DLVO (Derjaguin–Landau–Verwey–Overbeek) forces due to dispersion and electrostatic interactions, a long-range attractive interaction force operates as a
consequence of the hydrocarbon–water contact. Hence, they identified this surface force as
the “hydrophobic force,” a term originally coined by Blake and Kitchener (1979).
Several years later, Claesson and Christenson (1988) and Christenson and Claesson
(1988) experimented with mica surfaces that had been modified by Langmuir–Blodgett
(LB) deposition of dioctadecyldimethylammonium (DODA) bromide or, alternatively, a
double-chain cationic fluorocarbon surfactant. For these cases, the hydrophobic attraction
was considerably stronger and of longer range than observed by Israelachvili and Pashley
133
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FLOTATION FUNDAMENTALS
Fringes
Coarse
Separation
Distance
Control
Piezoelectric
Tube
Double-Beam
Cantilever
Curved
Mica
Surfaces
Clamp to
Adjust Spring
Stiffness
White
Light
NOTE: The interacting (mica) surfaces have cylindrical shape, and their separation is determined optically by means of an
interferometric technique. The lower surface is attached to a cantilever spring, the deflection of which serves to determine the
interaction force (Flinn 1997).
FIGURE 1 SFA based on the original design by Tabor and Winterton (1969) and Israelachvili
and Tabor (1972)
Laser Beam
Cantilever
Deflection
Flat Plate
Sample
Microsphere
Piezoelectric
Tube
NOTE: The top portion shows a drawing of a triangular cantilever with an attached glass sphere. The distance between the
sphere and flat plate (sample) is measured by monitoring the deflection of the cantilever spring using a laser beam. The deflection of the cantilever of a known spring constant gives the force at a given separation distance (Flinn 1997).
FIGURE 2
Schematic representation of AFM
(1982, 1984)—about 100 times stronger than the van der Waals attraction; decay length
about 15 nm—and measurable up to about 80 nm. Independently, measurements were also
conducted in Moscow by Rabinovich and Derjaguin (1988) using silanated silica filaments,
which yielded mostly equivalent results (Figure 3).
Subsequent control experiments employing a variety of prepared hydrophobic surfaces
have only partially confirmed the previous findings and perceptions. An abundance of
biased and bewildering results have been presented by many investigators, making it difficult
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
135
h, nm
0
10
20
30
40
50
60
70
80
–10
Rabinovich and
Derjaguin
(1988)
F / R , μN m–1
–102
–103
–104
NOTE: The points demark experimental values measured by Claesson and Christenson (1988). The solid line represents
Equation 56, obtained by inserting B = 0.6 mJ/m2 and b–1 = 15.8 nm from Equation 31. For comparison, the dotted line shows
the surface force function determined by Rabinovich and Derjaguin (1988) for silica filaments immersed in 0.1 mM KCl (potassium chloride) solution. The corresponding experimental points are scattered around the dotted line within a factor of about 3.
FIGURE 3 Plot of the attractive surface force, F/R, vs. the surface separation, h, for DODAcovered mica surfaces in contact with pure water
to summarize the current status of research. Nevertheless, Christenson and Claesson (2001)
presented a detailed account of the scientific state of the art from an experimental perspective. They classified the non-DLVO attractive forces observed between hydrophobic surfaces into the following classes:
1. A fairly short-range but strongly attractive force, much stronger than the van der
Waals force, between stable hydrophobic surfaces
2. An attraction of variable strength and range caused by the presence of small bubbles
sporadically adhering to hydrophobic surfaces
3. A very long-range attractive force with exponential decay operating between a variety
of hydrophobic surfaces, in particular those hydrophobized by means of Langmuir–
Blodgett deposition on mica or silica, or adsorption from solution of an ionic
amphiphile
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136
FLOTATION FUNDAMENTALS
Spalla (2000) presented a complementary paper, primarily dealing with the theoretical
understanding of the hydrophobic force. Furthermore, papers by Tsao, Evans, and Wennerström (1993) and Craig, Ninham, and Pashley (1998) incorporate illuminating discussions
about the possible origins of the hydrophobic force. More recently, Israelachvili and co-workers
have begun to critically re-examine the entire issue, and extensive reference lists are included
in their papers (Meyer, Lin, and Israelachvili 2005; Lin et al. 2005).
At this point, the chapter focus is on some typical results obtained for the main hydrophobic surfaces investigated so far, and in particular, for hydrophobic surfaces that are of
superior quality. The ultimate purpose is to scrutinize whether there is enough evidence
today in support of the existence of a “true” hydrophobic force (i.e., a surface force of long
range that is related to the structural response of a thin water film between hydrophobic surfaces). First, some recent findings as to the molecular organization of water interfaces and
hydrophobic interactions in general are discussed. After dwelling on the thermodynamics of
the surface force experiment, the concept of an ideal hydrophobic surface is introduced, thus
providing a frame of reference for judging experiments with real hydrophobic surfaces that
usually exhibit many deficiencies. One section is devoted to the possible formation of bridging air bubbles (and cavities) for hydrophobic surfaces with (equilibrium) contact angles
against water in excess of 90°.
M O L E C U L A R O R G A N I Z AT I O N O F WAT E R AT I N T E R FA C E S
Because of the advent of novel spectroscopic and computational methods in recent years,
considerable progress has been made in the probing and modeling of the molecular state of
bulk as well as of surface water. These novel methods include sum-frequency generation
(SFG) spectroscopy, X-ray absorption spectroscopy (XAS), and X-ray Raman scattering
(XRS) (Miranda and Shen 1999; Wernet et al. 2004). Hence, it has been established that
about one-fourth to one-fifth of the water molecules in the top monolayer of the water–air
interface have a non-hydrogen-bonded OH group as a unique feature.
The dangling –OH group spectral peak for a hydrocarbon–water interface remains
essentially the same as for the free water surface toward air, and can be rationalized on the
notion of oriented oxygen double layers similar to those present in the ice Ih lattice (Figure 4).
Evidently, a large part of the comparatively high surface energy of water is connected with
the excess of broken hydrogen bonds in the surface, compared to the situation in bulk water.
Rather surprisingly, however, Cavalleri (2004) has recently found that for bulk water
there is also a tendency to form linear aggregates of water molecules (based on two strong
hydrogen bonds rather than four of medium strength), resulting in rings and chains of water
molecules. On this point, one may note that according to Hill’s small-system thermodynamics, a distribution of linear aggregates may arise insofar as the free energy cost of introducing
one additional molecule in the central part of such an aggregate is small enough to become
compensated by the fluctuation entropy contribution that is inherent in the length distribution. In the absence of any long-range forces, a mechanism of this thermodynamic nature is
decisive for the formation of elongated, rod-shaped surfactant micelles. A similar scheme
can be applicable even for chains of water molecules and might be at the root of water effects
of long range. It remains to be assessed, however, as to what extent the novel findings and
ideas about the structural aspects of liquid water will necessitate revising the conventional
molecular picture based on four-coordinated water molecules in small clusters having external surfaces with less-well-bonded molecules (Ludwig 2001).
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137
712
THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
275
275
450
FIGURE 4 The ideal ice Ih structure, characterized by stacked oxygen double layers. The
numbers represent dimensions in picometers.
Thus, when considering hydrophobicity and hydrophobic surface forces, bear in mind
the ongoing endeavors to probe and to simulate interfacial and bulk water by means of
molecular dynamics, which may result in a substantially revised basis for a comprehensive
theoretical treatment of these phenomena (Cavalleri 2004). The classical simulations of Lee,
McCammon, and Rossky (1984), which are often referred to in this context, may have been
based on a simplistic water potential (of ST2 type). This has already been indicated by the
fact that the predictions of the melting point of ice using water models of this kind are, for
the most part, grossly in error. Moreover, to only explore density changes in a thin water film
might not be sufficient because minor, although thermodynamically significant, structural
rearrangements can occur without sensible volume changes.
Speaking in broad terms, an atomically smooth, chemically inert solid surface that is
subject to thermal motions of small amplitudes may be expected to
• Force the water molecules to pile up against the surface, as would actually be the case
for any liquid, and
• Force the hydrogen bond network to rearrange in its vicinity so as to limit the number of broken hydrogen bonds, thus forming a clathrate-resembling contact monolayer of water molecules.
On average, this would result in a roughly tangential alignment of the water dipoles. In
a cooperative fashion, such a contact monolayer, in turn, imposes bonding constraints on
the successive layers of water molecules beneath it so as to generate a surface-induced network. A crucial and much-debated question arises: How deep toward the bulk may such a
network prevail? Interestingly, the SFG spectra recorded by Miranda and Shen (1999) indicate that a well-ordered, ice-like water structure predominates at a solid hydrophobic surface,
whereas a more disordered water structure prevails for the water–air and water–hexane
interfaces.
The previously mentioned simulations by Lee, McCammon, and Rossky (1984) and
similar ones by other researchers (Forsman, Jönsson, and Woodward 1996) indicate that a
significant surface effect is noticeable only for short distances in liquid water, on the order of
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138
FLOTATION FUNDAMENTALS
1 nm. In contrast, many surface force measurements (detection limit ≈ 10 μNm–1) are
indicative of non-DLVO attractions extending as far as 50 nm and, occasionally, even
longer. Eventually, the reasons for these widely contrasting results will have to be determined
before a convincing theoretical description can be achieved for why and how the hydrophobic attraction arises.
H Y D R O P H O B I C E F F E C T S I N S O L U T I O N S A N D AT T H E WAT E R –
A I R I N T E R FA C E
The concept of hydrophobicity stems from the common observation that many substances,
such as oil and hydrocarbons, do not mix with water. Yet, for entropic reasons, there is
always a slight solubility, even for hydrocarbons, that reflects the local energetic conditions
of a single hydrophobic solute molecule surrounded by water molecules. By analyzing the
(mole–fraction-based) solubility in water of a number of straight-chain hydrocarbons,
Tanford (1980) was able to quantify the Gibbs (free) energy change associated with transferring
a –CH2– or a –CH3 group from bulk hydrocarbon to the dilute water state. These thermodynamic quantities are necessarily positive and of the following magnitude at room temperature (per methyl or methylene group):
0
Δμ CH
= 3.548k B T
3
0
Δμ CH
= 1.492k B T
2
(EQ 1)
where kB = R/NAvogadro and is the Boltzmann constant and T is the absolute temperature.
0
0
With sign reversed, Δμ CH
and Δμ CH
determine the driving force for aggregation of
3
2
hydrocarbon chains present in water, provided that the concentration of the chains exceeds
the normal water solubility of the bulk hydrocarbon, a condition that can easily be realized
by attaching a polar group to the hydrocarbon chain, thus forming an amphiphilic molecule.
Generally, the solubility of a hydrocarbon in water is nearly temperature-independent
at about room temperature. Referring to the thermodynamic Gibbs–Helmholtz equation
for the case under discussion,
0
d ln x hc
Δh hc
----------------- = ---------d( 1 ⁄ T )
kB
(EQ 2)
where xhc denotes the mole fraction of dissolved hydrocarbon. As xhc is approximately independent of temperature within the room-temperature range, this relation implies that the
enthalpy change for the process
hc (bulk hydrocarbon) → hc (water)
(EQ 3)
is small, which in turn means that the difference between the standard state chemical potentials is given by
0 = Δh 0 – T Δ s 0 ≈ – T Δ s 0
Δμ hc
hc
hc
hc
(EQ 4)
0 is largely determined by the (negative) entropy change, Δs 0 ,
In other words, Δμ hc
hc
accompanying the hydrocarbon dissolution in water.
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
139
According to Frank and Evans (1945), and Shinoda (1978), these thermodynamic
results can be understood in the following way. To insert a hydrocarbon molecule into water,
it takes a certain amount of energy to form a nano-cavity where the hydrocarbon molecule
can reside. At room temperature, however, this energy expense is almost exactly counterbalanced by the energy released upon structuring the surrounding water to create a more
0 is
ordered hydration shell, Δh structure . What is left, then, to obtain an estimate of Δμ hc
largely the concomitant entropy change:
0 ≈ Δs
Δs hc
structure ≈ Δh structure ⁄ T
(EQ 5)
that pertains to a process similar in nature to a freezing phase transition and is, hence, a negative quantity. The last approximate relation in Equation 5 follows because for such a phase
transition, there is almost no net change in Gibbs energy.
Conversely, T Δs structure may be identified as being the main driving force for hydrocarbon association in a water environment at ordinary temperatures. In other words, the overall
excess free energy due to the water structuring can be reduced by hydrocarbon association.
In this way, one can readily rationalize why, for example, ordinary surfactant micelles start to
form in solution (and hemimicelles on surfaces) at some fairly low “critical” surfactant concentration (i.e., the critical micelle concentration).
A few decades ago, the negative thermodynamic quantity Δs structure discussed above
was often attributed to “iceberg” formation that was considered to be the general cause of
hydrocarbon association in water, the prevailing belief being that release of structurally constrained water molecules furnishes the free energy required to accomplish the association.
The similarities with clathrate formation for inert gases in water were also frequently mentioned as supporting evidence. The full experimental picture is, however, somewhat more
complex; compare Figure 5, which shows that water exhibits “normal” liquid properties at
high temperatures where the solubility of a hydrocarbon increases with temperature but special, structurally related properties at lower temperatures. This behavior can, thus, be traced
back to the formation of hydrogen-bonded clusters and networks in water.
Basically, much the same explanation accounts for the relatively low surface entropy of
liquid water. On a molar basis, this entropy amounts to about 10 J/K·mole, whereas for normal,
non-hydrogen-bonded liquids, a value of about 25 J/K·mole (approximately one-fourth of
the entropy of vaporization) is typical (Eriksson 1966).
Moreover, at ordinary temperatures, the surface entropy of water increases with temperature, as appears from the following experimentally determined surface tension function for
pure water (Cini, Loglio, and Ficalbi 1972), recalling that the surface entropy is determined
by (the negative of ) the temperature derivative of the following function:
γ w, air = 75.653 – 0.1379t – 0.2717 × 10 –3 t 2
( mNm –1 )
(EQ 6)
where t denotes temperature in centigrade.
Because the surface entropy is an excess property relative to the entropy of water in the
bulk state, for an ordinary liquid the surface entropy is expected to be practically T-independent.
The fact that the surface entropy of water increases with temperature is thus indicative of a
gradual breakdown of an interfacial hydrogen-bond-dependent structure. This conclusion is
supported by the temperature dependence observed for the SFG spectrum of the pure
water–vapor interface that indicates a relative loss of ordered water as the temperature is
raised (Miranda and Shen 1999).
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140
FLOTATION FUNDAMENTALS
ºC
200 180 160 140
120
100
80
60
40
20
2
3
log X 2
Benzene
Toluene
4
Ethylbenzene
Extrapolated
Solubility
Curve
Experimental
Solubility
Curve
5
2.0
2.2
2.4
2.6
2.8
3.0
3.2
3.4
3.5
1/T × 103
NOTE: Typically, the solubility passes through a flat minimum at about room temperature. The effect of water structuring is
related with the vertical distances to the corresponding extrapolated lines.
Source: Data from Shinoda 1978.
FIGURE 5 Diagram displaying how the solubilities of alkylbenzenes in water vary with
temperature
For the case of a planar geometry, however, the enthalpy change associated with the
structure formation is insufficient to counterbalance the energy expenses because of rupturing hydrogen bonds toward air and the concomitant loss of dispersion interactions. Therefore, it can be assumed that the hydrophobic free energy is a curvature-dependent quantity,
being less for a strongly curved droplet of a hydrocarbon fluid than for a planar hydrocarbon–
water interface.
Although there is presently consensus among theoretical physicists and chemists about
the ability of liquid water to respond structurally to the formation of “internal” microscopic
as well as “external” macroscopic interfaces, obviously there is an unresolved issue about the
range that is affected. Calculations and molecular dynamics simulations using various water
interaction potentials made thus far show that the range having a molecular organization
different from that of bulk water may extend merely about 1 nm, whereas some surface force
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
141
measurements indicate attractive hydrophobic forces for water films as thick as 100 nm.
Hence, in an attempt to bridge the gap between theory and experiment, pertinent questions
to consider are as follows:
1. How realistic are the molecular model calculations/simulations for thin water films
carried out thus far? Do they comply with the notion of restructuring at a minor
free-energy expense? Would they be capable of predicting film tensions with the
necessary precision?
2. To what extent do the experimental hydrophobic surfaces employed actually mimic
an ideal hydrophobic surface?
The questions presented in no. 1 above are currently being pursued by means of novel
spectroscopic and computational methods. Certainly, the difficulties encountered along this
fundamentally oriented route are rather formidable. The latter question is more practically
oriented and leads to the almost arcane art of measuring contact angles in order to assess the
hydrophobicity/polarity of solid surfaces.
C O N TA C T A N G L E M E A S U R E M E N T S T O A S S E S S
HYDROPHOBICITY
It is common practice to quantify the hydrophobicity of a solid surface 1 by means of making contact-angle measurements, usually employing pure water 3 as the contacting liquid.
Assuming the solid–liquid interfacial tension (γ13), or more precisely, the reversible cleavage
work in water (compare Eriksson 1969) to be γ13 = 51 mJm–2 (i.e., the same as for an entirely
fluid hydrocarbon–water interface); taking the hydrophobic surface against air 2 to have a
corresponding interfacial tension value of γ12 = 21 mJm–2; and invoking γ32 = 72 mJm–2, followed by using the Young equation, one can readily predict the contact angle of the hydrophobic surface to be
–1
–1
θ = cos [ ( γ 12 – γ 13 ) ⁄ γ 32 ] = cos ( – 30 ⁄ 72 ) = 114.6°
(EQ 7)
where the term γ 12 – γ 13 (= 30 mJ m–2) is referred to as the “superficial tension” of water in
contact with the hydrophobic surface. This particular tension, introduced by Gibbs,
accounts for the tendency of water to contract on a solid surface. A prerequisite for the
above identification is that no, or just minute amounts of, water vapor adsorbs at the hydrocarbon–air interface. According to Equation 7, a contact angle of 90° is reached when the
superficial tension equals zero, whereas a completely water-indifferent, polar solid surface
resembling ice, with γ13 = 0 mJm–2 and γ32 = 72 mJm–2, would exhibit a contact angle of 0°.
Based on discussions in the previous paragraph, γ1 and γ13 have been plotted as functions of contact angle (cosθ), as shown in Figure 6. It is expected that γ13 = 36.0 mJm–2 at
θ = 90° and 18.0 mJm–2 at θ = 60°. Thus, one realizes that even surfaces exhibiting a contact
angle substantially less than 90° are presumably capable of exerting a structural hydrophobic
effect on water.
The interfacial tension, γ13, between solid 1 and water 3 can be considered to consist of
apolar and polar components. Thus,
LW + γ AB
γ 13 = γ 13
13
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(EQ 8)
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142
FLOTATION FUNDAMENTALS
60
γ13
50
γ, mJm–2
40
30
γ1
20
10
120°C
θ = 90°C
Hydrophobicity
Polarity
60°C
0
–0.5
–0.4
–0.3
–0.2
–0.1
0.0
0.1
0.2
0.3
0.4
0.5
cos (θ)
FIGURE 6 Diagram showing how, in theory, the cosine of the contact angle of water (cosθ )
and the solid–water (γγ13) and solid–air (γγ1) interfacial tensions depend on the polarity/
hydrophobicity of a solid surface
where LW refers to Lifshitz–van der Waals (i.e., apolar) interactions, and AB refers to acid–
base (polar) interactions. Good and Girifalco (1960), and Fowkes (1963) showed that
2
LW = ⎛
LW
LW ⎞
γ 13
⎝ γ1 – γ3 ⎠
(EQ 9)
where γ 1LW is the LW component of the surface free energy of solid, and γ 3LW is the LW
component of water. For the acid–base interactions, van Oss, Chaudhury, and Good (1987)
showed that
AB = 2 ⎛
+
+⎞ ⎛
–
–⎞
γ 13
⎝ γ1 – γ3 ⎠ ⎝ γ1 – γ3 ⎠
(EQ 10)
where γ 1+ and γ 3+ are the acidic components of the surface free energy of the solid and
water, respectively; and γ 1– and γ 3– are the basic components of the same.
Inserting Equations 9 and 10 into Equation 8, one obtains
γ 13 = ⎛⎝ γ 1LW – γ 3LW ⎞⎠ + 2 ⎛⎝ γ 1+ – γ 3+ ⎞⎠ ⎛⎝ γ 1– – γ 3– ⎞⎠
(EQ 11)
From the following fundamental relationships:
γ i = γ iLW + γ iAB
(EQ 12)
where the subscript i refers to phases of interest, and
γ iAB = 2 γ i+ γ i–
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(EQ 13)
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
143
one can rewrite Equation 11 to yield the following expression:
γ 13 = γ 1 + γ 3 – 2 γ 1LW γ 3LW – 2 γ 1+ γ 3– – 2 γ 1– γ 3+
(EQ 14)
The values of γ 1+ and γ 1– of many hydrophobic substances (e.g., Teflon, polypropylene,
and dodecane) are zero and therefore have high interfacial tensions. Surfactants are used to
render various solids hydrophobic. Because the acidic ( γ 3+ = 25.5 mJ/m2) and basic components ( γ 3– = 25.5 mJ/m2) of water are fixed under most experimental conditions, the major
role of a surfactant would be to decrease γ 1+ and γ 1– by blocking the hydrogen donor or
acceptor sites of the solid surface. Laskowski and Kitchener (1969) noted that hydrophobicity arises when water molecules are prevented from forming hydrogen bonds with the polar
sites on the surface of a solid.
Pazhianur and Yoon (2003) conducted surface force measurements between silica surfaces treated with octadecyltrichlorosilane (OTS) using an AFM and compared the results
with the changes in the surface free energies of the treated surfaces. The results are given in
Figure 7, in which K (of Equation 35) represents the magnitude of a hydrophobic force
measured. As shown, K increases with increasing advancing contact angle (θa) of the silica
surfaces. The sharp increase in K at θa close to 90° may represent a change in orientation of
the OTS molecule from flat to vertical orientation (Flinn, Guzonas, and Yoon 1994). As the
contact angle increases in excess of 90°, cavitation may occur and cause a strong, additional
non-DLVO attraction, providing an explanation for the change in slope of the K versus θa
plot. It is also possible that the second inflection point of the plot is caused by a flat (or flipflop) orientation of the additional OTS molecules adsorbing at high surface coverages. It is
important to note here that K increases with decreasing γ 1+ , γ 1– , and γ 1LW , as suggested by
Equation 14. These findings are consistent with the work of Ederth (1999), who showed
that both advancing and receding water contact angles on thiol-coated gold increase with
increasing fraction of CH3 groups relative to that of OH groups (Figure 8). Note also that
the surface free-energy data given in Figure 7 agree well with the approximate relationship
between surface and interfacial tensions (γ1 and γ13) and cos θ shown in Figure 6.
The interfacial tension (γ13) can be used to obtain the free energy of hydrophobic interaction (ΔG131) between two solid surfaces 1 in water 3 as follows:
ΔG 131 = – 2γ 13
(EQ 15)
Hence, from Equation 11, one obtains the following expression:
2
ΔG 131 = – 2 ⎛⎝ γ 1LW – γ 3LW ⎞⎠ – 4 ⎛⎝ γ 1+ γ 1– + γ 3+ γ 3– – γ 1+ γ 3– – γ 1– γ 3+ ⎞⎠
(EQ 16)
in which the first term is small for hydrophobic solids (e.g., hydrocarbons or alkanes adsorbing on a solid surface), because the values of γ 1LW and γ 3LW are close to each other. If γ 1+
and γ 1– are small (i.e., a solid is apolar), the free energy of the hydrophobic interaction arises
predominantly from the 4 γ 3+ γ 3– (= 102 mJm–2) term. On the other hand, if the value(s)
of γ 1+ and/or γ 1– is large, ΔG131 becomes positive (i.e., the hydrophobic interaction vanishes), which means that water molecules form hydrogen bonds with the surface hydroxyl
groups and render the surface hydrophobic. If a solid is basic (i.e., γ 1+ = 0, and γ 1LW =
40 mJm–2, the free energy of hydrophobic interaction vanishes at γ 1– > 28.3 mJm–2 (van Oss
1994).
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144
FLOTATION FUNDAMENTALS
45
40
10–16
γ1LW
35
10–17
25
K, Joules
Surface Free Energy, mN/m
30
γ1–
20
10–18
15
γ1AB
10
10–19
5
γ1+
0
10–20
50
60
70
80
90
100
110
120
θ2 (Water)
NOTE: K is a parameter of a power law (Equation 35). γ1LW and γ1AB are the nonpolar and polar components of the surface tension of the solid surface, γ1, whereas γ1+ and γ1– denote acidic and basic components, respectively, of γ1AB.
FIGURE 7 Changes in surface free energy of OTS-treated silica plate in air contact as a
function of the advancing water contact angle, and the changes in hydrophobic force constant,
K, as a function of the water contact angle
100
θ
90
80
70
55
60
65
70
75
80
85
Solution Fraction C16, %
Source: Ederth 1999.
FIGURE 8 Advancing ( ) and receding ( ) contact angles of water with surfaces prepared
from mixtures of C16 and C16OH-thiols of various mixing ratios
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
145
Thus, the hydrophobic interaction is seen to be related to the strong cohesive energy of
liquid water. The free energy gain due to the hydrogen bond interactions among water molecules is large as compared to the interaction with the hydrophobic surfaces, eventually causing the water film tension to diminish as the thickness shrinks, which in turn is manifested
as a hydrophobic attraction or a negative disjoining pressure. A question remains then as to
how the hydrophobic force decays with the distance separating two macroscopic particles.
Laskowski and Kitchener (1969) suggested that the multimolecular water layer on the surface of a hydrophobized silica is unstable, which is ascribed to a less favorable state of molecular association at a certain distance from the surface than in ordinary (bulk) water. These
investigators were the first to recognize the existence of a long-range, non-DLVO hydrophobic force and to suggest that the long-range character arises from the structural properties of water.
T H E R M O DY N A M I C A S P E C T S O F S U R FA C E F O R C E
MEASUREMENTS
Consider an idealized thin film/surface force experiment using two plane-parallel, atomically smooth, laterally homogeneous hydrophobic surfaces with a thin water solution film
between them (Figure 9). The thermodynamic variables of prime interest as determined in
the adjacent bulk water solution are the temperature T and the solute chemical potential
μS(or the concentration cS of solute).
To establish equality of the chemical potentials everywhere, a lateral tension γf develops
in the film which at large thicknesses of h will be equal to twice the interfacial tension of the
hydrophobic surface–water solution interface (i.e., about 100 mNm–1). In order to keep the
film at a certain thickness h, an extra pressure, positive (repulsive) or negative (attractive),
the so-called (Derjaguin) disjoining pressure has to be applied. This pressure, πD, has alternatively been called surface “interaction pressure.”
πD
Water Solution
γf
πD
NOTE: The excess (interaction) pressure operating perpendicular to the film is, using Derjaguin’s terminology, the disjoining
pressure, πD. In lateral directions, a film tension, γf, is acting. The surface force is defined as 2π times the difference in film tension between a thin film with interacting film faces and an infinitely thick film.
FIGURE 9
Schematic of a thin liquid film between two planar, solid surfaces
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146
FLOTATION FUNDAMENTALS
At constant temperature and pressure, the thin film system obeys the following thermodynamic fundamental equation to a very good approximation:
– d γ f = Γ sf,ex d μ s + π D dh
(EQ 17)
which constitutes an extension of the classical Gibbs surface tension equation to cover the
case of interacting solid surfaces separated by the film thickness h (compare to the appendix
at the end of this chapter).
The film thickness may most conveniently be defined as the separation between the
hydrophobic surfaces devoid of any loosely adsorbed solute. This means that the Gibbs
equimolecular dividing surfaces of the (hydrocarbon-covered) hydrophobic surfaces are
being employed to delimit the water film in the direction perpendicular to the film faces.
Furthermore, n sf,ex ≡ A Γ sf,ex (where A denotes the area of the film) is an excess quantity
in the following sense:
n sf,ex = n sf – n wf c s ⁄ c w
(EQ 18)
where n sf and n wf denote the actual mole numbers of surfactant and water in the film,
respectively, and cs and cw are the corresponding bulk phase concentrations. In other words,
n sf,ex represents the film content of the solute component in excess of what would be a corresponding slab of bulk solution containing the same amount of water as is present in the film.
From Equation 17 it can be seen that the disjoining pressure πD is generally related to
the film tension by the derivative
γ-f⎞
⎛ ∂-----= –πD
⎝ ∂h ⎠ T, p, μs
(EQ 19)
Hence, an attractive surface force, characterized by πD < 0, will result when the film tension γf decreases as the thickness diminishes, whereas πD > 0 means repulsion and a film tension that increases with decreasing h. In addition, Equation 17 includes an interesting
Maxwell relation,
f,ex
Γs ⎞
∂π D⎞
⎛ ∂-----------= ⎛ --------⎝ ∂h ⎠ T, p, μs
⎝ ∂μ s ⎠ T, p, h
(EQ 20)
stating that the amount of (surfactant) solute in the film will become less with decreasing
thickness if the disjoining pressure tends to increase when raising the chemical potential of
the solute.
The film tension γf is formally an Ω-potential per unit area,
γ f = Ω ⁄ A ≡ G f,ex ⁄ A = G f ⁄ A – Γ wf μ w – Γ sf μ s
(EQ 21)
Thus, when equilibrium prevails at constant T, p, and, μs, the film tension γf is necessarily at
a minimum.
To derive the film tension change, Δγf, arising because of diminishing the thickness
from h = ∞ where πD = 0 (i.e., beyond the range of the surface forces), Equation 19 can be
integrated as follows while assuming that the bulk phase state will remain the same:
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
147
h
Δγ f = γ f ( h ) – γ f ( ∞ ) = –
π D dh
∫
(EQ 22)
h=∞
In practice, it is not feasible to employ a plane-parallel experimental setup as that indicated
in Figure 9. Instead, one has to resort to sphere–sphere, sphere–plane, or crossed cylinder–
cylinder configurations. At this point, the amazingly powerful Derjaguin approximation
(1934) takes place that, in effect, constitutes a geometrical integration of Equation 19 (compare Figure 10), yielding an additional geometry-dependent factor such that the measured
surface force F becomes equal to F = 2πRΔγf or sphere–plane and crossed cylinder–cylinder
geometry and F = πRΔγf for sphere–sphere geometry. These simple relations effectuate the
transformation from a curved to a planar geometry. They are valid for various surface interaction forces insofar as the ranges of these forces are much less than the inverse curvature of
the surfaces involved. Although they are actually based on the very circumstance that (equilibrium) surface forces represent little else but film tension changes that arise when the interacting surfaces are brought close to contact, it has become common practice to report
surface forces “normalized” in the form F/R, corresponding to 2πΔγf and πΔγf for these two
standard geometries.
Next, consider a pure water film between hydrophobic surfaces. Because of the unfavorable water–hydrocarbon contact combined with the necessity of attaining a uniform value
of the water chemical potential everywhere, the local tangential pressure pT will have to
h1 = h + r 2/2R
R
Sphere
r
h
Ring of Radius r
Volume 2π drdh
dr
dh
FIGURE 10 Diagram of the geometrical disjoining pressure integration involving the Derjaguin
(1934) approximation
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FLOTATION FUNDAMENTALS
assume a variable negative value, corresponding to tension for each water layer located at x
outside the solid surface. Hence
p – p T ( x ) = [ p – p T ( 0 ) ] × exp ( – x ⁄ D )
(EQ 23)
Here, pT(0) denotes the tangential pressure next to the hydrophobic solid surface,
whereas p represents the atmospheric pressure of the surroundings. Equation 23 assumes an
exponential decay (decay length = D) of the thermodynamic effect caused by the presence of
the surface as a function of the distance x from the surface. Equation 23 might constitute a
reasonable first approximation, at least when the outer part of the surface zones begin to
overlap (compare Marčelja and Radic 1976).
Moreover, making the reasonable assumption that the surface force arising as the two
hydrophobic surfaces are brought closer is entirely due to the elimination of surface-affected
water molecules, integrate Equation 23 to yield (crossed cylinders):
F ⁄ ( 2 π R ) = Δγ f = – const. × exp ( – h ⁄ 2D )
(EQ 24)
Hence, for the large separation part of a surface force versus thickness isotherm on this
admittedly oversimplified basis, an exponential behavior of the (attractive) surface force
with respect to the film thickness h can be predicted. Physically, this means that upon
diminishing the film thickness, water molecules with slightly higher (Helmholtz) free
energy than in the bulk are being released from the thin water film and transferred to the
adjacent bulk water, thereby lowering the overall free energy of the film.
Essentially the same mechanism (i.e., replacement of surface-located water) is responsible for the adsorption of surfactant and polymers at the air–water interface in the dilute
Henry’s law region (Kumpulainen et al. 2005). In the following paragraphs, a more elaborate and more realistic treatment of the structural effect arising for the water in a thin film
between ideal hydrophobic surfaces is considered.
Diminishing film thickness from infinity to h = 0 will yield the adhesion force, as given
by the Derjaguin–Muller–Toporov relation (for nondeformed interacting cylindrical surfaces, compare Israelachvili 1991):
F adhesion ⁄ R = 2 πΔγ f ( h = 0 ) ≅ 4 πγ hc,w
(EQ 25)
The last of these relations assumes that there is no residual excess free energy associated
with the touching surfaces in direct adhesive contact. Accordingly, the maximal adhesion
force between fully hydrophobized solid surfaces in water is estimated to be 640 mN m–1.
Evidently, 2γhc,w represents all the free energy per square meter that is gained in the planar
case upon eliminating the hydrocarbon–water contact area. The additional 2πR-factor in
Equation 25 originates from the Derjaguin approximation.
However, when surface deformations are significant, the JKR ( Johnson, Kendal, Roberts) theory instead applies, which predicts a pull-off force equal to
F adhesion ⁄ R ≅ 3 πγ hc,w
(EQ 26)
that is, a maximal pull-off force of about 480 mN m–1 for two crossed-cylindrical, hydrocarbon-covered surfaces (compare Israelachvili 1991). Pull-off forces refer to the forces measured when two surfaces in contact with each other are separated.
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
149
The Ideal Hydrophobic Surface Versus Real Hydrophobic Surfaces
For experimental purposes, an almost-perfect hydrophobic surface and, hence, an “ideal”
hydrophobic surface in the theoretical, philosophical sense is often realized. The question
immediately arises about the most desirable properties of such a surface as to smoothness,
hydrophobicity, molecular constitution, motional state, fluidity, stability toward water solutions and other media, and so forth.
Regarding smoothness, the answer seems simple: a minimal roughness at the atomic
level is desired, on average approaching 0.1 nm in peak-to-valley distance. First of all, this
requires a smooth enough substrate; mica, silica, or glass is usually preferred. In addition,
however, the hydrocarbon chains used as modifying agents must be attached in such a manner that the smoothness is maintained. Chemical reaction and polymerization schemes for
the bonding of alkyl or silyl chains can be risky in this regard as they may involve strong
mechanical pressures/tensions to operate in the surface during the reactions.
Smoothness is likely to be a crucial factor for the occurrence of a long-ranged hydrophobic attraction force, and for this reason, it should be fully verified. The main experimental difficulty encountered in this context involves obtaining a sufficient degree of stability
for the hydrophobic layer attached, and at the same time preserving surface smoothness at
the atomic level.
Concerning hydrophobicity, the answer may seem almost self-evident at first: maximal
hydrophobicity calling for fluorocarbon rather than hydrocarbon chains. By making such an
extreme choice, however, one may easily end up with studying the effects of vapor/air cavities and bridging bubbles rather than the hydrophobic attraction per se. For this reason, one
might well prefer surfaces exhibiting a contact angle against water slightly less than 90° for
which capillarity phenomena of this extraneous nature, in principle at least, should not
occur.
As to the molecular constitution, one would preferably desire the hydrocarbon chains
employed to be sufficiently long (≥ C16 ), and have them fairly densely packed on the surface,
to make up a certain minimum thickness of the resulting surface-modifying layer. Yet, forcing them to adopt a crystalline state with tilted chains may cause complicating domain structures and grain boundaries to arise.
Concerning the motional state, one might prefer the hydrocarbon chains to be in a
“semifrozen” solid (rather than close-packed crystalline) state, having a packing density
between 0.22 and 0.25 nm2 per single chain. For a less crowded amphiphile monolayer in a
liquid-expanded state, one runs the risk of the thermal motions making the water–hydrocarbon phase boundary fuzzy.
Clearly, when ionic head groups are attached to the bonding sites of opposite charge on
a solid surface for which the thermal amplitudes are small, the chains may end up in some
“frozen” gel state, hampering the equilibration of the adsorbate, or at least making equilibration times exceedingly long. For mica and silica substrates, there seems to be a big difference
between C16 and C18 cationic surfactants in this regard, the latter approaching adsorption
equilibrium at an amazingly slow rate (Zhang et al. 2005).
In summary, an ideal, stable, hydrophobic surface would exhibit an advancing contact
angle against water of about 110°, and would in some solid state maintain a certain degree of
molecular mobility but nevertheless be smooth on the atomic scale. The fairly high mechanical tension acting in the water adjacent to the hydrocarbon film will further tend to
dampen the thermal amplitudes in the interface. If properly balanced in terms of its
mechanical properties to withstand mechanical loads, and if smooth enough from the
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FLOTATION FUNDAMENTALS
outset for such a surface contact angle, hysteresis might be limited and the risk minimized
for having bubbles attaching and bridging between approaching surfaces.
Real surfaces that exhibit pronounced hydrophobic properties to a variable degree are
listed as follows:
1. Cleaved mica with a deposited LB layer of a double-chain cationic surfactant (i.e.,
the hydrophobic surfaces employed by Claesson and Christenson [1988] and, more
recently, by Meyer, Lin, and Israelachvili [2005] and Lin et al. [2005]) fulfils the
quality criteria with respect to smoothness and motional state but shows lack of stability toward salt solutions and even toward passage through the water–air interface. One must realize that the chemical potential of the attached amphiphile is
raised by compressing the monolayer spread on the Langmuir trough to the desired
packing density of about 0.25 nm2 per single chain, which, by the way, matches
almost exactly the anionic site density of the mica surface (≈0.50 nm2). Consequently, upon screening (by adding salt), there is a tendency for the surfactant to
pass over to some less strained (bilayer) state. Further, the whole ion exchange process is prone to be facilitated kinetically by the presence of small ions.
2. A surface that is similar to point 1 but instead relies on adsorption from a cyclohexane solution to attach the double-chain cationic surfactant onto the mica. This
scheme was successfully applied by Tsao, Evans, and Wennerström (1993) and Tsao
et al. (1991), who obtained hydrophobized mica surfaces that were quite similar to
those made according to point 1. In particular, by means of AFM they convincingly
demonstrated that their surfaces were laterally homogeneous and free of defects
over great distances, on the order of microns. Interestingly, a substantially smaller
attraction force was recorded for C16 than for C18 chains, and significant temperature effects were noted, the attraction becoming weaker and less long-ranged at
higher temperature. However, stability toward salt solutions was less satisfactory
than for the corresponding LB monolayers prepared according to point 1.
3. Self-assembling alkylthiols dissolved in, for example, alcohol are known to bind
strongly to gold surfaces, a circumstance that was utilized by Ederth, Claesson, and
Liedberg (1998) to prepare hydrophobic surfaces on borosilicate glass substrates,
which started by forming (by electron beam evaporation) a 1-nm titanium layer, followed by a 10-nm gold layer. The gold surfaces made in this manner are slightly
rough with peak-to-through values of ~1.5 nm. The hydrocarbon chains become
tilted and are tightly packed in a crystalline state. These surfaces exhibit a hydrophobic attraction of medium range but have been plagued by sporadic bridging bubble
steps in the surface force curves. Stability is not a problem.
4. Glass, silica, or mica surfaces rendered hydrophobic by reaction with silanation
agents (e.g., (tridecafluoro-1,1,2,2-tetrahydrooctyl)dimethylchlorosilane; Parker
and Claesson 1994) and capable of reacting with surface OH groups. For mica surfaces, a water plasma pretreatment is necessary to introduce surface OH groups
(Parker, Cho, and Claesson 1989). As a rule, stability toward salt solutions is
obtained for this type of hydrophobic surface. Yet, to quote Christenson and Claesson (2001): “Experimental results with silylated surfaces have shown great variability depending on exact preparation conditions, and further underscored the critical
connection between details of surface chemistry, surface morphology and the
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
151
measured forces.” Hence, it seems hardly feasible to fully control a silanation reaction scheme so as to generate a sufficiently smooth, charge-free, and laterally
homogenous hydrophobic surface.
5. Wood and Sharma (1994) believed that one might succeed in making a more stable
and better characterized hydrophobic surface by first polymerizing OTS at the air–
water interface (pH = 2) on a Langmuir trough, followed by LB deposition and
chemical grafting to a mica surface. Certainly, the surfaces obtained proved stable
toward salt solutions but laterally inhomogeneous on the micrometer-length scale.
No long-range hydrophobic attraction was observed, presumably because of the lack
of molecular smoothness required.
6. Adsorption from water solution of a cationic surfactant such as CTAB or C18TAC
(octadecyl-trimethylammonium chloride) onto mica, glass, or silica. Though easily
prepared, these types of surfaces have a major drawback: they are uncharged (or
carry a minimal surface charge) only at one particular surfactant concentration,
often denoted as the charge neutralization concentration (CNC). Furthermore, for
the most desirable chain length, C18, adsorption equilibrium is only approached
very slowly at room temperature. Hence, in spite of the circumstance that the phenomenon of the long-range hydrophobic attraction is readily observed provided
that the right surfactant concentration is chosen, fundamental studies are hampered
by the variable state of the hydrophobic surfactant layer. Presumably, the best option
is to concentrate on those surface force curves that show the strongest hydrophobic
attraction rather than to try to sort out the exact electrostatic mechanisms that
operate at concentrations different from the CNC. The temperature dependence of
the hydrophobic force, for instance, could perhaps be investigated in this manner by
changing the temperature, followed by adjustment of the surfactant concentration
to recover a minimal surface charge.
This listing, though certainly not complete, may serve as a ranking list of hydrophobic
surfaces based on experiences from several laboratories during the past several decades. Preparing hydrophobic surfaces of sufficient quality is of crucial importance for making reliable
conclusions about the strength, range, and nature of the hydrophobic force. Although
hydrophobic surfaces prepared by adsorption of surfactants from solution are at the bottom
of the list, they are probably the most important ones in terms of applications, particularly
in flotation.
Hydrophobic Attraction Forces Under Ideal Conditions
For large separations between two interacting hydrophobic surfaces, an exponentially decaying, negative tangential pressure component, pT, gives rise to a likewise exponentially decaying surface force. However, for closer distances between the two hydrophobic surfaces, a
more elaborate model is necessary—in principle, similar to one by Cevs, Podgornik, and
Zeks (1982)—to account for repulsive hydration forces. In this type of model, changes of
state arising in the residual thin water film itself are also considered.
A quasi-thermodynamic/structural model for hydrophobic attraction forces was presented in 1989 by Eriksson, Ljunggren, and Claesson (1989). In contrast to most of the
other tentative explanations of hydrophobic force, this model has not been properly falsified
in the Popper sense. Rather, it has been refuted by the vague claim that long-ranged structural effects are virtually excluded for (normal) liquids. However, with its ability to form a
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FLOTATION FUNDAMENTALS
variety of hydrogen-bonding patterns, water is still far from being a well-understood liquid
at the molecular level. Hence, it is worthwhile to again scrutinize the basis of the waterstructure-based model and assess its predictive power by comparing it with the wider range
of experimental results that are available today. After all, this model might serve as a valid
point of departure for a deeper understanding of the hydrophobic force. Alternative
attempts to fully understand the hydrophobic attraction will also be addressed. The premises
are as follows:
1. The hydrophobic solid surfaces that interact in a liquid water medium are ideal:
they are geometrically smooth to within about 0.1 nm at 0 K, they exhibit small
thermal amplitudes at room temperature, and they are laterally close to being homogeneous in every respect.
2. Relatively speaking, there is a free-energy decrease associated with the molecular
reorganization of the first monolayer of water molecules next to the hydrophobic
surfaces because of minimizing the number of hydrogen bonds broken when contacting water with the hydrophobic surfaces.
3. The cooperatively enhanced tendency to avoid rupturing hydrogen bonds causes
the surface-induced structure to be propagated (with a certain decay rate) toward
the center of the thin film, resulting in a somewhat larger average number of hydrogen bonds per water molecule in the film than in the bulk. Hence, assuming only
short-range interaction forces, there is a free-energy increase arising throughout the
core of the thin film, owing to the imposed hydrogen-bond network formation.
The final (inhomogeneous) water state in the film reflects a balance between the favorable molecular reorganization occurring in the first (contact) layer of water molecules and
the induced, unfavorable restructuring of the remainder of the film. In the following discussion, these free-energy changes are accounted for by making use of a dimensionless order
parameter s(x), which is a measure of the local increase of the free-energy density in the thin
water film as compared with a corresponding slab of bulk water (x denotes the coordinate
perpendicular to the thin film). Alternatively, it may be assumed that s(x) mirrors the local
increase of the average number of hydrogen bonds per water molecule in the film, or the
associated decrease in local density.
Now imagine that the final equilibrium state of the thin water film sandwiched
between two hydrophobic surfaces is reached in a stepwise fashion. Starting from a thin film
cut out of the bulk state, the first step involves establishing the proper molecular interactions at hydrocarbon–water interfaces while retaining the average spherical-symmetric orientation of all the water molecules in the film.
The second step implies a change of the packing and of the average orientation of the
water molecules in each of the first molecular layers next to the hydrophobic surfaces to
yield a less dense and more ordered molecular state with an increased number of hydrogen
bonds and a preference for tangential alignment of the H–O–H bisectors of the water
molecules.
The film tension attained after these first two equilibration steps are denoted by γ 0f .
The third step comprises a reorganization of the hydrogen-bond network throughout the
film, whereby the parameter s becomes a function of the x coordinate and the final equilibrium value of γf is reached.
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
153
In an approach similar to the so-called square gradient approximation, frequently used
in the past to model liquid–vapor interfaces, the model features invoked previously can be
expressed in the following manner:
γ f = γ 0f – as 0 +
h⁄2
∫
[ c 2 s 2 + c 3 ( ds ⁄ dx ) 2 ] dx
(EQ 27)
–h ⁄ 2
The second term on the right-hand side of this expression (where a is a constant) accounts
for an assumed linear free-energy reduction because of changing the order parameter s0
(and, hence, the packing density) for the water layers in direct contact with the hydrophobic
surfaces, whereas the integral accounts for the free energy expense associated with structuring the core of the thin water film. Inclusion of the squared gradient term is essential because
it furnishes a mechanism of energetic coupling between successive layers of water molecules.
Hence, the constant c3 reflects the (average) tendency to cooperative structure generation.
Upon minimizing the film tension γf while taking into account the proper boundary
conditions, the following is readily derived
cosh ( bx )
a
s ( x ) = ⎛ ----------⎞ ---------------------------⎝ 2c 3 b⎠ sinh ( bh ⁄ 2 )
(EQ 28)
and
γ f ( h ) = γ 0f – ( a 2 ⁄ 4c 3 b ) coth ( bh ⁄ 2 ) = γ 0f – as 0 ⁄ 2
= γ 0f – ( B ⁄ 2 π ) coth ( bh ⁄ 2 )
(EQ 29)
The constant b stands for the quotient, whereas the interaction constant B introduced in
Equation 29 can also be written in the form B = π a 2 ⁄ 8c 2 c 3 .
For infinitely large film thicknesses, a lowering of the film tension caused by the
imposed structuring due to the hydrophobic surfaces is obtained:
γ f ( h = ∞ ) – γ 0f = – B ⁄ 2 π
(EQ 30)
implying that B/4π is the corresponding reduction of the interfacial tension between water
and a hydrophobic surface. It can be estimated to be ≈ 50 μNm–1, that is, to just about 0.1%
of the total interfacial tension value of approximately 50 mNm–1.
By taking the difference between Equation 29 and Equation 30 and making use of the
Derjaguin approximation, the hydrophobic attraction force as measured by means of an
SFA with cylindrically shaped hydrophobic surfaces having radii equal to R is given by the
following expression:
F ⁄ R = 2 πΔγ f = – B [ coth ( bh ⁄ 2 ) – 1 ]
(EQ 31)
For sufficiently large film thicknesses, the right-hand side of this equation becomes equal to
–2Bexp(–bh), that is, in the (weak overlap) regime, ln (–F/R) is predicted to be a linear
function of h, in line with Equation 23. Moreover, in this range, b–1 has the nature of a decay length.
The assumption invoked to derive Equation 23 was that the removal of surface-perturbed
water (due to the overlap of surface zones) is predominantly free energy-wise. On the other
hand, Equation 31 is more general because it also accounts for the free energy changes arising
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FLOTATION FUNDAMENTALS
in the residual thin film itself due to the elimination of the outer parts of the water surface
zones that are energetically perturbed by the hydrophobic surfaces. The major contribution
to this part of the lowering of the film tension is an additional reduction of the number of
broken hydrogen bonds for the contact layers of water molecules. As a consequence, the
hydrophobic surface force increases in magnitude at an accelerating rate when the film
thickness is diminished (compare Figure 17).
Equation 31 can be placed in an approximate double-exponential form, namely
F ⁄ R = – 2B ( e –bh + e –2bh )
(EQ 32)
where the second exponent accounts for the amplification of the surface force usually seen at
medium separations. However, unlike Equation 31, Equation 32 does not extrapolate properly
down to small h-values. Hence, the common practice of making use of a double-exponential
expression entailing four parameters instead of just two is understandable, namely
– F ⁄ R = C 1 e –h ⁄ λ1 + C 2 e –h ⁄ λ2
(EQ 33)
which, in many cases, has been shown to represent experimental surface force data satisfactorily in a wide separation range. The first exponent yields the largest contribution for small
separations, whereas the second one dominates at large separations. Typical values derived
are C1 = 0.2 Nm–1, λ1 = 2 nm, C2 = 1 mNm–1, and λ2 = 20 nm.
Another convenient, though somewhat less accurate, manner to represent experimental
surface force data is to invoke an expression of the same mathematical form as for the van der
Waals attraction, that is,
Δγ f = ( – K ⁄ 12 π )h –2
(EQ 34)
resulting in the one-parameter surface force expression:
– F ⁄ R = ( K ⁄ 6 )h –2
(EQ 35)
Obviously, the strength of the hydrophobic attraction can easily be judged by comparing the
value of the constant K with the corresponding value of the Hamaker constant, which is
usually on the order of joules. Typical K values range between 10–18 and 10–19 J.
The disjoining pressure expression, which can be derived from Equation 31 by differentiation with respect to h (compare Equation 19), is as follows:
2
π D = – ( bB ⁄ 2 π ) [ coth ( bh ⁄ 2 ) – 1 ]
(EQ 36)
Especially when displayed in this derivative mode, experimental hydrophobic interaction
curves may appear to belong to two distinct regimes: below and above 10–20 nm, respectively (Figure 11). The same holds true for the order parameter s in the middle of the thin
film (where x = 0), as derived theoretically. Below about 15 nm, s(x = 0) rapidly diminishes
with the film thickness h, whereas above about 15 nm, s(x = 0) decreases very slowly in an
almost linear fashion with h (Figure 12).
The hydrophobic surface forces arising within the large separation range are generally
rather weak, at most about 1 mNm–1, as compared with ≈500–600 mNm–1 for the maximum
adhesion force at h = 0, and can thus easily be concealed (e.g., by a surface force of electrostatic
origin).
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
155
106
DHDA
DODA
DEDA
–(2π)XπD, Nm –2
105
10
4
103
0
100
200
300
400
500
600
h, Å
FIGURE 11 Surface interaction pressures between mica surfaces modified with
dihexadecyldimethylammonium (DHDA), DODA, and dieicosyldimethylammonium (DEDA)
monolayers in water at 25°C as measured by Tsao et al. (1991). Note that the surface force
curves are indistinguishable for DODA (C18 chains) and DEDA (C20 chains).
0.10
0.08
s(x=0)
0.06
0.04
0.02
0.00
0
5
10
15
20
25
30
35
h, nm
FIGURE 12 Order parameter s (x ) in the mid-plane of a thin water film sandwiched between
hydrophobic surfaces (Eriksson, Ljunggren, and Claesson 1989)
Further, on the basis of Equation 36, it is estimated the excess free energy per water
molecule in the middle of the water film where the pressure tensor is likely to be approximately isotropic. Typically, in the medium separation range, h ≈ 10 nm, this excess free
energy amounts to about 4 × 10–4 kBT per molecule, as compared with the energy of a
hydrogen bond at room temperature, ≈ 7 kBT, again recognizing the very minute thermodynamic effects by means of a sensitive SFA or AFM setup.
According to the theory summarized, the strength of the hydrophobic attraction force
is determined by the interaction constant B = πa 2 ⁄ 8c 2 c 3 , which in turn is strongly
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FLOTATION FUNDAMENTALS
dependent on the constant a that is related to the free-energy change associated with the
reorganization of the contact layers of water molecules. Thus, B is anticipated to vary
strongly with the degree of hydrophobicity as well as the smoothness on the molecular scale
of the solid (or liquid) surface.
On the other hand, the decay length b –1 = c 3 2c 2 should depend primarily on the
properties of the water in the thin film, becoming large if structuring occurs readily (as for
bulk water at the freezing point) or when there is a strong tendency to avoid rapid changes
of the order parameter s(x), that is, when cooperativeness plays a significant role. However, if
the hydrophobic surfaces are also charged and (overlapping) electrostatic double layers are
present, one must anticipate competition between the (tangential) polarization of the water
molecules due to the hydrophobic surfaces and the (perpendicular) dipole alignment in the
electrostatic (mean) field. Such a coupling might result in a larger constant c2 and, hence, a
shorter decay length.
Salt effects for uncharged or nearly uncharged surfaces are expected to be rather minor,
provided that the hydrophobic surfaces themselves are stable in contact with salt solutions.
Generally, these features have been experimentally documented. In particular, note that the
strength of the hydrophobic attraction scales semi-quantitatively with the contact angle for
water on hydrophobic surfaces (Figure 6), and that Angarska et al. (2004) have shown that
Equation 31 applies even for thin foam films at high salt concentrations. Furthermore, generalization of the water-structure-based theory to the case of an unsymmetric aqueous thin
film between two different hydrophobic surfaces has been accomplished recently, an interesting case that has been studied experimentally by Yoon, Flinn, and Rabinovich (1997).
In conclusion, it is seen that the quasi-thermodynamic theory due to Eriksson, Ljunggren, and Claesson (1989), which focuses on the rather minor free-energy effects that are
associated with restructuring of water in contact with hydrocarbon surfaces, is capable of
systematizing several experimental findings concerning the attractive hydrophobic surface
force. Yet a fundamental problem related to this approach is that it does not provide an
understanding as to why the effect, in some instances, can be of such an amazingly long
range that it can be detected even for separations beyond 100 nm. One must bear in mind,
however, that the long-ranged hydrophobic surface forces are extremely weak and represent
very minute thermodynamic effects and that hydrogen-bonded networks and chains of
water molecules are known to be cooperatively stabilized, that is, larger clusters are inherently more stable than smaller ones. This may set the stage for extended water clusters of various shapes to occur, provided that the associated free-energy expenses are small enough to
be counterbalanced by corresponding size-fluctuation entropies, a thermodynamic scenario
that is familiar from the field of surfactant aggregation.
Nevertheless, there is definitely a need for more sophisticated models based on the concept of structure generation in water due to contact with a hydrophobic surface, as well as
for novel experimental methods by which one can investigate the detailed state of thin water films.
B U B B L E AT TA C H M E N T A N D C AV I T Y F O R M AT I O N AT
H Y D R O P H O B H I C S U R FA C E S
Upon observing more or less distinct steps in the surface force curves for hydrophobic surfaces
(plasma-treated mica silylated with (tridecafluoro-1,1,2,2-tetraoctyl)dimethyldichlorosilane)
submerged in water at surface separations in the range of 100 nm, Parker Claesson, and
Attard (1994) suggested that these steps might actually demark the onset of the hydrophobic
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
157
attraction, albeit an attraction that is totally different from the water-structure-based one
discussed thus far. Hence, they attributed the attraction to tiny adhering air bubbles that,
upon the approach of the hydrophobic surfaces, suddenly form bridging (quasi-cylindrical)
bubbles. Although the concept of bridging bubbles is, in many ways, appealing, there are
many difficulties associated with the claim that a capillary mechanism of this kind is the
chief reason for the long-range effects observed.
A primary difficulty is that very small air bubbles nucleated in a water phase are surprisingly short-lived. Because of the rather sizeable Laplace excess pressure, the air dissolves and
the bubbles diminish in size at an accelerating rate. Straightforward calculations show that
the lifetime of a small air bubble in water scales with the bubble radius squared. It is about
10 ms for a radius of 1 μm, and about 1 μsec for a 10-nm bubble, whereas a bubble that is
millimeter-sized may subsist for days or even months (Epstein and Plesset 1950; Ljunggren
and Eriksson 1997). Although air bubbles, as a rule, are generated upon contacting two
hydrophobic surfaces in an SFA device filled with water and then followed by pulling them
apart before starting the measurements, adhesion of small air bubbles is nonetheless a rare
event because of the limited life span of these bubbles.
Moreover, unlike the situation for larger air bubbles, which (because of the small excess
pressure) can be treated as if they were thermodynamically closed, very small open air bubbles are not expected to adhere to an ideal hydrophobic surface in a stable manner (Kralchevsky 1996; Eriksson and Ljunggren 1999; Ryan and Hemmingsen 1993). Most real
hydrophobic surfaces, however, contain defects that presumably play a crucial role in promoting bubble adhesion to occur to some minor extent.
What remains, following the “bubble-mechanism” line of reasoning, is the option that a
limited number of small bubbles, which have survived (because they started out as relatively
large bubbles), happen to adhere in an irreversible, defect-dependent manner to the hydrophobic surfaces, and that a few of these adhering bubbles rather quickly form approximately
cylindrical air bridges to the approaching hydrophobic surface. In this way, the excess
(Laplace) pressure will be efficiently cancelled, and a bridging bubble can consequently subsist for a long time. To assume irreversibility for the wetting behavior is essential here. Otherwise, the formation of one large, air-filled cavity, eventually giving rise to a huge surface
force, would have to be inferred.
Simple calculations show that in order to account for the surface force steps observed by
Ederth (1999), which amount to about 2 × 10–8 N in terms of force (rather than surface
force), one must invoke an adhering (spherical cap) bubble that fulfils the contact angle condition, having a radius of curvature of about 70 nm. The surface area covered by such a bubble would be ≈ 1.3 × 104 nm2. Supposing the three-phase contact line to be pinned in a fixed
position on some surface defects, the curvature is likely to decrease as some air dissolves, thus
lowering the Laplace pressure and extending the expected lifetime of the bubble. Gas supersaturation in the surrounding water phase may likewise prolong the lifetime of an adhering bubble.
As is easily confirmed, the formation of a bridging bubble out of an adhering bubble is
likely to be advantageous free-energy-wise. Nevertheless, it is significant as well as understandable, in view of the short lifetime and submicroscopic size of such a bubble, that it has
met with experimental difficulties to positively verify the presence of adhering air bubbles
on hydrophobic surfaces (Lin et al. 2005; Mao et al. 2004). To judge from the evidence
available today, the occurrence of adhering air bubbles depends strongly on the nature of the
hydrophobic surface. This might be anticipated, of course, because non-ideal features in the
hydrophobic surfaces themselves are likely to play a decisive role.
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FLOTATION FUNDAMENTALS
R
rc
FIGURE 13
h
Two approaching spheres with a bridging (quasi-cylindrical) air bubble
For a few years, it was a widely held belief that the long-ranged part of the hydrophobic
attraction is an artifact that can be attributed to a capillary bridging mechanism of the previously mentioned character. More recently, however, several investigators have emphasized
that the long-range hydrophobic attraction is actually present for several hydrophobic surfaces for which there are no signs of bubbles whatsoever. Hence, the supposition that small
adhering bubbles which collapse into bridging bubbles is the major cause of the “true” longrange hydrophobic attraction, must be rejected. Strong additional support for this conclusion has recently been obtained from careful degassing experiments (Meyer, Lin, and
Israelachvili 2005; Zhang et al. 2005).
Conversely, large-scale cavity formation is a capillary phenomenon that is fairly well
understood. It may occur when the surface tension of a solid surface in contact with air (or
vapor) is less than the surface tension of the solid surface in contact with a liquid so as to
make the cavity state the thermodynamically favored state. Expressed otherwise, it is a prerequisite that the (equilibrium) contact angle exceeds 90°. A somewhat simplified, though
still sufficiently accurate version of the classical treatment presented by Yushenko, Yaminsky,
and Shchukin (1983) is discussed in the following paragraphs.
Consider the case of two approaching hydrophobic spheres (radii = R) submerged in
water (Figure 13). Suppose a quasi-cylindrical (minimal-surface) cavity is formed that has
almost no excess air pressure associated with it. The free-energy cost of forming such a cavity
stems from forming the air–water interface, whereas the gain in free energy is due to replacing the hydrocarbon–water interface by hydrocarbon–air interface. For the free energy of
cavity formation, the following approximate expression is obtained:
ΔG
------- = r c ( h + r c2 ⁄ R ) γ air,w – r c2 ( γ hc,w – γ hc,air ) = γ air,w [ r c ( h + r c2 ⁄ R ) + r c2 cos θ eq ]
2π
(EQ 37)
where rc denotes the cylinder radius, and the r c2 ⁄ R term develops because the curvature of
the spherical hydrophobic surfaces is taken into account to the first order.
The expression in Equation 37 is apparently a third-order function of the cylinder
radius rc with a minimum for a rather large value (where the contact angle condition is, in
principle, fulfilled) and a maximum for some smaller rc value. The barrier associated with
the maximum is present, however, only when the distance of approach, h, is greater than
zero. (compare Figure 14).
Provided that the equilibrium contact angle is large enough, the free-energy value at the
minimum (representing the formation of a fully equilibrated air-filled cavity) will have a
negative value, implying that the cavity state constitutes the thermodynamically stable state
of the system. However, the intervening barrier will usually prevent the cavity from forming
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
159
1E – 15
1.0E – 09
θ = 105°
0
θ = 115°
∆G/2π, J
–1E – 15
0.0000
0.0002
0.0004
0.0006
0.0E + 00
θ = 105°
θ = 115°
–1.0E – 09
0.0
0.1
0.2
0.3
0.4
0.5
rc , mm
FIGURE 14 Interfacial free-energy function according to Yushenko, Yaminsky, and Shchukin
(1983), which accounts for the thermodynamically stable cavity formation between spherical,
hydrophobic surfaces submerged in water. The inset shows the barrier at short separations
that prevents cavity formation when the surfaces are brought close, though not touching, from
a great distance.
when the thin film state is approached by gradually diminishing the surface separation. In
other words, the thin film state of water between hydrophobic surfaces is actually a metastable state, as was noted earlier by Blake and Kitchener (1979).
The cavity state can easily be realized, however, by first bringing the hydrophobic surfaces into contact, thus reducing the barrier to zero, then slowly pulling them apart. The
attractive surface forces arising because of a large equilibrium cavity with an air pressure
somewhat less than atmospheric pressure (undersaturation) may approach as much as about
1 N when the cavity is formed in an SFA between crossed cylinders (R = 1 cm).
The disappearance of the cavity for some rather large h value is likewise readily understood on the basis of the previously mentioned simple equation. By increasing the surface
separation h, the entire ΔG function is rotated counterclockwise. Eventually, the second
(negative) term can no longer outweigh the first term, resulting in a minimum with a positive
ΔG value, thus providing the impetus for cavity annihilation.
In general, the experimental experience is in line with the thermodynamic description
of cavity formation as outlined here; however, lack of complete equilibration for the air dissolved as well as complications arising because of contact angle hysteresis must be considered.
Electrostatic Interaction Forces
To estimate repulsive electrostatic double-layer forces, the nonlinearized, mean-field Poisson–
Boltzmann (PB) scheme is usually applied in some numerical version. The theory behind it
assumes an evenly-smeared-out surface charge and a laterally homogeneous counterion distribution outside the charged (geometrical) surface. These assumptions seem more and
more questionable when the surface charge density is smaller. At the same time, however, the
magnitude of the repulsion becomes less, making the problems that may arise redundant in
any case.
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FLOTATION FUNDAMENTALS
The special case of one single (noninteracting) double layer is known as the Gouy–
Chapman approach, according to which can be obtained the electrostatic free energy per
charge, gel, from the following set of equations:
g el = e φ 0 + γ el a charge
(EQ 38)
where γel is the electrostatic contribution to the surface tension, and acharge is the area per
surface charge. This particular equation is, in fact, generally valid within the realm of the PB
theory ( Jönsson 1981; Evans and Ninham 1983). For the Gouy–Chapman case, the surface
potential is given by the following expression:
φ 0 = ( 2k B T ⁄ e ) [ S + ( S 2 + 1 ) 1 ⁄ 2 ]
(EQ 39)
and the electrostatic contribution to the surface tension is given by
γ el = ( – 2k B T ⁄ a charge ) [ ( S 2 + 1 ) 1 ⁄ 2 – 1 ] ⁄ S
(EQ 40)
where the negative sign indicates that this contribution is actually a surface pressure. Further, the dimensionless, reduced-charge parameter S introduced previously is defined by
S = ( σ 2 ⁄ 8RT ε 0 ε r c t ) 1 ⁄ 2
(EQ 41)
where σ is the surface charge density, e the proton charge, εr is the relative dielectric number
for water, and ct is the total (1:1) electrolyte concentration.
In this context, the Debye length is introduced by means of the following expression
(F denotes the Faraday constant):
κ –1 = F –1 ε 0 ε r RT ⁄ 2c t
(EQ 42)
Hence, the Debye length, which characterizes the extension of the diffuse part of the
double layer, according to the Debye–Hückel scheme, depends on the electrolyte concentration but not on the surface charge density.
From this background, briefly consider the case of an (1:1) ionic surfactant adsorbed at
an air–water interface in the dilute Henry’s-law regime. In this range, where γel constitutes
the chief contribution to the reduction of the overall surface tension, the surface tension
drops linearly with the surfactant concentration. Taking this circumstance into account and
making use of the Gibbs surface tension equation, one can readily show that the gas law–like relation
– γ el a charge = 2k B T
(EQ 43)
must hold exactly for thermodynamic reasons. Evidently, twice the –kBT/acharge can be attributed
to the pressure effects of both anions and cations in the surface. On the other hand, in Equation 40,
considering S to be >>1, about 1 kBT stems from the osmotic effect of the counterions,
whereas ≈ 1 kBT stems from the polarization of the water in the electrostatic mean field
(Ljunggren and Eriksson 1988). This discrepancy immediately conveys that the current theoretical description of electrostatic interactions has its weak points.
Furthermore, from Equation 40 it appears that the Gouy–Chapman theory incorporates an additional factor [(S2 + 1)1/2 – 1]/S, which deviates substantially from unity,
especially when S becomes less than about 10, which is the case for dilute ionic surfactant
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
161
layers. In other words, in spite of its many virtues, the Poisson–Boltzmann description has a
serious flaw, as it does not show a correct limiting behavior for charged equilibrium (Gibbs)
monolayers. This is in sharp contrast with, for example, the Flory–Huggins theory of polymer solutions.
The main reason behind these deficiencies is probably related to both the mean-field
concept as such and, in addition, to the particular model feature of a geometrical (i.e., volumeless), evenly charged surface devoid of entropy of mixing for the charged groups with water.
This latter shortcoming also shows up in the fairly low surface potentials and charge densities that are commonly derived from surface force data.
Nevertheless, in the PB scheme, the electrostatic disjoining pressure is primarily given
by the following expression (for simplicity, no added salt):
π Del = c ion RT ( e e φm ⁄ kB T + e –e φm ⁄ kB T – 2 )
(EQ 44)
where φm represents the mid-plane potential and cion for the bulk concentration (moles per
cubic meter) of positive and negative ions. In particular, in the much-employed weak-overlap
approximation that is applicable insofar as the surface potential is small (< 25 mV),
π Del ≈ ( ε r ε 0 κ 2 ⁄ 2 ) φ m2 ≈ 64k B Tc s Γ 02 e –κ h
(EQ 45)
Γ 0 = tanh ( eφ 0 ⁄ 4k B T )
(EQ 46)
where
By integrating Equation 45, the corresponding contribution to the surface force is easily found:
( F ⁄ 2 π R ) = 64k B Tc s Γ 02 e –κ h ⁄ κ
(EQ 47)
implying an exponential asymptotic behavior determined by the (salt-dependent) Debye
length. Thus, for large separations, there is a formal similarity (though the sign is opposite)
between the electrostatic repulsion and the water-structure-dependent hydrophobic attraction treated previously. Moreover, the electrostatic surface force contribution can readily
match the hydrophobic contribution in strength, provided that there is a significant surface
charge.
Still, a major difference from the theoretical standpoint is that the electrostatic doublelayer repulsion can be rationalized by referring to the well-established osmotic effect of
dissolved (point) ions in a structureless solvent medium, whereas up to this point, a firm
fundamental basis is lacking for an explanation of the hydrophobic attraction in terms of the
molecular properties of water.
In the 1970s, it was realized that correlation effects may become significant, especially
for interacting double layers encompassing divalent counterions (Guldbrand et al. 1984;
Kjellander and Marčelja 1984). An attraction of this origin is conceptually similar to the
dispersion interaction because it relies on instantaneous, laterally inhomogeneous counterion distributions outside the charged surfaces. On average, the fluctuations cause more
weight to be given to attractive rather than repulsive configurations, resulting in an overall
attraction. On this basis, the large deviations from PB theory for charged systems with divalent ions have been accounted for while at the same time verifying that the same PB theory
as a rule is reasonably accurate for systems containing monovalent ions.
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FLOTATION FUNDAMENTALS
For the correlation interaction between surfaces carrying mobile charges, the Ninham–
Parsegian expression holds true for the disjoining pressure. At large separations, it takes on
the following limiting form:
3
κ
π D = – kT
------------- e –2 κ h
κh
h » κ –1
(EQ 48)
The magnitude of the attractive interaction pressure predicted by means of this equation is, however, on the order of the van der Waals attraction. Hence, it is evident that charge
correlation attraction arising at the molecular (ion) length scale is not sufficiently strong to
rationalize the experimentally measured hydrophobic attraction. This was realized by Forsman, Jönsson, and Åkesson (1998), Miklavic et al. (1994), and Kekicheff and Spalla (1995),
who argued that more sizeable contributions of this nature can be obtained for charged surfaces that are heterogeneous, containing (mobile) charged surfactant aggregates of some
size, such as, for example, hemimicelles. In particular, Miklavic et al. (1994) were able to
demonstrate that at relatively high salt concentration, the correlation attraction should
decay with a decay length equal to κ–1/2(i.e., half of the Debye length) upon increasing the
surface separation. However, their calculations presuppose that the surface density of
adsorbed micellar aggregates is basically independent of the concentration of added salt, a
condition that for the most part is unlikely to be fulfilled.
Although these concepts about ionic correlation attractions may be of some relevance
for hydrophobic surfaces formed by adsorbing (e.g., a water-soluble cationic surfactant onto
mica, glass, or silica), it seems unacceptable to employ them to account for the archetypal
long-ranged hydrophobic attraction documented for the LB-monolayer-modified mica surfaces prepared by Claesson and Christenson (1988) or the corresponding mica surfaces prepared by adsorption from cyclohexane solution by Tsao et al. (1991), at least insofar as no
salt is added when making the surface force runs. The surface characterizations carried out
as well as the preparation protocols employed seem to leave little room for speculation in
this direction.
In a similar vein, it was suggested by Tsao, Evans, and Wennerström (1993), and Yoon
and Ravishankar (1996) that correlations among colloidal-grained, ordered dipole domains
across the thin film may give rise to sizable attractions, far stronger than the van der Waals
interaction. Although these matters, perhaps, are not yet resolved, it seems implausible that
all observations regarding the hydrophobic attraction can be accounted for on the basis of
correlation attractions alone.
Returning to the subject of hemimicelle formation, an aggregation of this kind is anticipated for hydrophobic surfaces prepared by surfactant adsorption from water solution,
starting at the critical hemimicelle concentration where the hydrophobic attraction is
already high. Herder (1990) demonstrated that upon adsorbing a cationic surfactant onto
Langmuir–Blodgett–DODA-modified mica, the hydrophobic attraction rapidly vanishes
and is replaced by electrostatic repulsion. Furthermore, Craig, Ninham, and Pashley (1998)
thoroughly investigated the interaction between silica surfaces in dilute CTAB and cetylpyridinium chloride (CPC) solutions in the presence of electrolytes, and firmly concluded
that a direct electrostatic mechanism for the hydrophobic attraction is hardly an option.
Parker and Claesson (1994) arrived at the same conclusion using the MASIF setup and
silanated glass spheres.
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
163
S U R FA C E F O R C E D ATA S U P P O R T I N G T H E WAT E R STRUCTURE MECHANISM
After about two decades of research efforts devoted to searching for the mechanism that
gives rise to the hydrophobic attraction, it is probably time to turn back to where the
surface–force-based development phase started—the measurements by Israelachvili and
Pashley (1982). These authors made use of mica surfaces submerged in dilute CTAB solution and interpreted the results rather straightforwardly in terms of the hydrophobic effect.
In the 1960s, the “mysteries” of water were still fashionable topics for study. Much
inspiration came from Pauling’s (1960) investigations of the hydrogen bond and from the
fascinating field of clathrate physical chemistry (Franks 1973). Later, the tendency to
restrict the discussion to structural aspects only—instead of employing the full statisticalmechanical machinery—and the polywater affair made this research direction fall in disrepute.
More recently, however, it has become abundantly clear that the molecular understanding of
liquid water is incomplete to an extent that renders it virtually excluded for making reliable
theoretical predictions about (transient) H-bond-generated structures.
In order to cope with the long-ranged hydrophobic attraction, a few different
approaches than the water-structure method have been rather thoroughly scrutinized. In
particular, electrostatics beyond the standard PB mean-field description have been reexamined, and more trivial, though quite cumbersome, capillary effects have been investigated in
detail. The latter phenomena include the formation of cavities and bridging bubbles
between (hysteretic) hydrophobic surfaces submerged in water. In the end, however, both of
these major alternative approaches have turned out to be untenable.
To further reveal the extent to which the water–structure-based theory of the hydrophobic force compares with experimental data, perhaps the most illuminating and most
well-documented experimental surface force study carried out so far will be examined—one
that clarifies the hydrophobic attraction (presented by Tsao et al. [1991]). These investigators made use of hydrophobized mica surfaces (mounted in an SFA), prepared by means of
adsorption from cyclohexane solutions of cationic surfactants of different chain lengths:
DHDA, DODA, and DEDA, with bromide or acetate as the counterions. The resulting
adsorption layers were characterized by AFM. Hence, it was verified that the double-chain
surfactant cations employed are electrostatically bonded, one to each anionic mica site,
resulting in a packing density close to 0.50 nm2 (0.25 nm2 per single chain), and that they
remain quantitatively bonded to their sites (though in a metastable state) even after contacting
the surface with water and raising the temperature to 50°C. Full stability toward salt solutions
was, however, not achieved using this surface preparation method.
Moreover, the aforementioned authors demonstrated that while the DEDA monolayer
preserves a frozen chain state at 50°C, this is not the case for the DHDA and DODA monolayers. In fact, DHDA appears to be present in a melted chain state already at 25°C, whereas
DODA melts in the range between 40° and 50°C.
Correlating changes were observed in the surface force curves, and it was concluded
that a well-organized, smooth hydrocarbon monolayer of frozen hydrocarbon chains causes
the strongest hydrophobic attraction. This finding was later corroborated by Rabinovich,
Guzonas, and Yoon (1993) by means of Fourier transform infrared (FTIR) measurements.
The chain length primarily matters in the sense that it determines where on the temperature
scale chain melting occurs. Moreover, for one and the same temperature, the decay length at
large separations was found to be nearly the same.
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FLOTATION FUNDAMENTALS
These observations fit well into the theoretical model description presented previously
that presumes structural water effects. Accordingly, the strength of the hydrophobic force is
determined primarily by the constant a of the (linear) response function –as0 , which participates in the expression for the constant B in Equation 31:
πa2
B = ---------------8c 2 c 3
(EQ 49)
In view of the cooperative nature of water structure generation, it seems very likely that a
stronger response (larger a) in terms of the lowering of free energy should result for the contact monolayer insofar as it is fairly unperturbed by the thermal motions of the hydrocarbon
surface. This dependence may suffice to account for the distinctly different surface force
behaviors recorded by Tsao et al. (1991) for hydrocarbon surfaces in a frozen (larger B) or
melted (smaller B) state.
Similarly, raising the temperature should tend to make a smaller, and, in addition, make
the structuring of the film core more costly in terms of free energy, yielding a higher c2, that
should likewise tend to diminish B. Also, the decay length b–1, which is given by c 3 ⁄ 2c 2 ,
depends inversely on c2 and should thus diminish with temperature, which is for the most
part observed. Concerning c3, just a modest dependence on temperature is anticipated.
On the other hand, the (unintentional) presence of a polarizing electrostatic field
would tend to increase c2 and hence shorten the decay length, whereas the presence of inert
gas molecules in the form of clathrate guest molecules might conceivably make it less costly
to restructure the water film core, implying a smaller c2 and a longer decay length.
More importantly, however, a virtually model-independent, surface-thermodynamic
analysis of surface force data as complete as those of Tsao et al. (1991) can be carried
through. Revisit the thermodynamic fundamental equation governing the present film case
(see appendix to this chapter). For a pure water film in contact with water,
d γ f = ( V f,ex ⁄ A )dp – ( S f,ex ⁄ A )dT – π D dh
(EQ 50)
where Vf,ex represents the volume in excess of the volume of a slab of bulk water that contains
the same number of water molecules as the film of a certain thickness. The excess entropy
Sf,ex is analogously defined. For an infinitely thick film, accordingly, obtain the following:
d γ f ( ∞ ) = ( V ∞f,ex ⁄ A )dp – ( S ∞f,ex ⁄ A )dT
(EQ 51)
Upon deducting Equation 51 from Equation 50, next obtain
d ( Δγ f ) = ( Δ V f,ex ⁄ A )dp – ( Δ S f,ex ⁄ A )dT – π D dh
(EQ 52)
This equation is the appropriate thermodynamic relation for dealing with surface force data
for thin films consisting of pure water. It obviously includes the partial derivative
( Δγ f )-⎞
Δ S f,ex
⎛ ∂--------------= – ------------⎝ ∂T ⎠ p, h
A
(EQ 53)
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
165
6
DEDA (s) and DODA (s) at 25°C
DEDA (s) at 50°C
DODA (s) at 40°C
DHDA (l) at 25°C
DHDA (l) at 40°C
5
–F/2π R, mJm –2
4
3
2
1
0
0
50
100
150
200
250
300
350
400
h, Å
FIGURE 15 Film tension changes for thin water films between hydrophobized mica surfaces at
different temperatures derived from the disjoining pressure data presented by Tsao et al. (1991)
Thus, if the surface force increases with temperature, which in the present case means that
the attractive hydrophobic surface force becomes less negative when the temperature is
raised, it necessarily results in ΔSf,ex as a negative quantity. In turn, this evaluation result
would imply that the thin water film is actually more ordered than an imaginary water film
of the same thickness encompassing two noninteracting surface zones of water next to the
hydrophobic surfaces. Provided that the hydrophobic surfaces themselves do not suffer any
changes in their intrinsic thermodynamic properties as the film thickness varies (which
should not be the case, especially if the surfaces are solid), there will be no contribution to
ΔSf,ex other than the one arising in the water film.
By integrating the disjoining pressure functions derived experimentally by Tsao et al.
(1991), the corresponding relative film tensions can be obtained (compare Figure 15).
These curves display the crucial feature discussed previously, that is, upon diminishing the
surface separation, the film tension reduction is larger at room temperature than it would be
if the room temperature were greater. Consequently, as soon as the surfaces interact, ΔSf,ex is
always a negative quantity. Also, Equation 31 with B-, b- values generated from the measurements of Claesson and Christenson (1988; who used similarly prepared hydrophobic surfaces), yields a reasonably good fit for the film tension curve obtained by Tsao and
colleagues for DODA at room temperature.
The relative excess film entropies per unit area, ΔSf,ex/A, can be quantified using Equation 53, resulting in the curves shown in Figure 16. Evidently, the excess entropy becomes
increasingly more negative when the water film gets thinner, especially for the frozen hydrocarbon surfaces, DODA and DEDA.
Regarding the numerical values, it is illuminating to compare with the entropy reduction one would get for a corresponding slab of bulk water that undergoes freezing to ice. By
invoking the entropy of fusion of ice (≈ 22 J mol–1 K–1) for a 5-nm water film, estimate
f
ΔS freezing
= – 6 mJm –2 K –1 , which is about 30 times more than the entropy reduction
quantified from the surface force measurements. In other words, the enhanced molecular
order in the thin film is equivalent to introducing about 3% of ice ordering.
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166
FLOTATION FUNDAMENTALS
250
DODA (s)
–∆S f,ex/A , μJm–2K–1
200
150
DEDA (s)
100
50
DHDA (l)
0
0
50
100
150
200
250
300
h, Å
FIGURE 16 Excess film entropies as functions of the separation between mica surfaces
hydrophobized with DHDA, DODA, and DEDA. The difference between the DODA and DEDA
curves is presumably due to experimental uncertainty.
Furthermore, comparing with the relative film tensions shown in Figure 15, observe
that for every thickness h, it holds true that
Δγ f << T ΔS f,ex ⁄ A
(EQ 54)
This inequality necessarily implies that the (negative) relative excess entropy of the film is
approximately counterbalanced by a corresponding enthalpy term, that is,
T ΔS f,ex ≈ ΔH f,ex
(EQ 55)
Such an enthalpy–entropy compensation is perhaps the most distinguishing feature accompanying structural changes in water. This was briefly discussed in relation with Equation 5.
For freezing of water at 0°C, ΔG = 0, which means that a relation similar to Equation 55 is
exactly fulfilled.
It is not necessary to invoke any mechanistic model whatsoever or to introduce any
speculative assumptions about the nature of the thin film system in order to make the preceding purely thermodynamic evaluation of the change in water film entropy. Conversely, it
would be meaningless to propose an interaction mechanism incapable of giving rise to a significant reduction of the water film entropy. This circumstance apparently restricts the number of alternative mechanisms and, hence, lends support to the idea that a water structuring
effect caused by the contact with hydrophobic surfaces is being addressed. It is envisaged
that the structure formation is most pronounced for smooth hydrophobic surfaces that are
free of charges and made up of hydrocarbon chains in a frozen state.
The most important assumptions on which these conclusions rest are the following:
1. The disjoining pressure data recorded by Tsao et al. (1991) are representative of a
water film state where the film itself is fully equilibrated, whereas the (smooth)
hydrophobic surfaces remain in the same metastable frozen state all the time.
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
167
2. No conceivable mechanism, other than the water restructuring mechanism, can
reproduce the strong temperature dependence and the characteristic enthalpy–
entropy compensation documented for the thin water film sandwiched between
hydrophobic surfaces.
The reversibility demonstrated with respect to temperature changes and the close quantitative agreement with surface force data from other laboratories using similarly prepared hydrophobic surfaces lend support to point 1. Point 2 is further discussed in the following paragraphs.
An additional advantage of the water restructuring model should be brought up explicitly. It is not necessary to assume that different physical mechanisms are operating at small
and large separations. By making use of Equation 31 and introducing two parameters (i.e.,
the surface force strength parameter, B, and the decay length, b–1), the seemingly different
mechanisms observed at small and large separations can be accounted for within one framework. Figure 17 shows a surface force function of the form given by Equation 31 and with
B-, b-parameter values obtained by fitting the data of Claesson and Christianson (1988),
namely,
h⁄2
--F- = – 0.600 coth ⎛ ----------⎞ – 1
⎝ 15.8⎠
R
( mJm –2 )
(EQ 56)
(where h is in nanometers) together with the surface force data reported by Lin et al. (2005).
It is evident that the function in Equation 56 produces an amazingly good fit, especially for
short separations.
Likewise, the disjoining pressure function (Equation 36) is given by:
2
h⁄2
π D = – 6.04 × 10 3 coth ⎛⎝ ----------⎞⎠ – 1
15.8
( Nm –2 )
(EQ 57)
0.01
F/R, mNm–1
0.1
1
10
100
0
10
20
30
40
50
h, nm
FIGURE 17 Surface force data points recorded by Lin et al. (2005) at 25°C for DODA-modified
mica surfaces in water, compared with the surface force function (Equation 31), where the
fitting parameters of B = 0.6 mJm–2 and b –1 = 15.8 nm
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168
FLOTATION FUNDAMENTALS
1,000,000
DODA
DEDA
(2π) x πD, Nm –2
100,000
10,000
1,000
0
10
20
30
40
50
60
h, nm
FIGURE 18 Comparison between the disjoining pressure data reported by Tsao et al. (1991)
and the disjoining pressure expression given by Equation 36 and inserting the parameter
values of B = 0.6 mJ/m2 and b –1 = 15.8 nm
which according to Figure 18 fairly well represents the data recorded by Tsao et al. (1991)
for DODA and DEDA at 25°C.
For h values that are small enough (less than about 2 nm), the B[coth(bh/2)–1] function can be approximated by 2B/bh; that is, the surface force is predicted to vary inversely
with h. The good fit in the short separation range displayed in Figure 17 is noteworthy. Still,
it is hardly conceivable that the model concepts introduced to derive Equation 31 would
apply below ≈2 nm; in other words, the agreement may appear reassuring but is presumably
little else than fortuitous.
For completeness, add an expression for ΔSf,ex, which can be derived by combining
Equations 31 and 53:
∂ ln B
Bbh ⎛ ∂ ln b ⁄ dT ⎞
ΔS f,ex = – ------------ Δγ f – --------- ⎜------------------------∂T
4 π ⎝ sinh 2bh ⁄ 2 ⎟⎠
(EQ 58)
where the (negative) terms on the right-hand side, primarily, are expected to contribute
about equally. Thus, approximately ΔSf,ex and Δγf are expected to be proportional. This particular
feature is clearly observed in the thermodynamic surface force data (compare Figures 15 and 16).
T H E S U R FA C E C L U S T E R M O D E L
Having recognized the significance of the strong temperature dependence of the hydrophobic force demonstrated by Tsao et al. (1991), and the ensuing conclusion concerning the
lowering of the water film entropy, questions are raised about alternative mechanisms to
account for the hydrophobic forces that are distinctly different from the one based on the
surface-induced water restructuring that have been favored thus far.
There are no strong indications that the hydrophobic surfaces primarily referred to here
(i.e., those used by Tsao et al. [1991]) would be laterally inhomogeneous. On the contrary,
AFM studies showed that they were homogeneous over a micrometer scale. Furthermore,
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
169
there is little reason to suspect that surfactant micelles or similar aggregates might be loosely
adsorbed onto the surfaces insofar as they are exposed to nothing else but pure water that is
continuously checked for contaminants. Consequently, it is not anticipated that the hydrophobic surfaces will grow inhomogeneous under the experimental conditions, thus setting
the stage for various forms of attractive ionic or dipolar correlation interactions (Tolstoi et al. 1966).
However, development of a well-bonded clathrate-like water monolayer next to a
hydrophobic surface might possibly involve the formation of small planar patches or islands.
These areas would comprise strongly correlated water dipoles pointing approximately in the
same tangential direction in the interface. In this way, large (cluster) dipole moments may
arise at each of the interfaces, which could interact by an electrostatic correlation mechanism and give rise to a strong attraction pressure. This is the key idea (in a slightly modified
version) of the hydrophobic force theory that was advanced by Pazhianur and Yoon (2003).
Making a Hamaker type of approach and assuming independent surface dipoles, one
can theoretically estimate the corresponding surface force by means of the following expression:
( ρ m2 ⁄ εr ε0 )2
--F- = – -----------------------------R
32k B Th 4
(EQ 59)
where ρ is the surface density of dipoles and m is the dipole moment. Equation 59 was
derived on the basis of the concept presented in Section 5.3 of Evans and Wenerström
(1999). For a surface separation of 10 nm, an average patch size 20 × 20 nm, and an effective
cluster dipole moment equal to 6 × 103 D, this expression yields an attractive surface force
amounting to 1.6 mJm–2, which is the correct order compared with experiments. However,
even if a good fit to experimental data can be obtained in the middle separation range, the h4
factor in the denominator makes it virtually impossible to simultaneously generate acceptable fits for short and large separations, respectively.
A distance dependence of h–4 type is likewise obtained using the more general Lifshitz
approach, regardless of the lateral dipole interaction mode assumed within the surfaces.
Hence, it does not appear straightforward to formulate a fully satisfactory theory of the
hydrophobic force on the basis presently being discussed. Nevertheless, it is an appealing
feature that the strong temperature dependence observed for the hydrophobic force might
be linked to the temperature-sensitive dipole moments of the correlated planar cluster
domains. Although it has been claimed that huge dipole moments can arise for hydrophobic
colloidal particles in water (Tolstoi et al. 1966), the implicit assumption of cluster dipole
moments on the order 103 – 104 D would need additional justification.
Incidentally, the K/h2 function that accounts for the hydrophobic surface force, though
evidently more appropriate than const.·h–4, does not yield an entirely satisfactory description
of experimental data. Primarily, it was invoked to facilitate making comparisons between the
strength of the hydrophobic force with the van der Waals attraction as well as devising theoretical descriptions of experimental film systems where both dispersion as well as electrostatic interaction forces must be considered.
EFFECT OF SOLUTES
Brief comments will be made on the effect of adding a solute to the water phase that is in
equilibrium with a thin film. The main thermodynamic relation governing this case is the
counterpart of the Gibbs surface tension equation, which at constant T and p reads:
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170
FLOTATION FUNDAMENTALS
d Δγ f = – ΔΓ sf,ex d μ s – π D dh
(EQ 60)
where μs denotes the chemical potential of the solute. Consequently, having access to equilibrium data on the surface force for a certain film thickness as a function of the solute concentration, the change in solute excess, ΔΓ sf,ex , for a film of that thickness as compared with
an infinitely thick film can be obtained.
Very recently, such SFA data were published by Meyer, Lin, and Israelachvili (2005),
aiming to clarify the effect of degassing (pure) water contacted with DODA-modified mica
surfaces. Within most of the separation range below 50 nm, a reduction due to degassing of
the magnitude of the hydrophobic force is clearly seen. Hence, referring to dissolution of the
main air components N2 and O2, the derivative is
Δγ-f⎞
⎛ ∂---------⎝ ∂μ s ⎠
T, p , h
= – ΔΓ sf,ex
(EQ 61)
has a negative sign. This obviously implies that ΔΓ sf,ex > 0; that is, the air gases are enriched
in the thin water film between hydrophobic surfaces in comparison with an infinitely thick
film that was scaled down to the same thickness.
By applying Equation 61 to films of 10 to 20 nm thickness, as examples, the film
excesses of N2 and O2 are about the same as the amount of dissolved gases in corresponding
(imaginary) thin films made up of bulk solution. This observation is in fair agreement with
the idea of an enhanced water structure in thin water films delimited by hydrophobic surfaces. The solvent properties of the water change as additional water cage volume becomes
available to the (clathrate-forming) gas molecules, similar to when the temperature is
reduced to below room temperature.
Relatively speaking, the higher content of the gas solutes for thin films will cause the
film tension to become more negative in accordance with Equation 61. Conversely, deaeration should tend to diminish the magnitude of the attractive hydrophobic force, as observed
by Meyer, Lin, and Israelachvili (2005). Moreover, it seems most likely that the effects of
deaeration documented by Pashley (2003) on the stability of colloids and emulsions can also
be understood, at least in a preliminary way, on this basis.
Referring to Equation 58, an analogous equation for ΔΓ sf,ex is
∂ ln B
Bbh ∂ ln b ⁄ d μ s ⎞
ΔΓ sf,ex = – ------------ Δγ f – --------- ⎛⎜-----------------------------⎟
4π ⎝
∂μ s
2
sinh ( bh ⁄ 2 ) ⎠
(EQ 62)
which should be relevant insofar as the solute concentration is kept low and the extent of
solute adsorption is limited on the hydrophobic surfaces. From this expression, an increase
in ΔΓ sf,ex , such as found in the case under discussion, is strongly coupled with the constant
B becoming larger as the chemical potential μs is raised: the H-bond network in the film is
reinforced by introducing inert gas molecules.
Concerning the addition of alcohols and surfactants that readily adsorb on hydrophobic surfaces, in the dilute regime the situation is principally opposite: The surface force rises
with μs, yielding a negative ΔΓ sf,ex that, according to the quasi-thermodynamic relation
(Equation 62), can be interpreted in terms of a smaller constant B and a diminishing decay
length b–1. This is true for alcohols as well as nonionic surfactants. For a contact monolayer
mixed with such surface-active species, there is simply less free energy to be gained by
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
171
0.10
0.05
F/R, mN/m
0.0
–0.05
Water
–0.10
12.5% Ethyl Alcohol
20% Ethyl Alcohol
–0.15
–0.20
0
10
20
30
40
50
60
Separation, nm
FIGURE 19 Surface force isotherms presented by Ederth (1999) for water and ethanol–water
mixtures between hexadecanethiol surfaces, showing the pronounced effect on the
hydrophobic attraction of adding alcohol. The advancing contact angles are 94° (12.5%) and 88°
(20%), respectively.
restructuring (smaller a). Evidently, for ionic surfactants, the electrostatic (double layer)
repulsion must also be invoked.
Surface force curves for water and ethanol–water mixtures recorded by Ederth (1999)
using C16-alkylthiolated gold surfaces are reproduced in Figure 19. It appears that the surface force is substantially reduced in magnitude by adding alcohol. Using Equation 61, estimate ΔΓ sf,ex to –4 × 1016 molecules m–2 for a 10-nm-thick film in contact with ≈16%
alcohol solution to be compared with the alcohol content of a corresponding bulk solution
film: ≈ 2 × 1018 m–2. The reason why the surface force diminishes as alcohol is added is that
the lowering of the film tension γf is larger for an infinitely thick film than for a thinner film.
In turn, this is because adding alcohol to the thin film counteracts the favorable structure formation.
On the other hand, adding salts should not change the situation very much with respect
to the hydrophobic force (provided that the metastable attachment of hydrocarbon chains
remains unaffected), the chief reason being that small ions do not mix with the water in the
contact monolayers adjacent to the hydrophobic surfaces. Hence, the B constant should be
left much the same, and likewise b–1, insofar as the ions do not interfere significantly with
the H-bond network formation. This agrees with a multitude of observations using hydrophobic surfaces that are sufficiently stable.
Hydrophobic Forces in Flotation
In flotation, hydrophobic particles are selectively collected on bubble surfaces and separated
from the hydrophilic particles suspended in aqueous slurry. Thermodynamically, the bubble–
particle adhesion occurs when contact angle of the particle is larger than zero. In view of
their high interfacial tensions in water, air bubbles should be considered hydrophobic. Thus,
the bubble–particle interaction occurring during flotation may be seen as a hydrophobic
interaction, the kinetics of which is controlled by the surface forces involved, namely electrostatic, van der Waals, and hydrophobic forces. The electrostatic forces are repulsive when
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FLOTATION FUNDAMENTALS
both particles and bubbles are negatively charged. The van der Waals forces operating in
wetting films are also repulsive according to the Lifshitz theory. Therefore, it is difficult to
explain flotation without considering the existence of the attractive hydrophobic force discussed in this chapter. In flotation, the role of the hydrophobic force is to reduce the energy
barrier so that the bubble–particle interaction becomes a “fast heterocoagulation” process.
The process can be “slow” if the hydrophobic force is too small to counterbalance the repulsive electrostatic force. It is well known that flotation rate increases with increasing particle
hydrophobicity and decreasing double-layer potentials.
In general, surface forces are weaker than the hydrodynamic forces operating in a flotation cell by orders of magnitude. As a particle approaches a bubble in close proximity, however, the hydrodynamic force is much reduced because of the hydrodynamic resistance
against film thinning. The hydrodynamic forces become comparable to the surface forces as
the separation distance between bubble and particle reaches the critical rupture thickness of
the wetting film between the two surfaces. By assuming that the probability of the bubble–
particle adhesion is determined by the hydrodynamic and surface forces, a flotation rate
equation has been developed both under quiescent and turbulent flow conditions (Schimoller, Luttrell, and Yoon 1994; Yoon and Mao 1996; Mao and Yoon 1997; Do and Yoon
2005). This approach made it possible to develop a model that can predict flotation rates
using both hydrodynamic and surface chemistry parameters.
Hydrophobic forces also play a role in particle–particle interactions. Hydrophobic particles coagulate at a pH well above the isoelectric point, which cannot be explained without
assuming the existence of a hydrophobic force (Xu and Yoon 1989, 1990). The hydrophobic coagulation, which is driven by the hydrophobic force, also plays an important role in
flotation. In general, flotation rate decreases with decreasing particle size. Gaudin (1957)
showed, however, that the flotation rate of galena particles stayed constant at particle sizes
<10 μm, which may be attributed to the hydrophobic coagulation. Thus, use of a stronger
collector that can increase the hydrophobicity of mineral particles and, hence, increase the
particle size may be conducive to improving fine particle flotation.
S U M M A RY
In Jacob Israelachvili’s book, Intermolecular and Surface Forces (1991), he writes, “It is the
energy (or entropy) associated with the H-bonding network and proton hopping defects,
which extends over much larger space than the molecular correlations, that is probably at
the root of the long-range solvation interactions of water.” After having scrutinized the scientific issues standing today regarding the origin of the amazingly strong and long-ranged
hydrophobic attraction, the authors agree with Israelachvili. No equally convincing
approach to gain a closer understanding of the hydrophobic surface force has so far been
proposed. Various electrostatic interaction mechanisms have been subjected to critical tests,
and likewise the adhering–bridging bubble track. Yet, none of these approaches has been
applicable for the most ideal hydrophobic surfaces one can prepare where the hydrophobic
attraction is strongest. In the end, they have to be abandoned as possible starting points of a
more elaborate theoretical modeling.
Two decades ago, the original proposition that the hydrophobic surface force is related
to the structural response of water contacted with a solid hydrophobic surface appeared as
quite a natural hypothesis. After some time, however, a paradigm shift occurred within the
scientific community, and eventually the concept of water restructuring was considered to
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
173
be nothing but a “bold conjecture,” maybe even too bold to be taken seriously. Today, however, a vast amount of experimental results has been accumulated. Together, these results
indicate that there is hardly any alternative route that might be fruitful when aiming at a
molecular understanding of the hydrophobic attraction.
The versatility exhibited by water molecules to form clathrate cages for guest molecules
of various size and shape indicates that reliable predictions of the structural state of water,
close to a real hydrophobic surface, will be difficult to make, particularly because the molecular description of liquid water is still much in dispute. However, water has long been
known to be an associated liquid where three-dimensional networks play a significant role.
Furthermore, the association process is of a cooperative nature, implying that the free energy
per water molecule depends on the cluster size. Therefore, it may not be surprising that the
introduction of an (infinite) hydrophobic solid surface can promote an extensive structure
generation in the adjacent water, and that confinement of water in a thin film between two
hydrophobic surfaces might strongly promote additional restructuring.
Given the background presented in this chapter to attempt resolving the long-standing
issue about the origin of the hydrophobic force, it seems most urgent to (a) make several
more studies of the temperature dependence of the hydrophobic force with the purpose of
fully assessing the thermodynamic water film properties; (b) probe the state of the water
next to carefully prepared and characterized hydrophobic surfaces, as well as in thin water
films sandwiched between hydrophobic surfaces, using, for example, the SFG or XAS technique; and (c) make extensive experimental studies of the effect on the hydrophobic surface
force of various clathrate-forming inert gases and other well-chosen solutes.
APPENDIX
T H E R M O DY N A M I C S O F A P L A N A R , T H I N A Q U E O U S F I L M
B E T W E E N H Y D R O P H O B I C S O L I D S U R FA C E S
Treating the thermodynamics of a (symmetric) thin aqueous solution film between two
plane-parallel hydrophobic solid surfaces starts with the thermodynamic fundamental equation of the film in the following form:
d γ f = – ( S f ⁄ A )dT + hdp – π D dh – Γ wf d μ w – Γ sf d μ s
(EQ A.1)
Here, it is presupposed that the properties of the thin film are estimated by using the
equimolecular dividing surfaces with respect to the solid component, that is, by using the
(solid) dividing surfaces in the Gibbs notation. Thus, the film considered that has the thickness h (equal to the distance between the two dividing surfaces) only encompasses liquid
matter in the form of water and a solute component denoted by subscript s. In the Equation
A.1, p is the ambient pressure, πD is the disjoining (interaction) pressure, and γf is the film
tension. This equation is a straightforward extension of the Gibbs surface tension equation.
As considered by Gibbs (1961) and treated in more detail by Eriksson (1969), the latter
equation only holds true as long as the state of strain of the solid surfaces remains
unchanged.
It is emphasized that γf is energetically (rather than mechanically) defined, by means of
the following relation:
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174
FLOTATION FUNDAMENTALS
γ f = G f ⁄ A – Γ wf μ w + Γ sf μ s ≡ G f, ex ⁄ A = Ω f ⁄ A
(EQ A.2)
Accordingly, γf is primarily the free energy per unit area compared to what would be
noted for a thin film of bulk phase properties containing the same amounts of water and solute. Stated another way, it corresponds to the reversible cleavage work up to a certain surface
separation h. In addition, it is formally the same as the Ω potential per unit area of the thin film.
For the special case under discussion, however, regarding a thin film between hydrophobic surfaces and made out of a dilute water solution, γf has also a rather clear-cut
mechanical significance. Because only minor contributions to the excess free energy as well
as to the lateral tensions arise in the hydrocarbon part of the two interfaces, the integrated
tangential pressure pT profile across the film thickness h to a good approximation corresponds to –γ f.
Next, Equation A.3 must be brought in harmony with the phase rule by invoking the
Gibbs–Duhem relation for the bulk phase (denoted by superscript b):
S b dT – V b dp + n w d μ w + n s d μ s = 0
(EQ A.3)
Upon eliminating dμw and switching to the solute concentration cs as variable from
(A.1), the following is obtained:
∂μ s ⎞
d ln c s
d γ f = – ( S f, ex ⁄ A )dT + ( V f, ex ⁄ A )dp – π D dh – Γ sf, ex ⎛ ----------⎝ ∂ ln c s⎠ T, p
(EQ A.4)
where the coefficients dT, dp, and dμs are given by
S f, ex ⁄ A = S f ⁄ A – Γ wf s wb – Γ sf s sb
(EQ A.5)
V f, ex ⁄ A = h – Γ wf ν wb – Γ sf ν sb
(EQ A.6)
Γ sf, ex = Γ sf – Γ wf c S ⁄ c w
(EQ A.7)
and
In these expressions, sb and νb denote partial molar volumes in the bulk solution, and cw and
cs represent the bulk phase concentrations of water and solute, respectively.
The (reversible) surface force measured in a surface force device is proportional to the
film tension difference Δγ f = γ f ( h ) – γ f ( ∞ ) for which the following equation is obtained,
assuming an ideal solution of the solute and noting that π D ( ∞ ) = 0 :
d Δγ f = – ( Δ S f, ex ⁄ A )dT + ( Δ V f, ex ⁄ A )dp – π D dh – Γ sf, ex RTd ln c s
(EQ A.8)
with obvious definitions of the coefficients dT, dp, and dμs as the differences for the properties in question between a thin and an infinitely thick film. Because almost all residual contributions from the solid surfaces to the thermodynamic film properties cancel upon
forming these differences, the constraint of a constant surface strain can be disregarded, and,
furthermore, Δγf corresponds almost exactly to the change in overall lateral mechanical tension
of the film when comparing a thin and a very thick film; compare the equivalent equation
presented by Gibbs for the noninteracting case (Equation 678 in work by Gibbs [1961]).
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THE NATURE OF HYDROPHOBIC ATTRACTION FORCES
175
For the case of 1:1 electrolyte being the solute, the last term in Equation A.8 must be
multiplied by a factor of 2. Hence, by plotting Δγf as a function of the state variables T, p,
and cs for fixed surface separations h, important information can be gathered as to the thermodynamic thin film (excess) properties ΔSf,ex/A, ΔVf,ex/A, and ΔΓ sf,ex . Using Equation A.2
and the free enthalpy definition results in
Δ G f, ex ⁄ A = Δ H f, ex ⁄ A – T Δ S f, ex ⁄ A = Δγ f
(EQ A.9)
implying that the corresponding enthalpy difference is readily derived. In particular, gaining
knowledge about ΔSf,ex/A and ΔHf,ex/A will be crucial for understanding the structural
aspects of thin aqueous films, whereas the effect of solutes on the surface force is intimately
related with ΔΓ sf,ex . In all likelihood, producing the proper sets of data and analyzing them
thermodynamically might greatly contribute to resolving the current research issues about
the nature and properties of thin water films between hydrophobic surfaces.
AC K N OW L E D G M E N T S
The authors wish to thank Atte Kumpulainen and Jinhong Zhang for many illuminating
discussions and for assistance with the figures. Appreciation is also extended to Per Claesson
for critically reading the manuscript and making several useful comments.
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Adsorption of Surfactants and Its
Influence on the Hydrodynamics
of Flotation
P. Somasundaran, L. Zhang, T.W. Healy, W. Ducker, R. Herrera-Urbina,
and M.C. Fuerstenau
INTRODUCTION
Flotation is a process that separates various particles based on the differences in their surface
properties. In this process, hydrophobic particles, or hydrophilic particles that are made
hydrophobic by surface-active reagents (surfactants), attach to the gas bubbles in the pulp
and are levitated to the froth, with most other particles remaining in the bulk. The former
particles can, thus, be separated from the matrix by separating the froth from the bulk pulp
(Gaudin 1957).
Flotation was first used in the mineral industry as an enrichment process for sulfide ores
about 100 years ago (Glembotskii, Klassen, and Plaksin 1972). Considerable advances in
flotation were made after the development of flotation reagents that can control or modify
the surface properties of minerals selectively. Because of its high selectivity and relatively low
cost, flotation became the economically competitive process to treat low-grade and complex
ores that would have otherwise been treated as waste. Today, billions of tons of ores are
treated by the flotation process annually, and about 95% of the base metals produced are
beneficiated through this process (Matis and Zouboulis 1995). Beyond the mining industry,
the flotation process has been modified and used in many other industries, such as chemical,
paper, food, pharmaceutical, recycling, and remediation (Matis and Zouboulis 1995).
Flotation depends, first of all, on the probability of attachment of the bubbles to the
particles. The ability of the gas bubbles to attach to the particles is a function of the hydrophobicity of the particles. When a solid particle is attached to a gas bubble in a liquid, the
resulting contact angle is given by Young’s equation in terms of the interfacial energies (see
Figure 1):
γ sg – γ sl = γ lg cos θ
(EQ 1)
where γsg , γsl, and γlg are solid–gas, solid–liquid, and liquid–gas interfacial tensions, respectively, and θ is the contact angle. Adhesion of the gas bubbles to the particles in the suspension requires a contact angle that is significantly larger than zero. The larger the contact
angle, the greater is the hydrophobicity of the particle, and the easier it is for the bubble to
attach and result in flotation. Thus, the difference in hydrophobicities of various particles
can be used as the basis of their separation by flotation.
Few minerals, such as natural sulfur and graphite, are naturally hydrophobic. In most
cases, hydrophobicity is imparted to the mineral particles by the adsorption of surface-active
agents (surfactants and some polymers). Adsorption of surfactants and polymers on minerals
179
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180
FLOTATION FUNDAMENTALS
γlg
Gas
Liquid
θ
γsl
γsg
Solid
FIGURE 1 Solid–liquid–gas interface depicting contact angle and surface tension forces
acting at the boundaries
can lead to changes in their wetting and dispersion behavior and the resultant flotation
response. The present review focuses on adsorption of surfactants at mineral–solution
interfaces.
The action of surfactants to control the hydrophobicity of the sulfide–water interface
differs in a fundamental way from the effect of surfactants on other mineral–water interfaces. In sulfide systems, the principal control variable is the redox potential imposed on the
mineral–water system. By adjusting that potential, either through the use of a polarized electrode or by the use of a redox couple in the solution phase, sulfides can acquire hydrophobicity. Often, but not always, that redox couple is supplied by a surfactant moiety, the most
famous being the xanthate–dixanthogen redox couple.
The principal variable is pH for minerals such as the inorganic oxides (e.g., silica, alumina, goethite), soluble salt minerals (e.g., halite, sylvite), alkaline earth carbonates (e.g., calcite), fluorides, (e.g., fluorite), phosphates (e.g., apatite), sulfates (e.g., barite), feldspars, and
complex aluminosilicates. This chapter will show that, in a both a simple and also in a derivative way, pH is able to define the parameter of proton activity that is always necessary to
begin to understand surfactant adsorption in such systems. The variable of pH is also important in sulfide flotation systems.
In contrast to sulfide flotation systems, it is rarely necessary to consider redox effects to
understand surfactant adsorption in nonsulfide flotation systems. In the nonsulfide mineral
adsorption–flotation area, if the redox potential is changed for any reason (i.e., the Eh is
changed), there can be a concomitant change in pH. That pH change can, in turn, affect the
interfacial processes in nonsulfide systems.
Surfactant molecules contain both hydrophilic and hydrophobic moieties. They can
modify the interfacial properties significantly at very low concentrations. Also, they can selfassemble into aggregates in solutions and at solid–liquid interfaces through the hydrophobic effect above certain concentrations. The conformation of surfactants at solid–liquid
interfaces also has a governing role in determining the interfacial properties. The effect of
surfactant adsorption on major interfacial parameters can be seen in Figure 2, where the
adsorption density of the cationic surfactant, dodecyl ammonium acetate (DAA), on quartz
is given along with contact angle, zeta potential, and flotation recovery (D.W. Fuerstenau,
Healy, and Somasundaran 1964). For DAA and quartz at neutral pH, increase in adsorption
due to association of surfactants adsorbed at the solid–liquid interface into two dimensional
aggregates called hemimicelles occurs at about 10–4 M DAA. This marked increase in
adsorption density is accompanied by concomitant sharp changes in contact angle, zeta potential,
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ADSORPTION OF SURFACTANTS
+80
Contact Angle
Adsorption Density
Zeta Potential
Flotation Response
80
30
20
60
50
0
40
30
–40
20
10
10
0
10–6
+40
Zeta Potential, mV
70
Relative Flotation
Contact Angle, degrees and
Adsorption Density, mol/cm 2, × 1011
40
181
10–5
10–4
10–3
–80
10–2
Concentration of DAA, mol/L
Source: D.W. Fuerstenau, Healy, and Somasundaran 1964.
FIGURE 2 Correlation of adsorption density, contact angle, flotation response, and zeta
potential for quartz as a function of the cationic surfactant DAA concentration at pH 6 to 7
and flotation recovery, suggesting that these phenomena depend primarily on the adsorption of surfactants at the solid–liquid interface.
Flotation, being the result of fruitful contacts among solution, solid, and gas phases, is
governed by various processes at all of the solid–liquid–gas interfaces in systems. In fact, in
addition to adsorption at the solid–liquid interface, surfactant adsorption at the liquid–gas
interface also plays an important role in affecting the flotation (Somasundaran 1968). Furthermore, flotation is controlled not only by the physical chemistry of the system but also by
the hydrodynamics of the pulp. In this chapter, the physicochemical as well as hydrodynamics aspect of the flotation process are addressed. Adsorption of surface-active agents (including surfactants and certain polymers) and their effects on flotation processes are discussed.
Important effects of water chemistry of surfactants and minerals on the flotation process are
also reviewed. Application of various spectroscopic and imaging techniques to explore the
characteristics of the adsorbed layers, and particularly conformation and orientation, is also
examined.
A D S O R P T I O N O F S U R FA C TA N T S O N M I N E R A L S
Adsorption of surfactants at solid–liquid interfaces can drastically alter interfacial properties such as wetting and hydrophobicity that are critical to the flotation process. Separation
of minerals from each other depends on the selective adsorption of surfactants on only the
mineral particles to be floated. An understanding of the mechanism of adsorption is necessary to choose appropriate flotation reagents, particularly collectors, for optimum separation conditions.
Adsorption is the selective partitioning of the surfactant into the interfacial region as
the result of energetically favorable interactions among the surfactant, the solid surface, and
solution. The driving force for adsorption involves one or more of the following factors:
coulombic interaction, ion exchange covalent bonding, desolvation of the polar group of the
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182
FLOTATION FUNDAMENTALS
collector, desolvation of the surface and hydrogen bonding, and hydrophobic and van der
Waals interactions. The adsorption density of the surfactant is commonly interpreted using
the Stern–Grahame equation:
0
– ΔG ads
⎞
Γ δ = 2rC exp ⎛ ----------------⎝ RT ⎠
(EQ 2)
where Γδ is the adsorption density in the Stern plane δ, r is the effective radius of the adsorbing ion, C is the bulk concentration of the surfactant, R is the gas constant, T is the absolute
0 is the standard free energy of adsorption (D.W. Fuerstenau 1971).
temperature, and – ΔG ads
The driving force for adsorption is the sum of several contributing forces mentioned previously and can be written as follows (D.W. Fuerstenau 1971):
0
0
0
0 + ΔG 0 + ΔG 0
– ΔG ads
= ΔG elec
+ ΔG chem
+ ΔG hyd
H
H2 O
(EQ 3)
0 is the electrostatic interaction term and is equal to zFψ where z is the valence of the
ΔG elec
δ
adsorbate species; F is the Faraday constant, and ψδ is the potential in the δ plane where the
0
surfactant head group is located; ΔG chem
is the chemical term due to covalent bonding;
0
and ΔG hyd is the free energy gained when hydrocarbon moieties are transferred from bulk
water into bulk (e.g., micellar or other bulk self-assembly structures) or interfacial domains
or regions. The energy gained through such transfers is often a linear function of the energy
gained per –CH2– group. The magnitude of the energy gained per –CH2– decreases from
that of the process of transfer from water to bulk saturated hydrocarbon oil, to that of transfer to a spherical micelle, to that of transfer to a two-dimensional aggregate at a hydrophobic
solid–water interface, to that of a two-dimensional aggregate at a hydrophilic solid–water
interface (Lin and Somasundaran 1971). In each case, there is a gain in entropy—that is, an
increase in the magnitude of ΔS0 for the system, hydrocarbon chain-water, as one transfers
the chain of –CH2– groups from bulk water to the new bulk or interfacial aggregate state.
0 is the hydrogen bonding term, and ΔG 0
ΔG H
H 2 O is the term owing to dehydration of the
adsorbate species or any species displaced from the interface because of adsorption. For each
surfactant–solid system, several of the previously mentioned terms can be contributing,
depending on the solid and the surfactant type, surfactant concentration, temperature, and
so forth. In the following section, the major forces involved in surfactant adsorption are discussed.
Electrostatic Adsorption of Surfactants
In systems where the surfactants and the particles are oppositely charged, electrostatic interactions play a governing role in the adsorption process. Charge on the particle surface, in the
case of oxides, is a result of the hydrolysis of surface species followed by pH-dependent dissociation of the surface hydroxyl groups:
MOH = MO – + H +
(EQ 4)
+
H + + MOH = MOH 2
(EQ 5)
where M represents the interfacial metal atom. The pH at which the surface charge is zero is
called the point of zero charge (PZC). Oxides carry a positive charge in solutions that are
more acidic than the PZC and a negative charge in those that are more alkaline. For salt-type
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ADSORPTION OF SURFACTANTS
183
minerals such as calcite and apatite, the charge generation could be due to preferential dissolution of lattice ions, followed by hydrolysis of dissolved species in the bulk and subsequent
adsorption of the resulting complexes.
Alkyl Sulfonate–Alumina. Much detailed research work on surfactant adsorption via
electrostatic attraction has been conducted on alumina using alkyl sulfonates (Somasundaran and Fuerstenau 1966; Wakamatsu and Fuerstenau 1968; D.W. Fuerstenau and Wakamatsu 1975) and the alkyl benzene sulfonate–alumina systems (Dick, Fuerstenau, and
Healy 1971; Healy, Somasundaran, and Fuerstenau 2003). Experimentally obtained adsorption isotherms of sodium dodecyl sulfonate (SDS) on alumina at fixed pH and ionic
strength are S-shaped and exhibit three distinct regions in the pH range 3.2 to 8.2 when the
isotherm is plotted on a log–log scale (see Figure 3). In these regions, the adsorption density
increases linearly with the equilibrium surfactant concentration, but it does so to different
extents. The increase in adsorption density with surfactant concentration is small in Region
I, but it abruptly becomes large with small surfactant concentration increments in Region II,
and again is small in Region III. These isotherms have strongly suggested that in Region I, at
low concentrations, the surfactant anions adsorb as individual counterions on the positivelycharged surface of alumina (PZC at pH 9.2). Measurements of the electrophoretic mobility
of colloidal alumina in aqueous SDS support the proposed physical adsorption mechanism
involving mainly an electrostatic attraction between dodecyl sulfonate anions and positivelycharged alumina through an ion-exchange process (Somasundaran and Fuerstenau 1966).
10–9
pH 3.2
4.2
5.2
6.2
2
Amount Sulfonate Adsorbed, mol/cm
7.2
10–10
8.2
10–11
8.6
10–12
10
pH 9.2
–13
10–6
10–5
10–4
10–3
Equilibrium Concentration of SDS, mol/L
Source: D.W. Fuerstenau and Wakamatsu 1975.
FIGURE 3 Adsorption isotherms for the aqueous SDS–alumina system at several pH values,
25°C, and constant ionic strength of 2 × 10–3 M NaCl (sodium chloride)
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184
FLOTATION FUNDAMENTALS
In Region II, the slope of the isotherm becomes steeper, suggesting that at higher concentrations specific adsorption phenomena may occur through association of the hydrocarbon chains of surfactant anions adsorbed electrostatically at the Stern plane (Somasundaran
and Fuerstenau 1966). This association enhances the adsorption of sulfonate anions considerably and leads to the formation of two-dimensional aggregates of adsorbed surfactant
anions at the surface, a possibility that was first proposed by D.W. Fuerstenau (1953) and by
Gaudin and Fuerstenau (1955), who termed these bi-dimensional surfactant aggregates
hemimicelles. In this region, the adsorption mechanism involves electrostatic attraction
between the sulfonate anions and the positively-charged alumina surface, and hemimicelle
formation through association of the hydrocarbon chains of the adsorbed sulfonate ions.
The equilibrium SDS concentration at which Region II begins has been identified as the
critical hemimicelle concentration (Somasundaran and Fuerstenau 1966).
Region III of the adsorption isotherm is also characterized by a low dependence of the
adsorption density on the surfactant concentration, a behavior that is more marked at lower
pH values where the slope decreases significantly. This adsorption behavior seems to indicate saturation of sulfonate ions adsorbed electrostatically at the interface. Under these conditions, a further increase in adsorption density with sulfonate concentration has been
explained in terms of bi-layer adsorption of sulfonate ions (Somasundaran and Fuerstenau
1966). In addition to the pH, the ionic strength and the hydrocarbon chain length also
affect the adsorption of alkyl surfactant ions at the interface.
Effect of Hydrocarbon Chain. The shape of the adsorption isotherm of alkyl sulfonate
on alumina is dependent on the hydrocarbon chain length (Wakamatsu and Fuerstenau
1968). As shown in Figure 4, three-region adsorption isotherms are obtained with decyl-,
dodecyl-, and tetradecyl sulfonates at pH 7.2. In the case of octyl sulfonate, its adsorption
isotherm exhibits single linear behavior (corresponding to Region I) all along the range of
equilibrium concentrations of sodium octyl sulfonate investigated, and indicates that there
is no hemimicelle formation for this relatively short-chain sulfonate. The alkyl sulfonates
containing a hydrocarbon chain length of 16 carbon atoms give an adsorption isotherm with
only two distinct linear portions (Regions II and III) with different slope. These isotherms
clearly show that Region I is independent of chain length and that the onset of hemimicelle
formation (beginning of Region II) shifts to lower sodium alkyl sulfonates as the chain
length increases. The onset of Region II corresponds to the following bulk concentration of
sodium alkyl sulfonates: 7 × 10–4 M for C10, 6 × 10–5 M for C12, and 1.4 × 10–5 M for C14
(D.W. Fuerstenau 2002).
The standard free energy of adsorption for Regions II and III is given by the following
expression:
0
0
0
ΔG ads
= ΔG elec
+ ΔG CH
= zFψ δ + nφ
2
(EQ 6)
where n is the number of carbon atoms in the chain, and φ is the free energy decrease upon
removal of 1 mol of CH2 group from water. In Region III, the electrostatic and hydrophobic
bonding contributions to the free energy of adsorption are opposite in sign.
Recently, Healy, Somasundaran, and Fuerstenau (2003) have reported the isotherms for
the adsorption of a variety of branched and straight-chain alkyl benzene sulfonates (ABSs)
on a low-surface-area alumina at different pH values. Similar to the adsorption isotherms
previously obtained with a high-surface-area alumina (Dick, Fuerstenau, and Healy 1971),
in all cases the adsorption density shows an abrupt increase at a particular equilibrium
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ADSORPTION OF SURFACTANTS
185
10–8
10–9
Adsorption Density, mol/cm 2
Monolayer
10–10
10–11
10–12
C16
C14
C12
C10
C8
10–13
10–6
10–5
10–4
10–3
10–2
Equilibrium Concentration of Sodium Alkyl Sulfonate, mol/L
Source: Wakamatsu and Fuerstenau 1968.
FIGURE 4 Adsorption isotherms for the aqueous sodium alkyl sulfonate–alumina systems at
pH 7.2, 25°C, constant ionic strength of 2 × 10–3 M NaCl, and several hydrocarbon chain lengths
concentration of the sodium alkyl benzene sulfonate, identified as the hemimicelle concentration (Figure 5). As opposed to the adsorption of alkyl sulfonates on alumina, which
shows significant pH dependence, the adsorption of alkyl benzene sulfonates is independent
of pH over a wide range. In addition, a very remarkable finding of these results is the close
similarities between the form of the adsorption isotherms of alkyl and alkyl aryl sulfonates,
suggesting that the benzene ring adds the equivalent of 3.2 CH2 groups of the hydrophobic
energy to the overall adsorption process, as was postulated by Dick Fuerstenau, and Healy
(1971). Because the C9 and C10 alkyl benzene sulfonate isotherms exhibit the characteristic Regions I, II, and III of alkyl sulfonate isotherms, the hemimicelle hypothesis is equally
applicable to ABS adsorption on alumina.
Alkyl Ammonium Adsorption on Quartz. The experimental adsorption isotherm presented in Figure 6 shows the adsorption density of dodecyl ammonium ions on quartz as a
function of the equilibrium concentration of DAA at neutral pH (deBruyn 1955). At concentrations below the critical micelle concentration (CMC), this alkyl ammonium salt dissociates completely at this pH, undergoes protolysis, and the total concentration of added
surfactant is virtually in the cationic form. This isotherm is typical of the cationic surfactant
adsorption from solution at the oxide mineral–aqueous solution interface, and shows that at
concentrations lower than 2 × 10–4 M DAA, the adsorption density follows a linear behavior that can be expressed mathematically by the following empirical relation:
flotation0.book Page 186 Tuesday, January 2, 2007 7:36 PM
186
FLOTATION FUNDAMENTALS
10–9
Adsorption Density of ABS Surfactant, mol/cm 2
III
10–10
II
10–11
SO3–Na+
C12
I
10–12
SO3–Na+
C10
C12 SO3–Na+
C9
10–13
10–6
SO3–Na+
10–5
10–4
10–3
Equilibrium Concentration of ABS, mol/L
Source: Healy, Somasundaran, and Fuerstenau 2003.
FIGURE 5 Isotherms for the adsorption of C9, C10, and C12 alkyl benzene sulfonates on
alumina at pH 7.2 and 10–3 M NaCl
Γ DAA = 8.1 × 10 –9 C 0.5
(EQ 7)
where ΓDAA is given in moles per square centimeter and C is the bulk concentration of DAA
in moles per liter. In this region, the dodecyl ammonium cations adsorb electrostatically, and
their adsorption density is proportional to the square root of the equilibrium concentration
of DAA, as predicted from the Gouy–Chapman equation.
The slope of the adsorption isotherm presented in Figure 6 changes abruptly at an equilibrium concentration close to 2 × 10–4 M DAA in solution. This break in the curve provides the concentration at which the alkyl chains of the adsorbed cationic surfactants begin
to associate at the interface through hydrophobic bonding to form hemimicelles (Gaudin
and Fuerstenau 1955). Above this concentration, the adsorption density increases significantly and even exceeds monolayer coverage. Under these conditions, this linear behavior of
the adsorption density can be represented by the following empirical equation:
Γ DAA = 2.2 × 10 –6 C 1.2
(EQ 8)
Excellent evidence supporting the hemimicelle premise is the flotation data provided by
D.W. Fuerstenau, Healy, and Somasundaran (1964). Flotation recovery increases dramatically above specific concentrations for each of the amines of varying hydrocarbon chain
length (Figure 7). Similar to deBruyn’s (1955) results, the rapid rise in flotation recovery
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ADSORPTION OF SURFACTANTS
187
Adsorption of Dodecyl ammonium Salt, mol/cm 2
10–8
10–9
Monolayer
10–10
10–11
10–12
10–13
10–8
10–7
10–6
10–5
10–4
10–3
10–2
DAA Concentrate, mol/L
Source: deBruyn 1955.
FIGURE 6
Isotherm for the adsorption of dodecyl ammonium ions on quartz at pH 6–7
100
Flotation Recovery, %
80
60
40
C18
C16
C14
C12
C10
C8
C6
10–3
10–2
C4
20
0
10–8
10–7
10–6
10–5
10–4
10–1
1
Concentration, mol/L
Source: D.W. Fuerstenau, Healy, and Somasundaran 1964.
FIGURE 7 The effect of alkyl chain length on the relative flotation response of quartz in the
presence of alkyl ammonium acetate solution
with dodecyl amine also occurs about 1 × 10–4 M. These authors showed that a plot of collector concentration at which a rapid rise in flotation recovery occurs as a function of the
number of carbon atoms in the hydrocarbon chain is a straight line with a slope that corresponds to a specific adsorption potential of –0.62 kcal/mol CH2 group. This is the freeenergy decrease associated with hydrocarbon chain removal from solution.
flotation0.book Page 188 Tuesday, January 2, 2007 7:36 PM
188
FLOTATION FUNDAMENTALS
100
+20
Contact Angle, degrees
Surface Coverage, % of monolayer
Zeta Potential, mV
Flotation Recovery, %
90
0
–20
70
–40
60
–60
50
–80
40
Zeta Potential, mV
Contact Angle, Surface Coverage, Flotation Recovery
80
–100
30
–120
20
10
0
0
2
4
6
8
10
12
14
pH
Source: D.W. Fuerstenau 1957.
FIGURE 8 Correlation of adsorption density, contact angle, zeta potential, and flotation
recovery of quartz in presence of 4 × 10–5 M DAA
The important role that surfactant adsorption exhibits in determining interfacial properties is shown as a function of pH for the dodecyl amine–quartz system in Figure 8. The
marked increase in adsorption density is accompanied by concomitant sharp changes in
contact angle, zeta potential, and flotation recovery. These results were explained on the
basis of hemimicelle formation involving co-adsorption of RNH3+ and RNH2(aq) on the
quartz surface. Equilibria are (Ralston 1984; Smith 1973)
RNH 2 ( s ) ↔ RNH 2 ( aq )
RNH 2 ( aq ) + H 2 O ↔ RNH 3 + + OH –
K = 2 × 10 –5
K = 4.3 × 10 –4
(EQ 9)
(EQ 10)
Laskowski, Vurdela, and Liu (1988) have suggested that the colloids of precipitated
amine formed in alkaline solution may be responsible for quartz flotation under these conditions (see Figure 9). The amine colloids are charged in solution, and they have a PZC at
relatively high pH; for example, dodecyl amine has a PZC at pH 11. The pH region in
which the amine colloids form is the same as that in which optimal flotation of quartz is
obtained.
The association of hydrocarbon chains either as hemimicelles or as a precipitate of collector salt on the mineral surface is desirable for flotation. Interactions in the interfacial
region depend essentially on the relative concentrations of surfactant required to form
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ADSORPTION OF SURFACTANTS
189
DAA-HCI
Break of Transmittance
Leveling Off of Transmittance
10–1
Micellization
CMC
Concentration, mol/L
10
–2
Critical pH of Precipitation
Precipitation
(Colloidal Solution)
10–3
No Precipitation
10–4
Solubility Limit
10–5
6.0
8.0
10.0
12.0
pH
Source: Laskowski, Vurdela, and Liu 1988.
FIGURE 9 Thermodynamic equilibrium diagram showing the CMC (30°) and the lines of
critical pH of precipitation and solubility limit at 25°C. The points shown are experimental
solubility ( ) and redispersion ( ) determined from the transmittance curves.
hemimicelles and to precipitate the surfactant salt. If the HMC (concentration of hemimicelle formation) is the lower of the two, the formation of hemimicelles would be preferred
over salt precipitation.
The adsorption density of dodecyl ammonium species on quartz has also been determined at pH 5 and pH 9.8 (Takeda and Usui 1987), and the extent of adsorption has been
correlated with the flotation behavior of –10 μm quartz particles. Complete quartz flotation at pH 5 was achieved at a DAA adsorption density of about 1% of a monolayer,
whereas at pH 9.8 greater than a monolayer was required.
Chemical Adsorption of Surfactants
“Chemical adsorption,” or “chemisorption,” is adsorption in which the forces involved are
valence forces of the same kind as those operating in the formation of chemical compounds
(Everett 1972). Chemisorption is involved in systems in which the metal ions do not leave
their lattice sites, and, hence, adsorption is limited to a monolayer of surfactant ions. Surfactant ions that adsorb on mineral surfaces through chemical forces show more selectivity
toward a particular mineral and, thus, provide the means for more effective mineral separations by froth flotation. Chemisorption takes place between the polar head group of the surfactant (cationic or anionic) and the anionic or cationic surface sites, respectively, through
flotation0.book Page 190 Tuesday, January 2, 2007 7:36 PM
190
FLOTATION FUNDAMENTALS
10
10–4 M
Collector Adsorption Density, μmol/m 2
PZC
3 × 10
1
10
–5
–3
10–5
10–4
10–5 M
0.1
PZC
0.01
Alumina (SDS)
Hematite (Sodium Oleate)
I = 2 × 10–3 M, 23°C
0.001
4
5
6
7
8
9
10
11
pH
Source: D.W. Fuerstenau 1984.
FIGURE 10 Isotherms for the adsorption of dodecyl sulfonate on alumina (physical
adsorption) and of oleate on hematite (chemical adsorption) as a function of pH at 23°C
covalent or coordinate bonds, and results in the formation of a new surface compound.
Therefore, when ionic surfactants chemisorb on mineral surfaces, they can adsorb even on
charged surfaces with an electric charge similar to that of the surfactant polar head group.
In systems where chemisorption takes place, the PZC of the mineral is not determinant
of surfactant adsorption as it is in those systems in which the collector ion adsorbs electrostatically. This is clearly apparent in Figure 10. It is noted that adsorption of dodecyl sulfonate is possible only below the PZC of alumina. In the case of hematite, however,
adsorption of oleate occurs greater than two pH units above the PZC under which conditions both the surface and oleate have the same charge. This was also clearly shown by
Iwasaki, Cooke, and Choi (1960), who investigated the flotation response of hematite
(PZC at pH 6.7) as a function of pH using three unsaturated fatty acids that are known to
react chemically with iron. These researchers found that this hematite sample floats well up
to about pH 9 with linolenic, pH 10 with linoleic, and pH 11 with oleic acid.
Surface Reaction/Precipitation
Another type of chemical adsorption is involved when metal ions comprising the mineral
move out of their lattice sites and interact with the collector; this has been termed “surface
reaction.” Surface reaction is probably the most dominant chemical adsorption mechanism
flotation0.book Page 191 Tuesday, January 2, 2007 7:36 PM
ADSORPTION OF SURFACTANTS
191
in flotation. What has been termed chemisorption in the past was probably surface reaction.
Examples include xanthate–sulfide minerals, fatty acid–semisoluble salts, and fatty acid–
hydroxamate-insoluble oxides and silicates.
Sulfide Minerals. In those sulfide systems in which surface reactions are involved (e.g.,
galena, chalcocite, and sphalerite), adsorption occurs in two stages. In the case of galena, the
first stage comprises chemisorption of xanthate ion on surface lead sites. Woods (1976) has
presented a voltammogram established with a galena electrode in the absence and presence
of ethyl xanthate. In both cases, the electrode behaves reversibly. The anodic curve in the
presence of ethyl xanthate exhibits a peak at about 0 V. The discharge reaction given in
Equation 11 is believed to account for this peak, and the total charge passed was found to
correspond to monolayer coverage.
X – ↔ X ads + e –
(EQ 11)
With surface oxidation, metal collector salts form at the surface as lead ions move from
the lattice into solution. An infrared adsorption study of ethyl xanthate on a precipitated
lead sulfide (PbS) film was presented by Leja, Little, and Poling (1962–1963). In Figure 11,
the top curve (a) shows the infrared spectrum of bulk lead ethyl xanthate. The curve in (b)
shows the spectrum of the PbS film that has been oxidized in air. After exposing this film to
ethyl xanthate, the spectrum located in (c) was obtained, which is nearly identical to that of
bulk precipitated lead ethyl xanthate. Prolonged rinsing with ether removed much of the
xanthate from the surface (d), but it took a strong solvent, pyridine, to return the surface to
that of oxidized PbS (e).
1,100
1,000
1,110
1,200
700
1,110
990
D
E
987
1,020
1,110
C
1,115
0
50
1,195
0
50
1,140
1,140
0
50
B
1,020
1,112
1,210
986
0
Absorption, %
800
A
50
50
900
1,014
1,212
1,300 cm–1
100
0
Source: Leja, Little, and Poling 1962–1963.
FIGURE 11 Infrared spectra showing adsorption of ethyl xanthate onto an evaporated PbS
film: (a) bulk lead ethyl xanthate, solid on Nujol mull; (b) freshly evaporated PbS film after
atmospheric oxidation; (c) above treated in aqueous solution of ethyl xanthate; (d) after
prolonged washing in ether; and (e) after washing in pyridine.
flotation0.book Page 192 Tuesday, January 2, 2007 7:36 PM
192
FLOTATION FUNDAMENTALS
14
12
Milliequivalents
10
8
rfa
6
4
U
nl
ea
a
ch
bl
e
on
M
in
er
a
ce
u
lS
Cuprous
Xanthate
2
Excess
Xanthate in
Filtrate
0
0
2
4
6
8
10
12
14
Xanthate, meq
Source: Gaudin and Schuhmann 1936.
FIGURE 12
chalcocite
Distribution of amyl xanthate species after adsorption of amyl xanthate on
In their work with chalcocite, Gaudin and Schuhmann (1936) observed an unleachable
layer of xanthate and a leachable layer of cuprous xanthate on chalcocite after contact with
xanthate. These results are shown in Figure 12.
The presence of bulk precipitated zinc dodecyl xanthate on sphalerite after contact of
sphalerite with dodecyl xanthate was shown by Yamasaki and Usui (1965). Similar phenomena were also observed for hexyl xanthate by M.C. Fuerstenau, Clifford, and Kuhn (1974)
who also showed that flotation of unactivated sphalerite is controlled by the formation and
adsorption of zinc xanthate for various xanthates. These observations are in agreement with
those of Plaksin and Anfimova (1954) who concluded that two forms of adsorption occur
in this system.
Sulfide minerals are electronic conductors and, as such, develop a potential when
placed in water. This potential is termed the “rest potential.” When the rest potential is
greater than the equilibrium oxidation–reduction potential of a sulfhydryl collector, such as
xanthate, oxidation of the collector to a dithiolate occurs. Examples of this phenomenon
when xanthate is used as collector are pyrite, arsenopyrite, and pyrrhotite. In the case of xanthate, the oxidized species is dixanthogen. Oxygen or ferric iron can serve as xanthate oxidant. In the case of oxygen, the following reaction occurs at the pyrite surface.
2X – + 1 ⁄ 2O 2 ( aq ) + H 2 O ↔ X 2 ( ads ) + 2OH –
(EQ 12)
The mechanism of adsorption of dixanthogen on pyrite is not clear, but it is thought to
involve interaction between the sulfur of the collector molecule and sulfur atoms contained
in the pyrite surface. What is known, though, is that chemisorption is not involved. This
flotation0.book Page 193 Tuesday, January 2, 2007 7:36 PM
ADSORPTION OF SURFACTANTS
193
40
Xanthate Addition
None
1 × 10–4 M C2
Zeta Potential, mV
20
0
–20
–40
–60
0
2
4
6
8
10
12
14
pH
Source: M.C. Fuerstenau, Kuhn, and Elgillani 1968.
FIGURE 13 Zeta potential of pyrite as a function of pH in the absence and presence of ethyl
xanthate in the presence of air
many may be seen from Figure 13, which shows that the zeta potential of pyrite is the same
in the absence and presence of xanthate. Chemisorption would have resulted in the formation of a new phase on the pyrite surface and a corresponding difference in zeta potential.
Insoluble Oxides/Silicates. Oleate species have been found to adsorb on relatively
insoluble oxide minerals, such as hematite, even when the surface is highly negatively
charged (Kulkarni and Somasundaran 1980; Yap et al. 1981). Under these conditions,
adsorption phenomena in this system can only be explained by specific adsorption of the
oleate anion on hematite. Because it has been shown that oleate reacts with iron species and
forms iron oleate compounds (Peck, Ruby, and Wadsworth 1966), specific adsorption of
oleate on hematite can be explained by chemisorption of oleate ion on iron surface sites,
thus producing an iron oleate compound at the interface.
With several insoluble oxides and silicates, flotation takes place in the pH region in
which hydrolysis of metal ions comprising the mineral occurs (Peterson et al. 1965; M.C.
Fuerstenau and Rice 1968; Palmer, Fuerstenau, and Aplan 1975). The adsorption potential
of hydroxy complexes is high, which is apparent in the adsorption of calcium species on
quartz presented in Figure 14 (Clark and Cooke 1968). Modest adsorption of Ca2+ occurs
below pH 10. However, at about pH 11, the pH at which Ca2+ starts to hydrolyze to
CaOH+ and Ca(OH)2(s), extensive adsorption takes place. A speciation diagram for calcium species is presented for comparison in Figure 15.
A flotation system illustrating these phenomena well is the pyrolusite–oleate system
(Figure 16). The flotation response at about pH 4 is due to physical adsorption of oleate,
whereas the response at about pH 9 is due to surface reaction between MnOH+ and oleate
ion (M.C. Fuerstenau and Rice 1968). This is the pH region of hydrolysis of Mn2+ so that
dissolution of Mn4+ and subsequent reduction of Mn2+ must occur. Because of the stability
of the hydroxy complexes, dissolution of the surface metal atoms is facilitated by hydroxyl
ion at these pH values.
Research on chrysocolla with hydroxamate as collector revealed the presence of multilayers of cupric hydroxamate by visual observation (Peterson et al. 1965). Chrysocolla
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194
FLOTATION FUNDAMENTALS
40
Initial Ca2+ = 100 mg/kg
2
10
Adsorption, mol/cm × 10
30
20
10
Surface
Precipitation
0
4
6
8
10
12
14
pH
Source: Clark and Cooke 1968.
FIGURE 14 Adsorption of calcium species on quartz as a function of pH from solutions 100 ppm
in calcium
10–3
Ca2+
Concentration, M
10–4
CaOH+
10–5
Ca(OH)2(s)
10–6
10
11
12
pH
FIGURE 15
Speciation diagram for 1 × 10–3 M Ca2+
13
14
flotation0.book Page 195 Tuesday, January 2, 2007 7:36 PM
ADSORPTION OF SURFACTANTS
195
100
Flotation Recovery, %
80
60
40
20
0
0
2
4
6
8
10
12
14
pH
Source: M.C. Fuerstenau and Rice 1968.
FIGURE 16
Flotation recovery of pyrolusite as a function of pH with 1 × 10–4 M oleate
100
3.3 × 10–4 M; 23°C
1.0 × 10–4 M; 22°C
5.0 × 10–5 M; 49°C
5.0 × 10–5 M; 22°C
Flotation Recovery, %
80
60
40
20
0
4
5
6
7
8
9
10
11
pH
Source: Peterson et al. 1965.
FIGURE 17 Flotation recovery of chrysocolla as a function of pH with various concentrations
of octyl hydroxamate
changed color from its characteristic blue to apple green, the color of cupric hydroxamate.
Confirmation was made with infrared absorption analysis. Maximal flotation response occurs
at about pH 6, the region of hydrolysis of Cu2+ to CuOH+ and Cu(OH)2(s) (Figure 17).
The beneficial role of elevating temperature can also be noted, enhancing the dissolution of
the chrysocolla. Nagaraj and Somasundaran (1979) also observed favorable flotation of
chrysocolla with a hydroxyoxime (benzophenone) at pH 5–6 and also at pH 10, but to a
lesser extent.
Adsorption isotherms of octyl hydroxamate on chrysocolla were established by HerreraUrbina (1985) at various values of pH (Figure 18). Of the pH values involved, maximal
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196
FLOTATION FUNDAMENTALS
12
Hydroxamate Adsorption Density, mol/g × 10 4
Chrysocolla
–200 Mesh
2 × 10–3 M KNO3, 2 g/L
10
8
6
4
pH
5.9
7.2
9.0
11.2
2
0
4
5
6
7
8
9
Equilibrium Hydroxamate Concentration, mol/L × 10
10
11
4
Source: Herrera-Urbina 1985.
FIGURE 18 Isotherms for the adsorption of octyl hydroxamate on chrysocolla at ionic
strength of 2 × 10–3 M KNO3 at several pH values
adsorption is noted at pH 5.9, corresponding to the flotation data in Figure 17. The change
in slope indicates a change in mechanism of hydroxamate adsorption (i.e., the concentration
at which cupric hydroxamate is formed and adsorbed).
Multilayers of ferric hydroxamate were also observed on hematite (M.C. Fuerstenau,
Harper, and Miller 1970). A change in color was observed before and after contact with
octyl hydroxamate. These authors also noted that increasing the conditioning time (and,
hence, dissolution) with hematitic ores increased flotation recovery significantly.
A good representation of chemisorption and surface reaction between relatively insoluble oxides and silicates, and chemically bonding reagents is given in Figure 19.
Semisoluble Salts. The adsorption of oleate on such semisoluble minerals as calcite
(Somasundaran 1969; Rao and Forssberg 1991a; Young and Miller 2000), apatite
(Moudgil, Vasudevan, and Blackmeer 1987; Rao and Forssberg 1991a; Lu, Drelich, and
Miller 1998), fluorite (Rao and Forssberg 1991a; Mielczarski et al. 2000), and scheelite
(Rao and Forssberg 1991b) has been investigated by various researchers using spectroscopic
techniques and the traditional depletion from solution method. Depending on the experimental conditions and method used, in many cases, the adsorption isotherms reported in
the literature for the same mineral are significantly different. Both the pH and the solid–
liquid ratio have been found to determine the extent of oleate adsorption on these minerals.
In addition, the shape of the adsorption isotherm even depends on the method of sampling
when depletion of oleate from solution is analyzed (Mielczarski et al. 2000). If the isotherm
is determined from the total uptake of oleate, the adsorption density increases monotonically with the residual oleate concentration, but if the precipitated metal oleate is considered, then the isotherm exhibits a maximum in adsorption density. Oleate reaction with
sparingly soluble salt minerals is more complex because in addition to true chemisorption,
surface reaction and even precipitation of oleate compounds in the bulk can occur. If the
metal cation from the mineral lattice forms insoluble metal oleates, one can predict a sharp
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ADSORPTION OF SURFACTANTS
197
Chemisorption
Me2+
+
Xm–
Me2+
2+
+
HXm
Me
2+
Xm
2+
+
OH
–
Me
2+
OH
Me
Me
Xm–
–
+
+
H
+
H+
–
Adsorption with Surface Reaction
2+
+
HXm
Me
2+
+
Xm
(MeXm)
Me2+
+
OH–
(MeOH)+
+
–
HCO3
+
Xm
Me
Me
2+
–
(MeXm)+
+
(MeHCO3)
+
Surface Reaction
(MeOH)+
–
(MeXm)+
+
OH–
+
+
H2 O
+
H+
(MeOH)
+
+
HXm
(MeOH)
+
+
–
OH
Me(OH)2
(MeOH)+
+
2HXm
Me(Xm)2
(MeXm)
+
H2O
Source: D.W. Fuerstenau and Pradip 1984.
FIGURE 19 Representation of chemisorption and surface reaction between relatively
insoluble minerals and chemical bonding reagents
increase of the adsorption density of the oleate under conditions where the metal oleate is
more stable (insoluble) than the mineral. This adsorption behavior is clearly shown in Figure 20, which presents the isotherm for the adsorption of oleate ions at the calcite–aqueous
solution interface at pH 9.6 (Somasundaran 1969). At low residual sodium oleate concentration in the bulk, the adsorption density increases steadily as this concentration increases.
At about 3 × 10–5 M residual sodium oleate, however, the isotherm rises steeply, which suggests that at concentration higher than this value, calcium oleate deposits onto the calcite surface.
Young and Miller (2000) have reported their results on the adsorption of oleate on calcite. These researchers determined an in situ near-infrared-based adsorption isotherm at pH
9.2 and 20°C, and found that it compared well with isotherms reported previously. Their
isotherm for the adsorption of oleate on calcite shows two regions. At concentrations lower
than 1 × 10–5 M, oleate adsorption is explained in terms of chemisorption. At higher oleate
concentrations, adsorption is attributed to multilayer formation and precipitation of calcium dioleate.
Oleate adsorption on apatite was studied by Moudgil, Vasudevan, and Blackmeer
(1987). An adsorption isotherm with four distinct regions is presented in Figure 21. In
Region I, in which the oleate concentrations are below that necessary for calcium oleate precipitation, adsorption is attributed to chemisorption. In Region II, precipitation of calcium
dioleate and its adsorption with oleate ions is postulated. In Region III, surface precipitation
is assumed to be dominant with some bulk precipitation of calcium oleate. Adsorption in
Region IV is assumed to approach the limit of surface saturation.
Soluble Salts. The nature of soluble salt flotation is considerably more complex than
that of other groups of minerals because of the high ionic strengths of the solutions, which is
on the order of 5 M. Flotation occurs only after precipitation of collector salt in some cases,
though it is not necessary in other systems. Various mechanisms have been advanced to
explain soluble salt flotation, which are ion exchange (D.W. Fuerstenau and Fuerstenau
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198
FLOTATION FUNDAMENTALS
10–4
Oleate Adsorbed, mol/g of calcite
10
–5
10–6
10
–7
10–7
10–6
10–5
10–4
Residual Oleate Concentration, mol/L
Source: Somasundaran 1969.
FIGURE 20
Isotherm for the adsorption of oleate on calcite at pH 9.6
102
5.0 wt %
10.0 wt %
Adsorption Density, μmol/g
101
100
10–1
10–2
I
10–3
10–7
10–6
II
10–5
III
IV
10–4
Residual Oleate Concentration, kmol/m
Source: Moudgil, Vasudevan, and Blackmeer 1987.
FIGURE 21
Isotherm for adsorption of oleate on apatite
3
10–3
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ADSORPTION OF SURFACTANTS
199
1956), insoluble reaction product between surface ions and collector (Halbich 1933), a
crystallographic properties model (Bachmann 1951), collector and surface hydration model
(Rogers and Schulman 1957), surface charge–ion pair model (Roman, Fuerstenau, and
Seidel 1968), and a water breaker/former model (Hancer and Miller 2000). A method
involving surface reaction is the dodecyl amine–sylvite system. As shown in Figure 22, flotation occurs only after precipitation of dodecyl ammonium chloride occurs. This is also true
for tetradecyl amine. In the case of octyl amine, however, flotation occurs prior to precipitation of octyl amine chloride, but relatively high concentrations are necessary. Adsorption
isotherms of various amines on sylvite were established by Schubert (1971) and by Matthe
and Schneider (1977; Figure 23). Comparison of the isotherm for dodecyl ammonium
100
Dodecyl
Tetradecyl
40
Solubility Limit
60
Solubility Limit
Flotation Recovery, %
80
20
0
10
–6
10
–5
10
–4
10
–3
10
–2
Amine Addition, mol/L
Source: Roman, Fuerstenau, and Seidel 1968.
FIGURE 22
Flotation recovery of sylvite with various amines
101
C10
C12
C18
Fractional Surface Coverage
100
C8
10–1
10
–2
Sylvite (KCl)
RNH3Cl
10–3
10–4
–8
10
10
–7
10
–6
10
–5
10
–4
10
–3
10
–2
10
–1
Alkylammonium Chloride Concentration, mol/L
Source: Matthe and Schneider 1977.
FIGURE 23 Adsorption isotherms of alkly ammonium chlorides on potassium chloride in
saturated brine
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200
FLOTATION FUNDAMENTALS
chloride with the flotation data reveals a change in slope of the isotherm at about the concentration at which precipitation of dodecylammonium chloride occurs, resulting in greater
adsorption of amine on the sylvite. The fatty acid–halite system is another that exhibits surface reaction (i.e., flotation occurs only after precipitation of the sodium carboxylate in solution). The complexity of these systems is further demonstrated by sylvite flotation occurring
with alkyl sulfonates prior to precipitation of collector salts.
Hydrogen Bonding
Hydrogen bonding between surfactant species and the particle surface species is responsible
for adsorption of a number of reagents, particularly those containing hydroxyl, phenolic,
carboxylic and amine groups. For example, adsorption of nonionic surfactant ethoxylated
alcohols and sugar-based alkyl glucosides has been proposed to involve hydrogen bonding
(Doren, Vargas, and Goldfarb 1975; Zhang et al. 2002). In relation to claims of hydrogen
bonding, the hydrogen bond formed between the collector functional group and a mineral
surface moiety is stronger than the hydrogen bond formed by that moiety and interfacial or
bulk water molecules.
Desolvation Energies
When a hydrated surfactant head group leaves the bulk water environment and inserts itself
into the interfacial region of the mineral–water system, there must be a partial desolvation
of the secondary salvation shell around the surfactant head group ( James and Healy 1972).
Desolvation is, thus, an unfavorable process, and the desolvation free-energy change is necessarily positive.
E F F E C T S O F WAT E R C H E M I S T R Y O N S U R FA C TA N T – M I N E R A L
I N T E R AC T I O N S
In addition to mineral-surfactant interactions, water chemistry also plays an overwhelming
role in the adsorption process by affecting the various interactions among mineral, surfactant
and solution. These interactions include the surfactant-solution equilibria, the mineralsolution equilibria and, consequently, the interactions between the surfactants and the
mineral particles. The interactions in mineral–surfactant–solution systems include dissociation,
micellization and precipitation of the surfactant, dissolution of solids in response to solubility product equilibria followed by hydrolysis, complexation and precipitation of the dissolved species, and the interactions between dissolved mineral species with surfactant in the
bulk in various forms. The dissolved species, including those introduced due to dissolution
from all the minerals present in the ore and those from the water source, are the major elements that affect the water chemistry.
Water Chemistry of Flotation Reagents
Long-chain fatty acids, such as oleic acid, are among the commonly-used flotation reagents
in nonsulfide flotation. Flotation processes using fatty acids are greatly affected by solution
conditions such as pH, as weakly acidic fatty acids will undergo associative interactions that
can influence their adsorption and flotation properties. For example, oleic acid undergoes
dissociation to form ions (Ol–) at high pH values and neutral molecules (HOl) at low pH
values. In the intermediate region, the ionic and the neutral molecular species can associate
to form dimer complexes [(Ol)2H]–. As the surfactant concentration is increased, micellization
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ADSORPTION OF SURFACTANTS
201
–4.0
Log (Total Oleate Activity)
–5.0
–6.0
–7.0
Ksol = 2.51 × 10–8
–8.0
–9.0
0
2
4
6
8
10
12
pH
Source: Ananthapadmanabhan 1980.
FIGURE 24
pH of oleic acid precipitation
or precipitation of the surfactant can also occur in the solution. In addition, surfactant species can associate to form other aggregates such as the dimer ion (Ol22–) in pre-micellar solutions. Also, oleic acid has very limited solubility, which is a sensitive function of pH
(Ananthapadmanabhan and Somasundaran 1988); Ananthapadmanabhan 1980). The pH
of precipitation of oleic acid calculated as a function of total oleate is shown in Figure 24.
The solution equilibria of oleic acid (HOl) are expressed as below. R represents oleate ion.
HOl l = H + + Ol –
pK sp = 7.6
(EQ 13)
RH ( aq ) = R – + H +
pK a = 4.95
(EQ 14)
2Ol – = ( Ol ) 22–
pK d = – 2.3
(EQ 15)
R – + RH = R 2 H –
pK = – 4.95
(EQ 16)
The species distribution of oleic acid as a function of pH based on the above equilibria
at a given concentration is shown in Figure 25.
1. The pH of the precipitation of oleic acid at the given concentration is 7.6, as shown
in Figure 24.
2. The activities of oleic monomer and dimer remain almost constant above the precipitation pH and decrease sharply below it.
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202
FLOTATION FUNDAMENTALS
–4
K-Oleate 3 × 10
–5
–
R
kmol/m3
Log (Activity of the Species)
R22–
–6
R2H–
–8
–10
RH
3
4
6
8
10
12
13
pH
Source: Ananthapadmanabhan 1980.
FIGURE 25 Oleate species distribution as a function of pH (total oleate concentration = 3 × 10–5 M)
3. The activity of the acid-soap (Ol)2H– exhibits a maximum in the neutral pH range.
4. The dominant equilibria at pH 8 and 10–4 mol dm–3 is HOl in equilibrium with
Ol–. Above pH 8 and 10–4 mol dm–3 total oleate, a small region of dimeric (solid)
exists before micellization or precipitation occurs.
Clearly, to understand the role of adsorption of collectors on minerals, the effects of
concentration, pH, ionic strength, and activities of the various possible collector species on
the adsorption process must be considered.
Water Chemistry of Minerals
When mineral particles come in contact with water, they undergo dissolution, the extent of
which is dependent on solubility product of the mineral and the solution composition. The
dissolved mineral species can undergo further reactions, such as hydrolysis, complexation,
adsorption, and precipitation. The complex equilibria involving all such reactions can be
expected to determine the interfacial properties of the minerals and their flotation behavior.
In a system containing soluble or sparingly soluble minerals in which the extent of dissolution is markedly higher than that in most oxide–silicate systems, the effect of dissolved mineral species can be drastic.
In the case of carbonaceous phosphate ores containing such sparingly soluble minerals
as apatite and calcite, depending on the solution conditions, the surface of apatite can be
converted to calcite and vice versa through surface reactions or bulk precipitation of the
more stable phase. The stoichiometry of the equilibrium governing the conversion of apatite
to calcite can be written as follows (Amankonah, Somasundaran, and Ananthapadmanabhan
1985a, b):
Ca 10 ( PO 4 ) 6 ( OH ) 2 ( s ) + 10CO 32– = 10CaCO 3 ( s ) + 6PO 42– + 2OH –
(EQ 17)
It can be seen from this equation that, depending on the pH of the solution, apatite can
be converted to calcite if the total carbonate in solution exceeds a certain value. In fact, the
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ADSORPTION OF SURFACTANTS
203
60
2 × 10–3 kmol/m3 KNO3
Calcite in Water
Apatite in Water
Apatite in Calcite Supernatant
Calcite in Apatite Supernatant
Zeta Potential, mV
40
20
0
–20
–40
5
6
7
8
9
10
11
12
13
pH
Source: Somasundaran, Amankonah, and Ananthapadmanabhan 1985.
FIGURE 26 Illustration of the effect of supernatants on the zeta potential and isoelectric point
of calcite and apatite
amount of dissolved carbonate from atmospheric CO2 does exceed that required to convert
apatite to calcite under high-pH conditions.
Surface conversion due to the reaction of the dissolved species with the mineral surface
has been proven experimentally; electrokinetic data obtained for the calcite–apatite system
in water and in the supernatant of each other are shown in Figure 26. When apatite is contacted with calcite supernatant, its zeta potential shifts to that of calcite and vice versa, suggesting surface conversion of apatite to calcite and calcite to apatite, respectively
(Somasundaran, Amankonah, and Ananthapadmanabhan 1985).
It is clear that dissolution equilibria of dissolved mineral species can alter the mineral
surface properties to such an extent that surface properties of various minerals become indistinguishable, leading to loss of selectivity in the adsorption of reagents.
Role of Water Chemistry in Adsorption of Surfactant on Minerals
Chemical equilibria in aqueous solutions containing both the minerals and the surfactants
can be expected to be much more complex than those in either of the individual systems discussed previously. In addition to surfactant adsorption at the solid–liquid interface, interactions between dissolved mineral species with various surfactant species can be expected. All
these interactions can affect the surfactant adsorption and the subsequent flotation of the
mineral phase.
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204
FLOTATION FUNDAMENTALS
From this discussion on apatite–calcite conversion, it is clear that a flotation separation
scheme designed on the basis of the surface properties of a single mineral is not likely to perform satisfactorily. The effect of dissolved species of calcite and apatite on fatty acid flotation of both minerals has, in fact, been studied using mineral supernatant solutions
containing various dissolved species. Both supernatants of calcite and apatite are found to
depress calcite flotation by oleic acid in the tested pH range, with apatite supernatant exhibiting a greater depressing effect. Similar results have also been obtained for apatite flotation.
The supernatants of calcite and apatite depress the apatite flotation under all tested pH conditions (Ananthapadmanabhan and Somasundaran 1984). These observations clearly show
the crucial role of water chemistry in the flotation of mineral–surfactant systems.
Adsorption Under Plant Conditions
Flotation is a dynamic process. Kinetics of surfactant adsorption on solids and flotation can
be expected to be affected by the total solution chemistry. In fact, in actual plant operations,
equilibrium adsorption/desorption may not be reached and, hence, it is the kinetics of these
processes that may dominate. The effect of water chemistry on oleic acid adsorption on
francolite during anionic conditioning has been studied under plant conditions along with
laboratory conditions for reference. (Figure 27). The laboratory process used low-solids
loading (10 wt %), whereas under plant conditions the solids loading was 72 wt %. Adsorption is markedly higher under laboratory conditions than under plant conditions. On the
other hand, under plant conditions, the adsorption is similar in distilled water and plant
water. Interestingly, the effect of dissolved species is reduced under plant conditions. It is evident, then, that the adsorption of surfactant on a mineral is a complicated process involving
1.E-04
Oleic Acid Adsorption Density, mol/m 2
1.E-05
1.E-06
1.E-07
1.E-08
Laboratory Condition, Distilled Water
Plant Condition, Distilled Water
Plant Condition, Plant Water
1.E-09
1.00E-06
1.00E-05
1.00E-04
1.00E-03
1.00E-02
1.00E-01
Conditioning Time, min
FIGURE 27 Adsorption isotherms of oleic acid adsorption on francolite in distilled water and
plant water under laboratory and plant conditions
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ADSORPTION OF SURFACTANTS
205
interactions, such as surfactant self-association, mineral dissolution, bulk precipitation,
adsorption, and surface precipitation. These interactions are further complicated by the
kinetic effects of the various reactions.
S P E C T R O G R A P H I C A N D I M A G I N G S T U D I E S O F S U R FA C TA N T
ADSORPTION
Because of the important role of adsorbed surfactant layers in adsorption and subsequent
flotation processes, considerable work has been done to elucidate the physicochemica1 and
hydrodynamic mechanisms that govern the formation of interfacial states. The structure of
these states plays a critical role in determining the interfacial processes, and information on
such structures, particularly at microscopic and molecular levels, can be helpful for controlling flotation. Various spectroscopic and imaging techniques such as fluorescence, electron
spin resonance (ESR), Fourier transform infrared, and excited-state resonance Raman have
been applied in mineral–surfactant systems to obtain information on microstructure of
adsorbed layers on a molecular level. Other advanced techniques that are being explored to
probe the nanostructures in solution and on solids include analytical ultracentrifugation,
small-angle neutron scattering, neutron reflection, and atomic force microscopy (AFM).
These techniques are reviewed here.
FLUORESCENCE SPECTROSCOPY
Fluorescence emission is the radiative emission of light by an excited molecule returning to
its ground state energy level. The nature of this phenomenon depends significantly on the
environment of the light-absorbing species and has been exploited for a long time in support
of exploring the solution behavior of surfactants. The fluorescence measurements are generally carried out by a steady-state fluorescence spectrofluorometer. The lifetime of fluorescence is determined by a time-resolved fluorescence lifetime instrument. The dependence of
fluorescence intensity and lifetime on the physicochemical environment of the fluorescing
molecules has been studied extensively. The technique has been adapted to investigate the
structure of adsorbed layers at the mineral surfaces to obtain information on the micropolarity, the microviscosity of the probe environment, and the aggregation number of the surfactant species adsorbed at the interface.
The surfactant aggregation process on the surface of a solid as described and shown in
Figure 3 has been a generally accepted model for surfactant adsorption on oppositely
charged solids. Through the fluorescence spectroscopic technique, the exact nature of these
aggregates in the adsorbed layers has been revealed.
In steady-state fluorescence spectroscopy, the ratio of relative intensities of the third
peak to the first peak, I3/I1, on a pyrene emission spectrum shows the greatest solvent
dependency. This ratio increases as the polarity decreases and, hence, can be used to estimate
the solvent polarity of an unknown microenvironment in which the pyrene probe is situated. I3/I1 values for pyrene were determined for alumina–SDS–water systems for various
regions of the adsorption isotherm. The data are shown in Figure 28, which is marked by an
abrupt change in the local polarity of the probe from an aqueous environment to a relatively
nonpolar micelle-type environment. This abrupt change occurs in a region that is well
below the CMC and corresponds with the transition in the adsorption isotherm from
Region I to II (i.e., the hemimicelle concentrations). In the plateau region, the I3/I1 value
coincides with the maximum I3/I1 value for SDS solutions (Figure 28a), indicating the
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206
FLOTATION FUNDAMENTALS
A
1.5
1.0
I3/I1
CMC
0.5
0.0
0
10–5
10–4
10–3
10–2
Dodecyl Sulfate Concentration, mol/L
B
1.5
SDS Micelle (1.1 M NaCl)
I3/I1
1.0
Water (0.1 M NaCl)
0.5
0.0
0
10–5
10–4
10–3
10–2
Residual Dodecyl Sulfate, mol/L
Source: Chandar, Somasundaran, and Turro 1987.
FIGURE 28 I3/I1 fluorescence parameter of pyrene in (a) SDS solutions in 0.1 M NaCl; (b) SDS–
alumina slurries in 0.1 M NaCl, pH 6.5; (I3 = 383 nm, I1 = 374 nm)
completion of aggregation on the surface. Thus, fluorescence studies provide a means to
measure the micropolarity of the adsorbed layers.
In the fluorescence lifetime method, when a fluorescence probe is excited by a short
nanosecond pulse of light, its decay is enhanced in the presence of molecules that act as
quenchers. The lifetime of the probe under quenching conditions would be determined by
the concentrations of the quencher and the probe as well as the rate at which they diffuse
and encounter each other. Kinetic analysis of the fluorescence decay profiles can therefore
provide information on the local concentration of the reactants and hence the size of the
local aggregates.
A kinetic analysis based on this relation was carried out from the decay profiles of
pyrene in the adsorbed layer of the alumina–SDS adsorption system. The SDS aggregation
numbers obtained are marked on the adsorption isotherm in Figure 29. The aggregates in
Region II appear to be of relatively uniform size (120 to 1,300). But in Region III there is a
marked growth in the aggregate size (160 to 360). These results are of special significance
with respect to the evolution of the structure of the adsorbed layer and their interfacial
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ADSORPTION OF SURFACTANTS
207
10–9
356
IV
196
Dodecyl Sulfate Adsorption, mol/cm 2
10–10
CMC
258
III
166
128
10
123
–11
121
II
49
10–12
66
10–13
I
SDS Only
SDS with Pyrene
10–14
10–5
10–4
10–3
10–2
Residual Dodecyl Sulfate, mol/L
Source: Chandar, Somasundaran, and Turro 1987.
FIGURE 29 Surfactant aggregation numbers determined at various adsorption densities in
alumina–SDS system in 0.1 M NaCl, pH 6.5 (average numbers at each adsorption density
shown along isotherm)
60
Zeta Potential, mV
40
20
II
I
0
III
–20
CMC
–40
IV
–60
–80
1.00E-06
1.00E-05
1.00E-04
1.00E-03
1.00E-02
Residual Dodecyl Sulfate, mol/L
Source: Chandar, Somasundaran, and Turro 1987.
FIGURE 30 Zeta potential of alumina as a function of equilibrium concentration of dodecyl
sulfate (designation of regions based on shape of isotherm in Figure 29)
effects in flotation. Regions II–IV are characterized by surfactant aggregates of limited size.
Here, the surface is still not fully occupied and enough positive sites are available. From the
zeta potential of about 40 mV, the surface charge is about 0.2 molecules/nm2 or perhaps as
high as 0.35 because the Debye length is so short. The transition between Regions II and III
has a surface excess of about 0.4 molecules/nm2. Hence, the transition point can be said to
occur when the surfactant approximately neutralizes the alumina surface charge (Figure 30).
Further adsorption occurs mainly by increasing the number of aggregates as revealed by a
near-constant aggregation number. The transition from Region II to III corresponds to the
zero zeta-potential condition of the mineral–surfactant system, and adsorption in Region III is
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208
FLOTATION FUNDAMENTALS
Region I
No Aggregation
Region II
Number of Aggregates
Increases ~120–130
Molecules per
Aggregate
Region III
Number of Aggregates
Increases >160
Molecules per
Aggregate
Source: Chandar, Somasundaran, and Turro 1987.
FIGURE 31 Schematic representation of the correlation of surface charge and the growth of
aggregates for various regions of the adsorption isotherm depicted in Figure 29
likely to occur through the growth of existing aggregates rather than the formation of new
ones. Here, the adsorption goes up by a factor of at least 5, while the aggregation number
goes up by about 2, so there must be about 2 1/2 times as many aggregates. The factor of 2 in
aggregate size indicates that adsorption possibly goes from a patchy monolayer (head facing
alumina) to a patchy bi-layer (one head facing solution, the other facing alumina). This is
possibly due to better gain in energy by its hydrophobic effect between the hydrophobic
tails of the already adsorbed surfactant molecule and the unadsorbed ones. Such a situation
can be expected to result in a reverse orientation of the surfactant molecules as illustrated in
Figure 31, where the whole process of adsorption has been schematically portrayed. These
studies on the adsorbed layer of SDS on alumina further confirm the earlier concepts of
hemimicellization. Surfactant aggregation occurs above a critical concentration referred to
as “hemimicellar concentration,” which is marked by a sharp increase in the adsorption isotherm, eventually leading to the formation of highly organized and finite size assemblies
even at relatively low surface coverages (Chandar, Somasundaran, and Turro 1987).
Electron Spin Resonance
The ESR spectroscopic technique deals with transitions induced between Zeeman levels of
a paramagnetic system situated in a static magnetic field. Only species with a magnetic
moment are capable of interacting with the magnetic field. Three types of ESR studies may
be applied to the micellar systems. They are spin-probing, spin-labeling, and spin-trapping
techniques. In the spin-probing technique, a molecule with a spin is externally added to the
system, whereas in the spin-labeling technique, a spin-bearing moiety through covalent
bonding forms a part of the molecule. The spin-trapping technique is mainly applied in the
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ADSORPTION OF SURFACTANTS
209
identification of radicals produced thermally, photochemically, or radiolytically by trapping
the radical through chemical reactions with a spin-trap and converting the radical into a stable free radical to be examined by ESR.
In the study of the alumina–SDS system, stable free-radical nitroxide spin labels were
chosen for use of ESR spectroscopy. These spin labels (in micromolar levels) were co-adsorbed
individually on alumina along with the SDS (Waterman et al. 1986). As shown in Figure 32,
the shape of the adsorption isotherm of SDS on alumina with probe is similar to that in the
absence of the probe, except at low SDS concentrations where an enhancement in SDS
adsorption is observed because of the synergistic co-adsorption of the surfactant with the
probe. The ESR spectrum of 16-doxylstearic acid probe in aqueous solution shows the typical isotropic three-line spectrum characteristic of the nitroxide. The spectra obtained from
the alumina–water interface, on the other hand, are distinctly different from the solution
spectrum, with three types of ESR spectra. At low SDS concentrations (<150 μM), the probe
aggregates heavily on the surface, leading to a spin-exchange narrowed spectrum consisting
of one broad peak (point A in Figure 32). As the SDS concentration is raised sufficiently to
allow significant SDS hemimicellar colloid formation at the interface, a sharper anisotropic
spectrum is obtained. In this region, SDS colloid formation leads to a breakup of surface
aggregates of the probes. Each nitroxide, then, no longer interacts strongly enough with
10–9
D
10–10
COOH
+
SDS Adsorption Density, mol/cm 2
H-O
B
10–11
C
10–12
A
50 Gauss
10–13
SDS + 10 μM 16-Doxylstearic Acid
SDS Only
10–14
10–6
10–5
10–4
10–3
10–2
SDS Equilibrium Concentration, M
Source: Waterman et al. 1986.
FIGURE 32 ESR spectra of 16-doxylstearic acid in SDS–alumina system (in 0.1 M NaCl, pH 6.5,
10 μM 16-doxylstearic acid) in various regions of the adsorption isotherm
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210
FLOTATION FUNDAMENTALS
80% Glycerol
+
Rotational Correlation Time × 1010
H-O
COOH
75% Glycerol
20 Gauss
SDS–Alumina
Hemimicelle
(0.1 M NaCl, pH 6.5)
33% Glycerol
SDS Micelle
(0.1 M NaCl)
Source: Waterman et al. 1986.
FIGURE 33 Comparison of ESR spectra of 16-doxylstearic acid in hemimicelles, micelles, and
ethanol–glycerol mixtures and corresponding rotational correlation times and corresponding
viscosity
other nitroxides to result in exchange narrowing of the spectrum, and the spectrum
obtained is anisotropic because of the high local viscosity. A calibration curve of relative
rotational correlation times of 16-doxylstearic acid probe in mixtures of ethanol–glycerol
(Figure 33) shows that the colloid spectrum is similar to that obtained in 75%–80% glycerol. At higher SDS concentrations, the spectrum remains essentially unchanged (point D
in Figure 32). This suggests that the aggregate structure does not change appreciably as a
function of SDS surface coverage. ESR spectroscopy with nitroxide spin probes is shown to
be a useful method for probing the microenvironments of surfactants and polymers.
Raman Spectroscopy
Unlike fluorescence and ESR spectroscopies—which depend on the use of an externally
added or labeled luminescent and free radical–bearing compounds or moieties and could
cause perturbation to the environment they are to probe—infrared and Raman spectroscopy are intrinsic probing techniques that can be used to study the system without disturbance. Raman spectroscopy has an edge over infrared absorption spectroscopy in that it is
ideal for an aqueous environment, versatile in ease of sample handling, and remarkable due
to the wide range (50–5,000 cm–1) over which surfactant spectra can be recorded.
Raman spectroscopy is essentially a scattering technique that relays information on the
vibrational modes of a molecule. Those vibrations causing a polarizability change of a molecule are Raman active. The Raman spectroscopic analysis or surfactant aggregates has
yielded information on the relative orientation of alkyl chains of the surfactant. The
excited-state Raman spectrum of Ru(bpy)32+ in micelles and in the adsorbed SDS layers on
alumina in situ has shown that several transitions are sensitive to the evolution of the nanostructure of colloids (Somasundaran et al. 1989) and that the surfactant adsorbs water on
reverse orientation in the final stages of the adsorption. These results suggest that timeresolved resonance Raman spectroscopy is a powerful diagnostic tool for exploring solid–
liquid interfaces.
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211
Atomic Force Microscopy
AFM Imaging. Atomic force microscopy, developed by Binnig, Quate, and Gerber
(1986), is an example of a near-field microscope. Prior to the invention of near-field microscopes, the detectors for microscopy were positioned many wavelengths away from the sample. Thus, particles (photons or electrons) that were reflected or transmitted through the
sample experienced diffraction. The principle behind a near-field microscope is that the
detector is kept very near (<5 nm) to the sample so that resolution is not diffraction limited.
The detector is also made very small, and the interaction with the sample has a very strong
distance dependence so that only a very small area of the sample is interrogated at any time.
AFM uses variation in surface forces to image a surface. Surface forces are ideal for producing contrast because they are ubiquitous, and they decay rapidly with distance (e.g., the van
der Waals force between two atoms decreases with distance to the power minus 7). The
basic elements within an AFM are listed as follows:
1. A very sharp probe—The dimensions of the sharp tip must be on the order of the
resolution required. For example, a tip of radius ~5 nm is required to image
adsorbed micelles.
2. A method to detect the force on the tip—The most common system employs a
weak cantilever spring. The force on the spring is determined from the end slope on
the spring. The end slope is measured by reflecting a low-power laser from the end
of the spring and measuring the displacement of the reflected beam using a positionsensitive detector (usually a split photodiode).
3. A scanning system—An AFM image is a two-dimensional array of measurements of
force as a function of position of the tip above the sample. Piezoelectric translation
stages are used to move the tip across the sample (X–Y dimensions) with a resolution of about 0.1 nm.
4. A translation stage for moving normal to the sample (Z dimension)—A coarse
motion controller brings the tip to within about 1 mm of the sample, and a fine
piezoelectric drive is used to change the displacement in the last few micrometers
with a resolution of about 0.1 nm.
5. A computer—The computer is used to control the translation stages and to record
and display the force as a function of position across the sample. This type of image
is called a “deflection image.”
Instead of measuring and recording the force at each point, a feedback loop can be
employed in which a force set-point is specified, and the feedback loop controls the Z-displacement of the fine piezoelectric drive. The computer is then made to display the signal to
the Z-drive. If one assumes that the force field above the sample is only a function of the
height above each point on the sample, then each movement of the Z-drive mimics the
undulating topography of the sample. Thus, a record of the Z-drive displacement as a
function of the X–Y displacement is a map of the surface topography. Such an image is
called a height image. A weak feedback loop is also used in a deflection image to prevent tip
crashes and to keep the tip near the surface of a sloped sample.
Many variants of AFM imaging have been developed, of which the most important are
tapping-mode AFM and friction- or lateral-force AFM.
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FLOTATION FUNDAMENTALS
Tapping-Mode AFM. In tapping-mode AFM, the tip is oscillated near the resonant frequency of the cantilever spring and the change in amplitude or phase of the vibration is
recorded. The principal advantage is that the tip does not slide across the sample in contact.
Thus, damage to both the tip and sample are reduced. In addition, the phase and amplitude
can be used to measure the viscoelastic properties of the sample (Hansma et al. 1994).
Friction- or Lateral-Force AFM. In friction- or lateral-force AFM, the lateral force on
the tip is measured and a map of the friction is recorded (Meyer and Amer 1990). The main
application of AFM imaging to flotation has been to study the distribution and shape of
adsorbed surfactant molecules. For flotation, the most important range of surfactant surface
excess is from a partial to a full monolayer of surfactant, where the solid has been rendered
hydrophobic though adsorption. Unfortunately, researchers have been quite unsuccessful in
imaging partial layers of reversibly adsorbed surfactant. The problem appears to be that the
surfactant can move around on the surface when subject to the force of the AFM probe.
When the mineral has been made hydrophobic by adsorption of surfactant, it is difficult to
image at a low force because the tip is pulled toward the mineral. The situation is better
when the surfactant adsorbs irreversibly (e.g., by reaction with the mineral or polymerization). This situation has been studied by the Schwartz group, who have monitored the
growth of islands of surfactant (Schwartz et al. 1992; Schwartz 2001).
When the concentration of surfactant is greater than about one-half the CMC, AFM
imaging can reveal the organization of the adsorbed surfactant (Ducker 2003). The principal finding is that surfactants usually associate into finite-sized micelles on surfaces (Manne
and Gaub 1995). This is direct evidence for that which was inferred from the aggregation
numbers measured by fluorescence measurements (described previously). The geometry of
the micelle is not always the same as in bulk solutions, but the following list offers some
guiding principles:
1. On hydrophobic surfaces, the micelles have a lower curvature, and the adsorbed
layer is often flat. The driving force for a flat layer is to minimize the area of contact
between the hydrophobic surface and water.
2. When the mineral and the surfactant have an opposite charge, the micelle often has
a lower curvature than in bulk, because the bottom half of the micelle is pulled
down onto the mineral by electrostatic interactions.
3. On graphite, the alkyl chains of the surfactant can adsorb epitaxially. This epitaxial
layer templates the adsorption of hemicylindrical micelles for most surfactants.
4. The addition of salt can cause an increase in the size of micelles, as it does in bulk.
However, when the mineral and the surfactant have opposite charge, the addition of
salt weakens the influence of the mineral (item 2) and causes a decrease in micelle
size.
Recently, experiments on surfaces with charges that are fixed in place have shown that
the lateral mobility of surface ions is important to form discrete micelles of charged surfactants at interfaces (Ducker, Senden, and Pashley 1991). If the counterions to the surfactant
cannot move around, it is more difficult to compensate the micelle charge, so the surfactant
does not cluster as easily into a micelle. Mobile counterions can be either as salt added to
solution or surface counterions, such as H+, which allow the effective moment of an anionic
lattice charge.
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ADSORPTION OF SURFACTANTS
213
EA (attach)
Separation, S
Energy, V
EA (detach)
FIGURE 34 The probability of particle attachment to a bubble and the probability of
detachment depend on the surface forces that occur between the particle and bubble
Atomic Force Measurement
An “AFM image” is a collection of force measurements as a function of position in the X–Y
plane. An “AFM force measurement” is the extent of the force as a function of displacement
in the Z dimension (normal to the mineral surface). This force measurement can be integrated to obtain the energy profile that is central to understanding the interaction between a
particle and a bubble (Figure 34). For AFM imaging, a sharp tip is required to obtain maximum lateral resolution. Often, high lateral resolution is not required in AFM force measurement, so a large-radius (~5 μm) probe of known geometry (a sphere) is used (Ducker,
Senden, and Pashley 1991). The large radius increases the force (and, therefore, the signalto-noise ratio) and allows the use of the Derjaguin approximation (Derjaguin 1934) to
obtain an intensive interaction property: the energy per unit area between flat surfaces.
AFM has been used to study the forces relevant to flotation. One approach is to directly
measure the force between a particle and a bubble. This was first performed by Ducker, Xu,
and Israelachvili (1994) for a silica particle. Unfortunately, this approach is complicated by
the fact that the deformation of the bubble during the interaction makes it difficult to determine the separation between the particle and the bubble. This deformation has been examined in more detail by Dagastine and White (2002). Ducker, Xu, and Israelachvili (1994)
found that the addition of the surfactant, SDS, to solution caused a repulsive force between
a silica particle and an air bubble. The SDS does not adsorb to the particle but does adsorb
to the air–water interface, creating a repulsive double-layer force.
An alternative approach is to simulate the particle–bubble interaction with a particle–
hydrophobic solid interaction. The hydrophobic solid is designed to mimic the air–water
interface but does not deform. Yoon’s research group has measured the forces between silica
surfaces that were made hydrophobic by the irreversible attachment of silane agents
(Rabinovich and Yoon 1994). They measured a long-range attraction between a hydrophobic
particle and a hydrophobic solid. This force should lower the activation energy to particle–
bubble coalescence, and is described in more detail elsewhere in this volume. Ducker’s group
has measured the forces between silica surfaces that were made hydrophobic by the reversible
adsorption of dodecyltrimethylammonium bromide (C12TABr) (Subramanian and Ducker
2001; Lokar and Ducker 2004). At low concentrations, the addition of surfactant decreases
the repulsive force that is experienced by silica particles (Figure 35). At about CMC/10, the
force is purely attractive and much greater in magnitude than expected for the van der Waals
force between silica particles. The origins of the attractive force remain controversial. At
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214
FLOTATION FUNDAMENTALS
0.6
0.0
4
[C12TABr]/mmol·Kg
1.8
3.6
6.1
9.1
13
18
27
46
64
3
2
a
–2
0.2
a
E (mJ·m )
0.4
5
4
–2
[C12TABr]/mmol·Kg
0.0071
0.014
0.021
0.036
0.071
0.11
0.18
0.36
0.71
1.4
E (mJ·m )
0.8
4
1
–0.2
0
–0.4
–0.6
–1
0
10
20
30
40
(S-S0)/nm
50
60
0
5
10
15
20
25
(S-S0)/nm
Source: Subramanian and Ducker 2001.
NOTES: When there is little surfactant in solution, the force is repulsive because of long-range double-layer forces. As the concentration increases, adsorption of surfactant decreases the net charge on the surface, leading to loss of the repulsion. However, at about 1 mM concentration, the force is much more attractive than van der Waals force between silica particles,
suggesting that another attractive force is present. When the surfactant concentration is increased further, the force becomes
more repulsive. This is the result of both the buildup of a large positive charge and the energy required to push the surfactant
off the surface at short range.
FIGURE 35
solution
Forces between silica surfaces as a function of the concentration of C12TABr in
even greater surfactant concentrations, the force becomes less attractive and ultimately
repulsive again. The latter is due to the generation of a large charge on both particles.
Small Angle Neutron Scattering
Small-angle neutron-scattering (SANS) experiments generally follow this procedure:
1. Irradiate a sample with a neutron beam.
2. Measure the resulting scattering pattern.
3. Determine the structure that caused the observed pattern.
Scattering patterns are caused by the interference of secondary waves that are emitted
from various nuclei structures when irradiated. Scattering of neutrons is caused by differences in scattering power of different nuclei. SANS is used to explore microstructure as well
as particle size distribution of the colloidal length scale, because the larger the diffraction
angle, the smaller the length of scale probed.
SANS has been applied to determine the stability of colloidal suspension (Wong,
Cabane, and Somasundaran 1988). Addition of inorganic electrolyte calcium chloride
(CaCl2) to negatively charged silica spheres results in fractal coagulation, which is a reversed
process in contrast to flotation and is featured with no preferential distance between particles. Unlike the fractal aggregates (diffusion controlled, cluster–cluster aggregates), the
aggregation of silica spheres has exhibited both fractal and nonfractal behaviors in the presence of copolymer of acrylamide and trimethylaminoethyl acrylate at different mole fractions (Wong, Cabane, and Somasundaran 1988; Wong et al. 1989). For example, the local
distance between particles with copolymer containing 5% of cationic acrylamide is determined on the order of 1 to 2 particle diameters, indicating that the polymer constructs a
bridge between silica spheres. On the contrary, an increase in the mole fraction of cationic
acrylamide (30%) produces a totally different behavior. Figure 36 shows the scattering spectra
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ADSORPTION OF SURFACTANTS
215
Log I (Q), arbitrary units
Polyacrylamide 1.7 × 106
5% Cationic Monomer
6
Polyacrylamide 1 × 10
30% Cationic Monomer
–3
–3
–3
Log Q
Source: Wong et al. 1989.
FIGURE 36 Scattering from two classes of structures found in aggregates of silica (380 Å
diameter) flocculated with organic polymers
of silica spheres with these two polymers, where two peaks are observed for 5% acrylamide
and only one peak is present in the 30% acrylamide system. In the former case, the peak at
large Q is identified as the interparticle peak, as seen in free spheres, whereas the peak at low
Q results from interferences between objects at 520 Å apart, which is more than 1 particle
diameters (410 Å) but less than 2. The uniform short-range order produces interferences
resulting in a peak at log Q = –1.55. Heterogeneous fractal structures at distances much
larger than the sphere diameter are characterized in the latter case.
On the other hand, the structures of aggregates formed by silica spheres adsorbed by
surfactants have been surveyed using SANS. Effects of concentrations, structures of surfactant molecules, as well as the mixing ratios in binary systems have been identified. A significant and systematic increase in the apparent fractal dimension is noted with the increase in
the concentration of cationic single-chain alkyltrimethylammonium bromide surfactants.
With a double-chain cationic surfactant, reordering to a liquid-like structure and redispersion are observed. The pattern of the scattering spectra does not change as a function of the
hydrophobic chain length, as shown in Figure 37. This indicates that the property-governing structure is not of hydrophobic origin. Charge neutralization is considered as the main
feature of adsorbed ions.
In the neutron scattering study of nonionic surfactant mixtures—nonyl phenol, ethoxylated decyl ether (NP-10) and n-dodecyl-β-D-maltoside (DM)—the nanostructure of
adsorbed layer is attributed to the packing of molecules of different structures (L. Zhang and
P. Somasundaran, unpublished data). The scattering spectra on silica spheres are recorded in
Figure 38 under the condition that the total concentration is the same, saturation adsorption is reached, and the composition of the active component NP-10 varies. The curves are
in similar pattern for the surface aggregates above 25% NP-10. Interestingly, the fit using the
polydispersed core-shell model accounts well for the scattering up to Q at 0.055 Å–1. Then,
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216
FLOTATION FUNDAMENTALS
0.00
CxTABr 2 mM
x = 18
x = 14
x = 12
x = 10
x =8
Arbitrary Log Intensity
–1.00
–2.00
No Surfactant
–3.00
–4.00
–5.00
–6.00
–3.60
–2.80
–2.00
–1.20
–0.40
Log Q
Source: Wong et al. 1989.
FIGURE 37 Scattering spectra of silica (diameter 380 Å) aggregates made with
trimethylammonium bromide alkanes, 2 × 10–3 M, with number of tail carbons x shown for
comparison
25%
50%
75%
100%
10
I(Q), cm –1
1
0.1
0.01
5
6
0.01
2
3
4
5
6
0.1
2
3
4
Q, Å –1
Source: L. Zhang and P. Somasundaran, unpublished data.
FIGURE 38 Scattering spectra for the mixed-surface aggregates at NP-10 at the similar
volume fraction of total surfactant concentrations on silica surfaces in H2O (0.39)/deuterium
oxide (0.61) solvent at a temperature of 25°C at pH 7.4
a damped decrease is followed by a shoulder at Q around 0.08 Å–1 and a Porod tail at high
Q. On the basis of the discrepancy between the predicted and measured intensity, the bi-layer
structure is ruled out. The surface aggregates are in the shape of flattened beads.
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ADSORPTION OF SURFACTANTS
217
Analytical Ultracentrifugation
Analytical ultracentrifugation is used to quantitatively determine speciation in surfactant
mixtures and surfactant–polymer mixtures in solutions (Acar and Somasundaran 1989).
This technique is nondestructive and is particularly powerful for distinguishing the size and
shape of various species in mixtures. Recent results have revealed coexistence of two types of
micelles in polyethylene oxide solutions and its mixtures with sugar-based surfactant,
whereas only one micellar species is present in sugar-based surfactant solutions (Zhang and
Somasundaran 2004). Also, unlike ionic surfactants, the micellar growth of the nonionic
sugar-based and polyethylene oxide surfactants is found to occur at a concentration immediately above the CMC. Both dynamic and equilibrium characteristics of surfactant mixtures,
polymer–surfactant mixtures, can be obtained using this technique.
Nonlinear Optical Technique: Second Harmonic Generation
A powerful nonlinear optical technique that is useful for studying surfactant adsorption
and, particularly, their orientation on solids involves second harmonic generation (SHG;
Wang 1996).
SHG is a second-order nonlinear process that is useful for the study of surfaces and
adsorbed monolayers. This technique not only possesses the versatile accessibility and the
chemical selectivity of optical spectroscopic techniques, but also overcomes the poor selectivity between interfacial and bulk molecules of conventional linear optical techniques, such
as Fourier transfer infrared spectroscopy. Furthermore, because the frequency of the outgoing optical signal lies in a different spectral region from that of the incident light, the latter
can be easily discriminated from the desired optical signal.
In the dipole approximation, SHG is forbidden in the bulk of a medium because of lack
of asymmetry, whereas at the surface inversion, symmetry is broken and SHG is allowed. In
an SHG process, two photons at the same frequency ω simultaneously interact with a nonlinear medium to generate radiation at a frequency of 2ω. The typical experimental geometry for SHG is shown in Figure 39.
E(ω)
E(2ω)
k1(ω)
Bulk 1
Interface
θi
θr
Bulk 2
k2(ω)
ω) is the incident
FIGURE 39 Typical experimental geometry for SHG at an interface where E(ω
ω) is the radiation generated from the nonlinear polarization. θi and θr are
electric field and E(2ω
ω) and k2(ω
ω) are the incident and refractive beam,
the incident and refractive angle; k1(ω
respectively.
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218
FLOTATION FUNDAMENTALS
Adsorption
B
A
A
B
Concentration
FIGURE 40 Schematic illustration of surfactant monolayers at low (a) and high (b) surfactant
concentration
Surfactants play a key role in many processes such as flotation, as well as detergency, oil
recovery, lubrication, painting, and cosmetics. Chain orientation and conformation of
adsorbed surfactant species at interfaces will play a major role in determining the final outcome. The SHG signal depends on the asymmetry of the interface, which is related to orientation and conformation of surfactant molecules.
Surfactant molecules at interface are more sparsely distributed at low adsorption density and associate to form colloids at higher densities. The SHG signal can reveal the average
orientation of a symmetric stretch transition dipole component perpendicular to the surface
and the number of gauche defects in the alkyl chain affects the intensity of signal, as large
numbers of gauche defects cause random alkyl-chain orientation. With the increase in surface concentration, a reduction in gauche defects takes place because of chain–chain interaction or self-assembling shown schematically in Figure 40. This can be expected to affect the
SHG signal yielding in-situ unique information and conformation of adsorbed species on
minerals and bubbles.
H Y D R O DY N A M I C S O F M I N E R A L F L O TAT I O N
In addition to the physical chemistry of the interfaces, the collision between particles and
bubbles in the liquid suspension, attachment of the particles to the bubbles, and levitation
of the particle–bubble aggregates to the surface of the suspension assume vital roles in flotation. The collision, attachment, and stability of the aggregates depend on the hydrodynamics in the flotation cell. Thus, flotation is a dynamic process that involves many
hydrodynamic phenomena in a system that contains solid, liquid, and gas in a state of varying turbulence.
Film Rupture Between Minerals and Bubbles
Rupture of thin liquid films between liquids, gases, and solids plays a governing role in
determining the stability of bubbles and droplets in foams, emulsions, froths, and even biological cells. Behavior of liquid films on solids is particularly relevant to many processes such
as detergency and froth flotation of minerals. Whereas much work has been accomplished
on the rupture of liquid films and aqueous films on solids, very little has been done on the
films of surfactant solutions on solid minerals, even though the rupture of these films is the
critical step that determines the efficiency of processes such as flotation.
The thinning and rupture of aqueous surfactant films on silica (Somasundaran et al.
1999) was investigated using the interferometric technique and free bubbles, which proximate
flotation more realistically than the captive bubble method. Different rupture mechanisms
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ADSORPTION OF SURFACTANTS
219
are observed at diverse pH values for aqueous films of dodecyl amine on silica, explaining
the pH dependence of quartz flotation using amine. At low pH, rupture is accompanied by
the breaking off of a large drop in the center and subsequent formation of large irregular
drops; at high pH, ruptured spots grow to large circular drops.
Particle–Bubble Interactions in Solutions
Attachment of particles to bubbles involves various interactions resulting from electrical
double-layer forces between the particle and the bubble (with adsorbed surfactant on both),
van der Waals forces, the energy change due to the transfer of the hydrocarbon chains
adsorbed on the particles to the gaseous phase, and the steric repulsion between adsorbed
surfactant layers on the two interacting surfaces. The energies of these interactions for the
alumina–dodecyl sulfonate system are calculated using data for the zeta potential of the
mineral, the surface tension of the solutions, and the surfactant adsorption density. Estimated
total interaction energy is correlated with the results of the bubble pickup experiments.
Hydrophobic Interactions. The free energy, VH, involved in the transfer of dodecyl
chains of the surfactant adsorbed on the solid surface to the gaseous phase upon attachment
of the bubble to the particle can be estimated as
S ⁄ L – L ⁄ G × 12A
V H = Γ S ⁄ L Φ CH
C
2
(EQ 18)
S ⁄ L – L ⁄ G is the transfer
Γ S ⁄ L is adsorption density at the solid–liquid interface, Φ CH
2
energy per molecule of CH2 groups from the solid–liquid to liquid–gas interface, and AC is
the area of contact between bubble and particle.
Electrical Double-Layer Interactions
Expressions for the interaction energy of overlap of double layers differ essentially in the
choice of boundary conditions, namely, constant charge or constant potential of the interacting surfaces during their neutral approach. For the case of constant potential at both surfaces, the following equation derived by Hogg, Healy, and Fuerstenau (1966) is appropriate:
V E1 – Ψ
1 + exp ( – κH 0 )
⎧ 2ψ 1 ψ 2
⎫
- ln ( 1 – exp ( – κH 0 ) ) +⎪
- ln -----------------------------------εa 1 a 2 ( ψ 12 + ψ 22 ) ⎪ -----------------------2
2
= -------------------------------------- ⎨ ( ψ 1 + ψ 2 ) 1 – exp ( – κH 0 )
⎬
4 ( a1 + a2 ) ⎪
⎪
1
κH
ln
(
–
exp
(
–
)
)
⎩
⎭
0
(EQ 19)
where V E1 – ψ is the electrostatic energy of interaction under constant potential (ψ) conditions at two interfaces; ε is the dielectric constant; a1 and a2 are radii of particle 1 and 2,
respectively; ψ1 and ψ2 are electrostatic potential of particle 1 and 2, respectively; κ–1 is the
double layer thickness; and H0 is the distance between the two interacting surfaces.
On the other hand, if the charge densities at both interfaces are assumed to remain constant, the equation derived by Weise and Healy (1970) using the procedure suggested by
Frens and Overbeek (1972) applies:
εa 1 a 2 ( ψ 12 + ψ 22 )
- ln ( 1 – exp ( – κH 0 ) )
V Eσ – σ = V EΨ – Ψ – ------------------------------------4 ( a1 + a2 )
(EQ 20)
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220
FLOTATION FUNDAMENTALS
where V Eσ – σ is the electrostatic energy of interaction under constant charge conditions at
two interfaces. The use of neither the constant potential nor the constant charge condition
will be valid, and an intermediate condition between these two extremes will be more appropriate. For the present treatment, the value of the double-layer interaction energy is taken as
the arithmetic average of the values calculated using Equations 22 and 23.
Van der Waals Interactions
Because the present system is comprised of three different media (solid, air, and intervening
solution), an energy change will be associated with van der Waals forces when the bubble
and particle are brought together. For the case of a flat plate and a sphere (ap << ab), this
free-energy change is given by
V V = – Aa p ⁄ 6H 0
(EQ 21)
where Vv is the van der Waals energy of interaction; A is the Hamaker constant for the system; ap and ab are the radii of the particle and the bubble, respectively; and H0 is the distance of separation. The Hamaker constant for the system is given by
⎛
⎞⎛
⎞
A = ⎝ A bb – A ll ⎠ ⎝ A pp – A ll ⎠
(EQ 22)
where b, l, and p represent bubble (gas), liquid, and particle, respectively.
Interaction Due to Steric Repulsion
When the interacting spheres (bubble and particle, in this case) contain adsorbed layers,
their adhesion will be sterically hindered, essentially because of the physical size of the molecules. Therefore, the treatment of the aggregation phenomenon has to be modified to
include a volume-restriction term in the expression for overall interaction energy. Mackor
(1951) has estimated this interaction energy by considering the loss in entropy due to the
restriction of the molecules during the attachment. The expression derived by Mackor is
V s = θN ∞ RT ( 1 – H 0 ⁄ δ )
(EQ 23)
where Vs is the steric energy of interaction, N∞ is the maximum number of sites available for
adsorption, R is gas constant, T is the absolute temperature, θ is the fractional surface coverage, and δ is the adsorbed layer thickness. The steric repulsion has therefore been calculated
using the following expression with δ = 20 Å.
V s = Γ S ⁄ L RT ( 1 – H 0 ⁄ 20 ) × A C
(EQ 24)
The overall interaction energy is the sum of contributions from the attractive hydrophobic forces resulting from the partial transfer of dodecyl sulfonate chains adsorbed at the
solid–liquid interface to the liquid–air interface; and the net repulsive forces (throughout
most of concentration range studied) resulting from the van der Waals interaction, the overlap of the electrical double layers of the particle and the bubble, and the steric repulsion
between the surfactant layer adsorbed on the two surfaces.
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ADSORPTION OF SURFACTANTS
221
Particle–Particle Interactions in Solutions
Aggregation can be considered to be caused either by Brownian motion or by external forces
such as stirring. In other words, the energy barrier resulting from the overlap of double layers
can be overcome by either the thermal energy or, as in the present case, the hydrodynamic
energy. Although aggregation due to Brownian motion—called “perikinetic aggregation”—
has been at least qualitatively dealt with by the Derjaguin–Landau–Verwey–Overbeek
(DLVO) theory, the orthokinetic aggregation of stirred mineral systems is not satisfactorily
explained by it. Major assumptions limiting the application of DLVO theory to such systems include those of thermal equilibrium between interacting double layers, low pulp density (dilute suspensions), and insignificant stirring effect. An attempt has been made to
extend the application of the theory to a system where orthokinetic aggregation of highly
charged anatase and calcite particles are present in turbulent flow. In addition, hydrophobic
layers present on the interacting particles are expected to influence the aggregation process.
The hydrophobic interaction of adsorbed oleate has also been taken into account when
determining the criteria for selective aggregation in carrier flotation.
The energetic criteria for aggregation of particles in the carrier flotation of anatase fines
using coarse calcite particles are developed. Flotation and aggregation test results are correlated with predictions based on the criteria. The role of hydrophobic interactions in governing the aggregation and that of selective aggregation in turn in carrier flotation are discussed
in the following section.
Repulsive Energy of Interaction
The case of constant surface potential interaction was treated in 1966 by Hogg, Healy, and
Fuerstenau and the following simplified solution was obtained for energy of interaction
between two dissimilar spheres, VR , for conditions of low surface potential (ψ0 < 25 mV),
and thickness of the double layer as small compared to the particle size ( κα ≤ 10 ) for dissimilar spheres:
2 + ψ 2 ) 2ψ ψ
εa 1 a 2 ( ψ 01
1 + exp ( – κH 0 )
02
01 02
----------------------- ln ------------------------------------ + ln ( 1 – exp ( – 2κH 0 ) )
V RΨ0 = -----------------------------------------2
2
4 ( a1 + a2 )
ψ 01 + ψ 02 1 – exp ( – κH 0 )
(EQ 25)
where ε is the dielectric constant (in SI units), a1 and a2 are the radii of respective spherical
particles, ψ01 and ψ02 are their surface potentials, and H0 is the minimum separation of particle surfaces.
In the case of constant surface charge interaction, interaction between dissimilar spherical particles under constant surface charge conditions was treated by Wiese and Healy
(1970) to yield the following expression for repulsive energy V Rσ for dissimilar spheres:
2 + ψ2 )
εa 1 a 2 ( ψ 01
02
V Rσ = V RΨ0 – -----------------------------------------[ln ( 1 – exp ( – 2κH 0 ) ) ]
2 ( a1 + a2 )
(EQ 26)
where V Rσ and V RΨ0 represent the repulsive energy at constant charge and at constant
potential, respectively.
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222
FLOTATION FUNDAMENTALS
Attractive Energy of Interaction
The potential energy of interaction between dissimilar spheres at small H0 (i.e., H0 << a)
was derived by Hamaker (1937) for dissimilar spheres as
Aa 1 a 2
V A = – -----------------------------6 ( a 1 + a 2 )H 0
(EQ 27)
Under most practical mixing conditions, the flow is turbulent and its velocity gradient
is not constant.
Assuming homogeneous isotropic turbulence, two different mechanisms were considered depending on whether the particle size is smaller or larger than microscale of turbulence, λ0:
ν3
λ 0 = ⎛ -----⎞
⎝ ε⎠
(EQ 28)
where ν = n/ρ is the kinematic viscosity and ε is the average rate of energy dissipation per
unit mass of suspension.
For conditions of particle size much smaller than the microscale of turbulence, the relative velocity, U, at the moment of collision is related to radius of particles a1 and a2 (Harris
and Mensah-Biney 1977) as shown in the following equation:
ε 1⁄2
U = 0.206 ⎛ ---⎞ ( a 1 + a 2 )
⎝ ν⎠
(EQ 29)
and the rate of aggregation ( J) in the absence of a barrier is given by
ε
J = 1.29 ⎛ ---⎞
⎝ ν⎠
1⁄2
( a1 + a2 )3 n1 n2
(EQ 30)
where n1 and n2 are the concentrations of particles of radii at a1 and a2, respectively.
If particle size is larger than microscale of turbulence λ0, an inertia mechanism is applicable. The relative velocity of particles whose centers are separated by (a1 + a2) is given by
Delichatsios and Probstein (1975):
U = 1.37ε 1 ⁄ 3 ( a 1 + a 2 ) 1 ⁄ 3
(EQ 31)
The aggregation rate under this condition is
J = 1.37ε 1 ⁄ 3 ( a 1 + a 2 ) 3 n 1 n 2
(EQ 32)
The results are summarized as follows:
1. Repulsive energies between various particle systems comprised of anatase, calcite,
and kaolinite were estimated for constant potential conditions on the basis of the
Debye–Huckel approximation, and attractive energies between them were calculated taking the retardation effects into consideration. The maximum total potential
energy of interaction was compared with the relative kinetic energy of the colliding
particles calculated from the estimated data for relative velocity and reduced mass.
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ADSORPTION OF SURFACTANTS
223
2. The analysis, based on DLVO theory and aggregation theory in turbulent flow, predicts aggregation between anatase and coarse calcite particles as was actually observed
in carrier flotation systems. The prediction for beaker–scale aggregation systems tests
using oleate was, however, in contradiction with the experimental findings. This discrepancy can be attributed to the anisotropic turbulent flow near the impeller.
3. Interaction of adsorbed oleate on particles that can be expected to contribute
toward the total attractive energy was estimated to be significant. Including the
hydrophobic interactions at the closest distances of approach, the total interaction
energy is attractive and is of the same order of magnitude as the relative kinetic
energy for colliding particles in the aggregation tests. Conversely, in the absence of
the hydrophobic contribution, the total energy is repulsive. Thus, aggregation
occurs under intense agitation conditions only for particles coated with oleate.
4. Selective anatase–calcite aggregation in carrier flotation is attributed to strong
adsorption of oleate on anatase and calcite and lack of such adsorption on kaolinite
at least under the intense mixing conditions used.
In addition to the hydrodynamic factors discussed previously, other hydrodynamic
parameters, such as pulp aeration, agitation intensity, bubble size, and particle size, also have
an effect on the flotation process.
AC K N OW L E D G M E N T S
The authors acknowledge R. Zhang and S. Lu for their contributions and thank the
National Science Foundation, the Department of Energy, and the NSF Industry/University
Cooperative Research Program (Grant No. EEC-03-28614) for their support of this work.
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Pulp and Solution Chemistry
John Ralston, Daniel Fornasiero, and Stephen Grano
A B S T R AC T
Pulp and solution chemistry play a pivotal role in bubble–particle capture. Surface heterogeneity, degree of surface hydrophobicity, degree of oxidation, bubble nucleation, and related phenomena influence the stability of the aqueous wetting film between a particle and an
approaching bubble. For metal sulfide minerals, which oxidize relatively easily in many cases,
monitoring and control of oxidation is often the key to selective flotation. Surface spectroscopy,
electrical double-layer studies, and cyclic voltammetry together enable various mechanisms to be
understood. The application of this information to several sulfide ore operations is discussed,
illustrating how rather straightforward tests can be used to diagnose and correct problems in
operating plants. Oxidation product dissolution and re-precipitation, as well as metal ion
hydrolysis formation and adsorption, are central issues. The use of pulp and solution chemistry
as diagnostic tools in selective flotation is examined in several case studies.
B A S I C P R I N C I P L E S O F B U B B L E – PA R T I C L E C A P T U R E
The efficient capture of specific particles by rising bubbles and the delicate control of interfacial chemistry that leads to selective capture are central issues in froth flotation. For efficient capture to occur between a particle and a bubble, they must undergo a close encounter,
a collision process dominated by hydrodynamics (Figure 1). Should this encounter bring the
particle and bubble within the range of attractive forces, the intervening liquid film between
the bubble and the particle will drain, leading to a critical thickness at which rupture occurs.
The liquid–vapor interface then retreats over the solid particle surface until a stable wetting
perimeter is established. The particle and bubble are then joined after this attachment process. The particle may only be dislodged from this state if it is supplied with sufficient
kinetic energy to exceed the detachment energy in this detachment process. These various
events are now well documented and may be followed using high-speed videomicroscopy
(Schulze 1984; Nguyen, Evans, and Jameson 2001). Imaging these processes under turbulent conditions at high pulp densities still presents a real challenge, however, with fast tomographic methods as one option.
The process of collision can now be described quantitatively for single bubbles colliding
with bubbles under various flow regimes (Dai et al. 1998) and appropriate scale-up is possible for multiple bubble particle collisions occurring under turbulent conditions (Pyke, Fornasiero, and Ralston 2003; Duan, Fornasiero, and Ralston 2003). This chapter addresses the
role of pulp and water chemistry and how they influence the flotation process. Basic principles and concepts are discussed, with later focus on oxidation processes for metal sulfides.
Table 1 summarizes key publications relating to fundamental aspects of pulp and water
chemistry in flotation from the mid-20th century to the present time.
It is important to first recognize that the forces that control the interaction between a
bubble and a particle are mainly repulsive (e.g., Ralston 2000). This is especially true of the
227
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228
FLOTATION FUNDAMENTALS
Particle to Bubble Center-to-Center
Relative Distance
1.24
Particle Movement
into the Bubble Surface
1.19
Movement of Three-Phase
Contact Line
1.14
Rupture
1.09
Collision
Sliding
1.04
0
20
40
60
80
100
Time, msec
Adapted from Schulze 1984; Nguyen, Evans, and Jameson 2001.
FIGURE 1
Bubble–particle distance as a function of time
van der Waals force of interaction for an asymmetric system of a bubble interacting with a
particle through an aqueous film: the van der Waals force or energy is repulsive and only
becomes attractive when the refractive index of the solid phase is less than water (Israelachvili 1991). Given that the electrostatic forces are generally repulsive, the primary reason for
attachment between a particle and a bubble in mineral flotation is due to the existence of an
attractive hydrophobic force. This attraction is caused by the addition of collectors that
adsorb at the appropriate solid–solution interface and/or by subtle changes in metal-to-sulfur
ratios in the early stages of oxidation of sulfide mineral surfaces (Fuerstenau, Miller, and
Kuhn 1985; Prestidge and Ralston 1995). If oxidation of a metal sulfide surface progresses
too far, so that hydrophilic oxidation products form or hydrophilic species adhere to a
hydrophobic surface as a result of the interaction with or between other minerals—either
from speciation processes in process waters or from added reagents or organic impurities in
the water—then the degree of hydrophobicity and the attachment affinity between a particle and a bubble will decrease.
The “neglected interface,” which is the air–water interface where frothers, collectors,
organic depressants, and organic impurities can congregate (Leja 1982), should not be forgotten; their presence at this interface can establish barriers to the particle–bubble attachment process (Tjus et al. 1988–1989). Hence, the correct balance for the forces of attraction
between a bubble and a particle must be obtained, otherwise true flotation will not occur.
Because flotation is a kinetic process, the flotation rate constant for the value material
should exceed that of the gangue by as large an amount as possible. The rate constant is very
much influenced by the speed of the subprocesses that take place during attachment and
detachment (Schulze 1984; Crawford and Ralston 1988). To understand the role of interfacial chemistry in more detail, the elements of thermodynamics, surface heterogeneity, bubble formation, surface force measurements, gas solubility, and spectroscopy must be
involved. It is critical to first understand the behavior of thin wetting films on solid surfaces.
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229
TABLE 1 Key publications for understanding fundamental aspects of flotation from mid-20th
century to present
Date
1948
Area of Research
A fundamental paper by Sutherland (1948) on the kinetics of the flotation process
appeared in Australia. This paper invoked induction time, described particle-size effects in
flotation, and catalyzed other similar approaches. Although it was preceded by other
efforts, this paper was the first comprehensive effort to describe recovery, size, and time
data in a fundamental manner.
1955
The Australian textbook, Principles of Flotation, by Sutherland and Wark (1955), made a
strong impact. Many aspects of pulp chemistry were thoroughly analyzed.
1960–1961 In Moscow, Derjaguin and Dukhin (1960–1961) produced a key paper on the theory of
flotation of small- and medium-sized particles. Hydrodynamics, surface forces, and
diffusiophoresis were all used in this theory. This seminal work resulted in an acceleration
of fundamental flotation research worldwide.
1972
Blake and Kitchener (1972), working together in London, published some very careful
measurements of the thickness of aqueous films on hydrophobic quartz surfaces. Film
thicknesses, measured as a function of salt concentration, were shown to depend on the
electrical double-layer force. Film instability occurred on hydrophobic surfaces at film
thicknesses less than about 60 nm. This value, which was larger than the range of the
electrical double-layer force, represented the combined effects of hydrophobic force,
surface heterogeneities, and external disturbances. Blake and Kitchener’s film thickness
studies hinted at the length dependence of hydrophobic forces, information that was
subsequently obtained by surface force experiments after 1982.
1976
In Bulgaria, Scheludko and colleagues (Scheludko, Toshev, and Bojadjiev 1976)
considered how particles might become attached to a liquid surface and developed the
capillary theory of flotation, which complemented the Russian efforts.
1977
In London, Anfruns and Kitchener (1977) published the first measurements of the absolute
rate of capture of small particles in flotation. This was the first critical test of collision theory
under conditions where the bubble and particle surface chemistry was characterized and
controlled.
1983
In Freiberg, Schulze (1983) published a key textbook on the physicochemical substeps that
are important in flotation, drawing on a wide range of hydrodynamic, surface chemical, and
engineering information. Although originally published in German, after being translated
into English, the book captured an international audience.
circa 1970– Major progress has been made toward understanding the electrochemistry of metal
present
sulfides (Woods 1972) and using surface analysis to characterize mineral surfaces and
interfacial chemistry to explain flotation behavior (Buckley, Woods, and Wouterlood 1989).
There has been a strong interest in developing reliable collision models (Dai et al. 1998).
The surface force apparatus and, recently, the atomic force microscope colloid probe
technique have provided very useful insight into electrical double-layer, van der Waals, and
hydrophobic forces (Israelachvili 1991; Fielden, Hayes, and Ralston 1996). Thin-film
drainage has been investigated between a rigid and a deformable interface (Miklavcic,
Horn, and Bachmann 1995). Attachment efficiencies have been measured (Hewitt,
Fornasiero, and Ralston 1995). Reliable methods for measuring contact angles on particles
have been developed (Diggins, Fokkink, and Ralston 1990). Major theoretical and
experimental advances in describing dynamic contact angles on well-defined surfaces
have taken place (Blake 1993). The prediction of flotation rate constants for single minerals
and valuable minerals in ores has now been determined (Pyke, Fornasiero, and Ralston
2003; Duan, Fornasiero, and Ralston 2003) and is being extended to flotation plants
(Ralston, Fornasiero et al. 2005).
A Q U E O U S W E T T I N G F I L M S O N S O L I D S U R FA C E S
Studies of wetting films on solid surfaces have frequently considered silica surfaces (Read
and Kitchener 1969), which give a very good understanding of solid surfaces in general.
Smooth surfaces can be readily prepared, electrical double-layer and optical data are available, and the optical transparency of silica facilitates measurements. The surface silanol
groups that are present on a silica (or quartz) surface and readily observed by infrared spectroscopy (Tripp and Hair 1991) make the surface strongly hydrophilic. Therefore, a high
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FLOTATION FUNDAMENTALS
t = 0 msec
t = 0.97 msec
t = 1.95 msec
t = 2.92 msec
t = 3.90 msec
t = 4.87 msec
t = 5.85 msec
t = 6.82 msec
Source: Schulze 2002, reproduced with permission from Snap Printing, Adelaide, Australia.
FIGURE 2 High-speed-video sequence of the rupture of an aqueous wetting film on methylated
glass with a contact angle of 59°. Film diameter: 200 μm. The rupture occurs at 0.97 msec.
heat of wetting, a zero contact angle for clean silica, and a water vapor adsorption isotherm
that rises asymptotically to the axis at a pressure equal to the saturation value are significant
factors. This is strong evidence for a high work of adhesion. Thick wetting films are perfectly stable on clean, hydrophilic silica surfaces. In this case, a high negative potential at the
silica–pure water or silica–dilute electrolyte solution interface and a clean air–water interface are the result. The equilibrium thickness of these films is dependent on ionic strength,
pH, and hydrostatic pressure (Blake and Kitchener 1972). The thickness of such a film may
be calculated (Hewitt et al. 1993), assuming that the interaction between the silica and the
air bubble occurs at constant potential. If the thickness is known, then the potential at the
air–solution interface can be obtained. It is generally much smaller than indicated by extant,
flawed* determinations of the bubble zeta potential, which are obtained through electrophoretic mobility measurements. The problem exists wherein gas–aqueous solution surface
potential data are very limited and no convincing mechanism has yet been identified for
charging mechanisms at any gas–solution interface.
If the silica surface is heated to 1,000°C or so, the surface silanol (–SiOH) groups are
mainly eliminated by dehydroxylation reactions of the following form (Iler 1979):
2 ( – S iOH ) → – SiOSi + H 2 O
The siloxane (SiOSi) groups that are formed are comparable to an ether in nature—even
though they are weakly dipolar, they are less hydrophilic than the silanol groups. Consequently, dehydroxylated silica shows an advancing contact angle against water of approximately 40°C, intercepts the P0 axis, and, below saturation pressure, results in far less
adsorbed water than does hydrophilic quartz. Similar behavior occurs for silica that has been
strongly methylated by reaction with a reagent such as trimethylchlorosilane, although the
effects are more pronounced. It is only when these surfaces adopt some hydrophobic character that they show any inclination to collapse. When collapse occurs on these surfaces, the
* The conversion of electrophoretic mobility data to zeta potentials in the case of bubbles is
flawed because the bubble surface polarizes during measurement; hence, existing theory, which
does not account for this fact, gives erroneous zeta potential data (Kelsall et al. 1996a, b).
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PULP AND SOLUTION CHEMISTRY
231
t = 0 sec
t = 1 sec
t = 2 sec
t = 3 sec
t = 4 sec
t = 5 sec
t = 6 sec after redraw of the bubble
Source: Schulze 2002, reproduced with permission from Snap Printing, Adelaide, Australia.
FIGURE 3 Ruptured wetting film of electrolyte solution with 10–2 M KCI (potassium chloride)
and 10–4 M AlCl3 (aluminum chloride) on oppositely charged glass surface. Pinning of the
three-phase contact line on multiple holes occurs during bubble removal, whose distance
indicates the wavelength of the critical fluctuation.
negative zeta potential remains virtually unchanged; the electrostatic interaction between
bubble and solid is essentially unaltered by the introduction of surface hydrophobic groups
(Grieser et al. 1984). The behavior of such an unstable film is shown in Figure 2.
A key feature in this high-speed-video sequence (Figure 2) is that only one embryonic
hole is sufficient to cause destabilization and dewetting of the entire film (Schulze, Stockelhuber, and Wenger 2001; Stockelhuber, Schulze, and Wenger 2001). The mechanism of
rupture is attributed to the presence of gas nuclei formed on heterogeneous surface sites
(Figure 3). The behavior of these wetting films on silica surfaces is mirrored for other solids.
Hydrophobicity cannot be explained simply by the Derjaguin-Landau-Verwey-Overbeek
(DLVO) theory. In this case, attraction will only occur if the solid has a refractive index less
than that of water (a rare occurrence) or if charge reversal occurs at the air–water interface
when a bubble approaches a solid of like but unequal potential. Although there is some evidence of this occurring for interacting metal sulfides (e.g., Toikka, Hayes, and Ralston
1997), the case for charge reversal for a bubble and a solid remains tenuous. Surface force
measurements performed between a captive bubble and a clean silica sphere, both with negative but different surface potentials, show evidence of a weak attraction (Fielden, Hayes,
and Ralston 1996), an observation subsequently made for mica and an air bubble (Pushkarova and Horn 2005). A stable wetting film exists at short range because of repulsive van
der Waals forces with an electrostatic attraction at intermediate range.
Heterocoagulation between a negatively charged bubble and a positively charged solid
surface has also been invoked as a reason for so-called “contactless” flotation (Derjaguin,
Dukhin, and Rulyov 1984). Experiments investigating the stability of a thin aqueous wetting film on a positively charged silica surface show that the film rupture may occur by a
mechanism described as spinodal dewetting, which occurs as a result of growing fluctuation
waves (Stockelhuber, Schulze, and Wenger 2001).
Although numerous attempts have been made to describe the drainage rate between a
particle and a bubble (or between a bubble and a plate), a quantitative description remains
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232
FLOTATION FUNDAMENTALS
elusive, as does the rate of retreat of a water–gas–solid contact line over a heterogeneous surface (Petrov, Ralston, and Hayes 1999). For this reason, induction time is still a very useful
parameter (Dai, Fornasiero, and Ralston 1999).
S U R FA C E H E T E R O G E N E I T Y
Two kinds of surface heterogeneities can be examined. The first is geometric and is actually
surface roughness; real solids are inevitably rough, with steps, kinks, recesses, and so on
(Lubetkin 1994). These geometric surface defects can influence bubble and particle interaction forces (e.g., by modifying the van der Waals interaction; Suresh and Walz 1996). Surface asperities can stimulate rupture of a thin liquid film between a particle and a bubble and
enhance flotation (Anfruns and Kitchener 1977). The movement of the three-phase contact
line over a rough surface following rupture of the thin film can be inhibited with the possibility that the line can be pinned at a heterogeneity (Semal et al. 1999). Surface defects can
also act as sites where gas can be trapped (Lubetkin 1994).
Sulfide mineral surfaces oxidize in a manner that ensures the process is centered on
impurity sites and/or surface defects. The inevitable consequence is that a chemically heterogeneous surface is produced, with parallel roughening on the surface of the nanometer and,
eventually, micrometer scale (Kim et al. 1995). Mild oxidation of most metal sulfides will
produce an intrinsically hydrophobic surface, due to the formation of a sulfur-rich surface
(Toikka, Hayes, and Ralston 1996). This hydrophobicity may be enhanced in the presence
of collector molecules, which adsorb in an uneven fashion on the solid surface. This patchy
distribution of collectors was first recognized by Plaksin (Plaksin and Shafeyev 1958) and
has been subsequently verified by modern scanned probe microscopy and surface analysis
techniques. This surface patchiness is also evident for nonsulfide surfaces, such as oxides and
silicates. Contact angle hysteresis on such surfaces is substantial and may significantly contribute to the ease or difficulty with which a particle may be detached from a bubble after
the two have formed a genuine contact (Schulze 1984).
Much can be learned about the interaction of water with patchy surfaces from water
adsorption isotherms and infrared spectroscopy studies (Zettlemoyer 1969; Gregg and Sing
1982; Griffiths and de Haseth 1986). On a hydrophobic surface, the first water molecules to
adsorb adhere to hydrophilic impurity sites, and the next adsorbate molecules cluster
around them. Low-energy adsorption is evident but with high entropy, pointing to the fact
that the adsorbed water molecules are highly mobile. This clustering process is very evident
in infrared studies of solids that act as good nucleants for ice formation (e.g., in applications
such as cloud seeding). Interacting surface hydroxyls show up very strongly for good nucleants, and isolated hydroxyls are far less populous; the reverse is true for poor nucleants. In
these instances, water molecules that are present in a continuous phase of air will adsorb on a
solid surface. This behavior inevitably leads to the question: How do patchy surfaces impact
the interaction between a particle that possesses such a surface and a bubble? In particular,
in this case where the continuous phase is water, what impact does the presence of dissolved
gas have and would nucleation have a role?
B U B B L E F O R M AT I O N O N S U R FA C E S
The emphasis here is on bubbles consisting of a gas in a continuous liquid phase and, in particular, for small bubbles present in the absence of surfactants. Whether or not bubbles
“cream” or can be treated as a classical colloidal system is principally determined by their
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PULP AND SOLUTION CHEMISTRY
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Peclet number (Leja 1982). In the latter case, this means that the upper limit of bubble
diameter is about 1 μm. Bubbles maintain sphericity up to diameters of about 1 mm,
beyond which their shape becomes rather complex and they may become axisymmetric
(Leja 1982). Theories exist for the heterogeneity and homonucleation of bubbles (Lubetkin
1994). Of particular importance here is the evidence relating to the influence of a surface on
the process of bubble formation, which is an important consideration for determining the
influence of the hydrophobic force in bubble–particle interaction.
In a series of papers, catalyzed by the very important problem of understanding and
controlling gas bubble formation in living organisms, Harvey and associates demonstrated
that the presence of nuclei in solution or on solid surfaces was a prerequisite for bubble formation (Harvey et al. 1944a, b, 1945; Harvey 1945; Harvey, Cooper, and Whitely 1946;
Harvey, McElroy, and Whitely 1947). Dissolved gas can exist in solution as a genuine
molecular solution in the form of tiny bubbles, purportedly down to nanometer dimensions, “bubstons,” (Bunkin et al. 1997), “gas bells” (Wrobel 1952), and as somewhat larger
bubbles. The formation of larger bubbles from preexisting sources of bubbles can occur
without a nucleation step (Wrobel 1952). Small bubbles with a small Peclet number (submicrometer in diameter), Harvey nuclei, and entrained or sparged bubbles can act as a source.
In the specific case of Harvey nuclei, the source can be any gas-filled re-entrant cavity, where
the trapped gas cannot be displaced by the surrounding liquid phase (Lubetkin 1989).
The size of the cavity must be larger than the critical bubble nucleus size under the
experimental conditions (Urban 1978). The trapped gas may be the same as or different
from the dissolved gas or liquid vapor. Any solid particle, flat surface, or container wall will
generally have been in contact with air, and air-filled cavities can then act as Harvey nuclei
for bubble formation in the system. If a solid surface is made hydrophobic, it can then act as
a greater source of Harvey nuclei, for the potential to trap gas in cavities is increased. Surface
geometry is important, as rough surfaces are a greater source of re-entrant cavities.
These tiny bubbles, existing as a dispersion or as Harvey nuclei resident on surfaces,
may be removed by high-speed centrifuging and subjection to high pressures (above 700
atmospheres or boiling followed by exposure to high pressures) (Harvey, McElroy, and
Whitely 1947; Urban 1978). By experimentation, gas nuclei have been found to be more
difficult to remove from a rough and irregular hydrophobic surface than from any other
type of surface (Lubetkin 1994; Wrobel 1952). These results are in accord with extant theory of bubble nucleation and detachment processes (Lubetkin 1989). Metal sulfide mineral
surfaces that had been treated with recrystallized potassium ethyl xanthate solutions and
then subjected to high pressures showed a reluctance to being attached to a captive air bubble, as compared with the same surfaces that had not been exposed to such pressures (Wrobel 1952). In the first case, induction times of 50 or more seconds were common, whereas in
the second instance, attachment was virtually instantaneous. Of course, the importance of
dissolved gas in enhancing mineral flotation and the “frosting” of surfaces with tiny bubbles
is well known (Klassen and Mokrousov 1983).
The key feature of these bubble formation investigations is the significant role performed by the solid surface. Surprisingly, Harvey nuclei, which are present on solid surfaces,
have often been overlooked by interfacial scientists who have been intent on measuring surface forces between solid hydrophobic surfaces. Very small bubbles present at the hydrophobic solid–water interface influence the attraction between two hydrophobic solid surfaces or
between a bubble and a particle (Dai, Fornasiero, and Ralston 1998; Ishida et al. 2000; Tyrrell and Attard 2002; Yang et al. 2003).
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234
FLOTATION FUNDAMENTALS
A
B
C
D
E
Source: Yang et al. 2003, reproduced with permission from the American Chemical Society.
FIGURE 4 TMAFM bubble coalescence images: (a) was observed at 20 hours after CO2–
saturated water was injected into the fluid cell; (b), (c), (d), and (e) were observed at time
intervals of 20 minutes after (a). Scan area: 10 μm × 5 μm; height = 300 nm.
The presence of very small gas bubbles (so-called nanobubbles) at structured solid–
water interfaces has been studied using the tapping mode atomic force microscopy
(TMAFM) imaging technique. Small bubbles do not form on smooth, hydrophilic, or
dehydroxylated silicon oxide wafer surfaces when immersed in aqueous solutions under
known levels of gas supersaturation. Randomly distributed small bubbles are observed over
the whole surface of observation on methylated surfaces and are larger and less densely distributed than on a smooth surface of similar hydrophobicity. The existence of these very
small gas bubbles on the surface is demonstrated by the observation of bubble coalescence
over a period of time (Figure 4).
The macroscopic contact angle, measured with respect to the aqueous or gas phase, is
very different from the microscopic contact angle detected by TMAFM and is possibly due
to the influence of line tension at the three-phase contact line. Surface heterogeneities act to
pin the contact line. The line tension (magnitude ~10–10 N) acts to stabilize the small bubbles, flattening them and thereby reducing the Laplace pressure (Yang et al. 2003).
D I R E C T E V I D E N C E F O R B U B B L E – PA R T I C L E AT T R A C T I V E
FORCES
Long-term fascination with hydrophobic attractive forces in flotation theory and practice
has existed, and some of the most famous names in colloid and surface science have contributed to the debate. Therefore, it is pertinent to recall some comments made by Sutherland
and Wark (1955) when discussing the rupture of an aqueous film on a patchy, hydrophobic
surface:
How is it that air suddenly takes the place of the water over the hydrophobic spot?
It seems certain that air itself is not the dominating factor, nor the water close to
the air/water interface. It is to the water–solid interface that we must look for the
first change which leads finally to air–solid contact. Whatever the correct explanation, we must be careful not to assume that air is able to “sense” some favorable surface condition through a thick film of water.
The question to ask now is “How thick is thick?”
The technique of gently pressing a bubble against a hydrophilic or hydrophobic plate
has been widely used in studying the variation of disjoining pressure with aqueous film
thickness, as well as for thin film drainage studies (Read and Kitchener 1969). Blake and
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PULP AND SOLUTION CHEMISTRY
235
10.0
F/R, mN/m
1.0
0.1
10–2 M
10–1 M
0.01
0
20
40
60
80
100
Separation, D, nm
Source: Fielden 1997.
FIGURE 5 Normalized (F/R = force/probe radius) force versus separation (D = distance) data for
the interaction of a silica sphere with an air bubble, and corresponding DLVO fits. Ψp = –100 mV
in 10–4 M NaCl (sodium chloride) and –27 mV in 10–2 M NaCl.
Kitchener showed that metastable, aqueous films were formed on methylated silica surfaces
(Blake and Kitchener 1972). These films were stabilized by electrostatic forces and became
increasingly unstable as the salt concentration increased. The minimum thickness at which
the smallest films were stable was 64 nm. Blake and Kitchener pointed out that the results
should not be interpreted as showing that hydrophobic forces extend to a bubble–particle
separation distance of this magnitude. The thickness represents an average for rupture that
was likely to occur at points where the film was locally thinner (e.g., where there were asperities, small particles of dust, external vibrations establishing variations in film thickness).
There have been only a few successful attempts to measure the interaction between a
particle and a bubble directly (Ducker, Xu, and Israelachvili 1994; Fielden, Hayes, and Ralston 1996; Preuss and Butt 1998). For a hydrophilic particle and an air bubble, the electrostatic interaction is generally repulsive (but refer to previous discussion) and sensitive to salt
concentration.
Although the distance between the bubble and the particle cannot be determined
exactly at this stage, the essential behavior fulfils DLVO expectations (Figure 5). When the
surface is strongly hydrophobic and moderately rough at the nanoscale level, the interaction
is repulsive at large separations, but an attractive “jump” into contact is evident as the separation
is decreased (Figure 6). The repulsion is electrical in nature, whereas the attraction, whose
time evolution is shown in the inset of Figure 6 and is completed in about 5 msec (see also
Figure 1), appears due to the presence of very small bubbles decorating the solid surface
(Figure 4).
These small bubbles alter the van der Waals interaction from repulsion (hydrophilic silica) to attraction (Mishchuk, Ralston, and Fornasiero 2002). During the time evolution of
the jump, thin film drainage and wetting perimeter formation takes place (Schulze 1984;
Nguyen, Evans, and Jameson 2001). In all of these cases, the results are semiquantitative,
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FLOTATION FUNDAMENTALS
10
Cantilever
Deflection, nm
0
F/R, mN/m
1
–200
–400
–600
–4
10 M
–2
10 M
–800
–1,000
0
2
4
6
8
Time, msec
0.1
10–4 M
10–2 M
0.01
0
20
40
60
80
100
Separation, D, nm
NOTES: Ψp = –100 mV (10–4 M) and –32 mV (10–2 M) (the Ψp value measured on
an OTS-coated silica plate interacting with a bare silica particle in 10–2 M NaCl
(–50 mV) was considered unreliable, so the value for dehydroxilated silica was
arbitrarily assigned based on similar behavior with a change in pH). A Hamaker
constant of –1 × 10–20 J was used. Inset: Oscilloscope measurement of
cantilever movement after the attractive jump.
Source: Fielden, Hayes, Ralston 1996, reproduced with permission from the American Chemical Society.
FIGURE 6 Normalized force (F/R) versus separation (D) between an octadecyltrichlorosilane
(OTS)-coated silica sphere and an air bubble as a function of NaCl concentration
because the bubble surface deforms as the particle approaches. To achieve more quantitative
results, the position of the liquid vapor interface must to be measured independently, along
with the position of the solid–liquid interface and the force, which are measured by the
atomic force microscopy (AFM) technique at this point (Ralston, Larson et al. 2005).
Moreover, only clean bubble surfaces have been studied to date.
P R O B I N G I N T E R FA C I A L C H E M I S T R Y
This section explores the means by which the chemistry of molecules is adsorbed at the
water–solid and water–vapor interfaces.
Although this list is not exhaustive, the chemical identity of adsorbed molecules at the
solid–water interface can be studied by in situ techniques such as attenuated total reflectance (ATR) and single-reflectance Fourier transform infrared spectroscopy, Raman spectroscopy, and second harmonic generation. These spectroscopic techniques are frequently
accompanied by dynamic electrochemical methods such as cyclic voltammetry. Ex situ spectroscopic techniques, such as x-ray photoelectron spectroscopy and time-of-flight secondary
ion spectroscopy, provide information on the chemical nature and distribution of adsorbed
molecules. Several of these techniques are applied to sulfide mineral oxidation, discussed
later in this chapter.
At the liquid–vapor interface, static and dynamic surface tension data reveal adsorption
densities and surface elasticity, whereas in situ, nonlinear optical techniques such as sumfrequency generation (SFG) provide complementary structural information.
At both interfaces, in situ scanned-probe microscopy (e.g., AFM) imaging provides
much useful dispositional information and some chemical information.
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Adsorption isotherms indicate the capacity of an interface to adsorb material and can
also yield the energetics of interaction. The solid–water interface hydrophobicity—the ultimate indication, in general, of whether or not attachment will take place—is perhaps most
usefully determined through contact angle measurement. The latter may now be reliably
determined for particles as well as flat surfaces and is a key component in flotation modeling.
In the following sections, the use of surface spectroscopy in mineral surface investigations is described, along with interfacial electrochemistry and solution chemistry. The use of
pulp and solution chemistry to improve selective flotation in several case studies is then
examined.
O X I D AT I O N O F S U L F I D E M I N E R A L S
It is well established that the flotation of a mineral particle depends on the proportion of
hydrophobic and hydrophilic species covering its surface, with the flotation rate and recovery increasing with the proportion of hydrophobic species. For a heterogeneous sulfide surface, the hydrophilic entities are mainly oxide/hydroxide/sulfoxy (e.g., sulfate, carbonate)
species, whereas the hydrophobic entities in the absence of a collector include sulfur-rich
species such as polysulfide and elemental sulfur (e.g., Fairthorne, Fornasiero, and Ralston
1997; Guy and Trahar 1985; Shannon and Trahar 1986).
The surface of a sulfide mineral is typically very reactive and starts to oxidize as soon as
the mineral is placed in contact with water and oxygen. The accepted mechanism for the
initial oxidation stage of sulfide minerals involves the migration of the metal from the outermost layers to the surface, followed by its dissolution in acidic solutions; whereas in alkaline
solutions, a layer of metal hydroxide is formed above the sulfur-rich mineral surface (Buckley and Woods 1984; Chander 1991; Ronngrem et al. 1991; Fornasiero, Eijt, and Ralston
1992; Fairthorne, Fornasiero, and Ralston 1997; Fornasiero, Li, and Ralston 1994; Fornasiero et al. 1994). The sulfur-rich surface consists of a metal-deficient sulfide lattice, which is
polysulfide or elemental sulfur depending on the extent of oxidation (Buckley and Riley
1991). The initial oxidation stage of a metal sulfide mineral (MS) surface may be represented by the following reactions.
1
2
in acidic conditions: MS + --nO 2 + 2nH + ↔ M 1 – n S + nM 2+ + nH 2 O
1
2
in alkaline conditions: MS + --nO 2 + nH 2 O ↔ M 1 – n S + nM ( OH ) 2
(EQ 1)
(EQ 2)
Upon further oxidation, polysulfide may oxidize to thiosulfate, sulfite, and sulfate ions
in neutral to alkaline solutions. The metal-deficient, sulfur-rich surface, M1–nS, is hydrophobic and is therefore responsible for the mineral collectorless flotation, whereas the metal
hydroxide surface layer is hydrophilic and depresses mineral flotation (Guy and Trahar
1985; Shannon and Trahar 1986; Chander 1991; Fairthorne, Fornasiero, and Ralston
1997). A thick layer of metal hydroxide may also prevent collector adsorption.
The products of Reactions 1 and 2 have a profound effect on mineral recovery, as illustrated in Figure 7, for the collectorless flotation of chalcopyrite. At low pH values, iron dissolves from chalcopyrite, leaving a hydrophobic, iron-deficient, sulfur-rich surface
(Equation 1), which results in a higher rate of chalcopyrite flotation. At high pH values, iron
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0.4
16
0.3
12
0.2
8
4
0.0
4
5
6
7
8
9
Metal Dissolved, 10 –5 mol dm 3
FLOTATION FUNDAMENTALS
Flotation Rate Constant, min –1
238
0
10
pH
Source: Fairthorne, Fornasiero, and Ralston 1997.
FIGURE 7 Collectorless flotation rate constant of chalcopyrite as a function of conditioning
pH. The dissolution of iron and copper is represented by the empty and filled bars, respectively
(N2 gas conditioning).
hydroxide is formed on the chalcopyrite surface (Equation 2), which results in a lower rate
of flotation. Copper dissolution in acidic pH conditions is much less than that of iron.
According to Reactions 1 and 2, the extent of surface oxidation depends on variables
such as solution pH and oxygen content, but also on the pulp oxidation potential (Eh), as
species are oxidized or reduced in solution and at the mineral surface (e.g., O2 + 4 H+ + 4 e–
↔ 2 H2O). Furthermore, interactions between minerals themselves and between minerals
and grinding media (galvanic interaction) during the grinding and conditioning stages will
also affect mineral oxidation through the transfer of electrons from the most anodic to the
most cathodic material (Rao, Labonte, and Finch 1992). Because pyrite is the most cathodic
sulfide mineral, the selective depression of its flotation in alkaline pH conditions and in the
presence of oxygen has been attributed to galvanic interactions with less cathodic sulfide
minerals present in the pulp, which results in increased oxygen reduction on the pyrite surface to produce ferric hydroxide (Shen, Fornasiero, and Ralston 2001). Galvanic interaction
will also increase the dissolution of the anodic sulfide mineral, which in the case of copper
sulfide or galena may lead to the inadvertent copper or lead activation of sphalerite and iron
sulfides. One way of limiting the effects of galvanic interactions is to reduce the oxygen content
of the pulp by purging with nitrogen gas (Martin et al. 1989). Table 2 lists the various sulfide
minerals as a function of their cathodic character.
The pulp solution may also contain metal ions, either dissolved from grinding media
and minerals present in the ore or intentionally added. With increasing pH, these metal ions
will hydrolyze and adsorb or precipitate on the mineral surface. The minimum pH value for
metal hydroxide precipitation in solution is well characterized (Fuerstenau and Palmer
1976) and corresponds quite well to the adsorption edge of these metal hydroxides on mineral surfaces, as measured by zeta potential or x-ray photoelectron spectroscopy (XPS) measurements (Fornasiero, Li, and Ralston 1994; Fornasiero et al. 1994; Clarke et al. 1995).
The pKa of the metal hydroxide species formation is well documented for most of the metal
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PULP AND SOLUTION CHEMISTRY
239
TABLE 2 Approximate order of collectorless flotation of sulfide minerals (descending order of
floatability) and their rest potential in water at pH 4
Mineral
Molybdenite, MoS2
Stibnite, Sb2S3
Argentite, Ag2S
Galena, PbS
Bornite, Cu5FeS4
Covellite, CuS
Sphalerite, ZnS
Chalcopyrite, CuFeS2
Marcasite, (Zn,Fe)S
Pyrite, FeS2
Rest Potential, Volts, SHE*
0.11
0.12
0.28
0.40
0.42
0.45
0.46
0.56 (anomalous)
0.63
0.66 (most cathodic)
Source: Majima 1969; Hayes and Ralston 1988.
*SHE = standard hydrogen electrode.
ions and has been used to calculate, as a function of pH, the amount of adsorbed/precipitated metal species on the surface of several sulfide minerals, for example, pyrite, galena, and
sphalerite (Sun et al. 1991; Ronngren et al. 1991; Fornasiero, Eijt, and Ralston 1992; Fornasiero, Li, and Ralston 1994; Fornasiero et al. 1994).
M I N E R A L S U R FA C E C H A R A C T E R I Z AT I O N
The oxidation of sulfide minerals and the underlying mechanisms have been studied in great
detail as a function of conditioning time, pH, Eh, and dissolved gas in the pulp, with complementary analytical techniques probing the species on the surface, as well as those that
have dissolved from the surface into solution. Surface-sensitive, in situ analytical techniques
include cyclic voltammetry and zeta potential measurements, while ex situ, high-vacuum
techniques include XPS and time-of-flight secondary ion mass spectroscopy (TOF-SIMS).
Because these techniques probe different types of species on the surface and to different
depths, using a combination of these surface-sensitive techniques should provide a more
complete picture of the mineral surface, similar to what a gas bubble “senses” when it collides with a mineral particle in a flotation cell. For example, zeta potential measurements
will only give information on surface-charged species and cyclic voltammetry will only
detect oxidation and reduction reactions occurring on the surface. XPS and TOF-SIMS
have provided the most detailed information on the type and proportion of surface species
on particles taken from mineral mixtures and ores in the laboratory or plant (Smart 1991;
Smart et al. 2000; Kanta, Sedev, and Ralston 2005).
Zeta Potential Study of Sulfide Mineral Oxidation
Microelectrophoresis, which enables electrophoretic mobility to be measured and from
which the zeta potential is derived, is an in situ surface-sensitive technique that probes the
electric potential and, hence, the charge on the mineral surface in contact with the solution
phase. The zeta potential of sulphide minerals is very sensitive to surface oxidation, with
both a sign reversal and a change in zeta potential value of up to 80 mV occurring between a
nonoxidized and fully oxidized surface. Healy and Moignard’s review (1976) on the zeta
potential of sulfide minerals was the first comprehensive analysis of this subject. It highlighted the deficiencies in the experimental procedures and in the interpretation of the zeta
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240
FLOTATION FUNDAMENTALS
20
A
ZnO
20
0
3.0
Zeta Potential, mV
4.1
6.5
–20
–2
8.5
0
ΔpH 2
S0
–20
40
B
40
0
20
–40
0
0 ΔpH 4
–4
–20
–40
–60
2
4
6
8
10
12
pH
Notes: The filled circles and triangles represent the experimental zeta
potential to infinite sphalerite concentration and the zeta potential of
nonoxidized sphalerite (IEP = 1.6), respectively. Insets: The same
experimental data as in (a) and (b) are shown, but as a function of
ΔpH (= pH–IEP).
FIGURE 8 Zeta potential of sphalerite as a function of pH (experimental data from Healy and
Moignard 1976): (a) at various conditioning pH values and for zinc oxide (ZnO) and elemental
sulfur, S0 (b) for sphalerite concentration of (curves shifting to higher pH values) 0.1, 0.25, 0.50,
1.0, 2.0 g dm–3 and for ZnO
potential data from earlier studies. Their detailed analysis of the electrophoretic mobility
(or zeta potential) of sphalerite in various pretreatment conditions served as a model for
subsequent zeta potential studies on the oxidation of sulfide minerals (e.g., Fornasiero, Eijt,
and Ralston 1992; Fornasiero, Li, and Ralston 1994; Fairthorne, Fornasiero, and Ralston
1997). During mineral oxidation, the changes of the zeta potential versus pH curves for
most sulfide minerals follow the same trends with increasing conditioning time, pH, or oxygen content of the solution (Healy and Moignard 1976; Fornasiero, Eijt, and Ralston 1992;
Fornasiero, Li, and Ralston 1994; see Figure 8).
The isoelectric point (IEP) of a mineral is characteristic of that mineral (Parks 1965)
and is the pH value (or pI if I is another potential-determining ion) where the zeta potential
is zero. It is generally not possible to measure the IEP of sulfide minerals because of the fast
oxidation of their surface, and, in most cases, an apparent IEP is measured, representing the
extent of mineral oxidation. For example, in the review of Healy and Moignard (1976), the
IEP of sphalerite increased with conditioning pH value or time from a low value, comparable
with the IEP of elemental sulfur (IEP = 1.6), up to that of zinc oxide (IEP = 8.5) (Figure 8).
During oxidation, both elemental sulfur and zinc oxide-hydroxide are formed at the sphalerite surface (Moignard, Dixon, and Healy 1977). The zeta potential and IEP represent the
contributions of both surface species.
In the first step of sulfide mineral oxidation, it is generally recognized that after preferential metal dissolution, a metal-deficient sulfur-rich surface is formed (Reaction 1), which
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PULP AND SOLUTION CHEMISTRY
241
later oxidizes to polysulfide and elemental sulfur (Buckley and Walker 1988; Buckley and
Woods 1984).
nS 2– ↔ S n2– + 2 ( n – 1 )e –
(EQ 3)
S n2– ↔ nS 0 + 2e –
(EQ 4)
and
These groups on the sulfur-rich surface may be attributed to sulfides (–SH) that can
gain a proton and become positively charged (–SH2+) or lose a proton and become negatively charged (–S–) for pH values lower or higher than their IEP of 1.6, respectively, in a
similar manner to metal oxides (de Bruyn and Agar 1962). The sulfide (SH) group has been
detected on the sphalerite surface by Raman spectroscopy (Gard, Sun, and Forsling 1995).
Upon oxidation, the sulfide mineral surface becomes increasingly covered with metal
oxide/hydroxide species, and the corresponding zeta potential versus pH curves become less
negative and even become positive (Fornasiero, Li, and Ralston 1994; Fairthorne, Fornasiero, and Ralston 1997), falling between the extremes of elemental sulfur and the corresponding metal oxide/hydroxide species (Figure 8).
Similar results were obtained in a zeta potential study of the oxidation of the copper
sulfide minerals chalcopyrite, bornite, chalcocite, covellite, tennantite, and enargite conditioned at pH 11 with nitrogen gas, oxygen gas, or hydrogen peroxide (Figure 9) (Fullston,
Fornasiero, and Ralston 1999a, b). The zeta potential was also monitored as a function of
pH in acid and base titrations to study metal hydroxide dissolution and adsorption/precipitation
in acidic and neutral pH conditions, respectively. From this detailed analysis, it was shown
that the oxidation of these minerals follows the order: chalcocite > tennantite > enargite >
bornite > covellite > chalcopyrite, which indicates that the separation by flotation of the
arsenic minerals of tennantite and enargite from the other copper minerals based on differences in surface oxidation is only possible if chalcocite is absent from the mineral mixture.
Other Surface-Sensitive Analytical Techniques to Measure the Oxidation
of Sulfide Minerals
Oxidation and reduction reactions occurring on the surface of sulfide minerals may be identified by cyclic voltammetry (Gardner and Woods 1979; Hamilton and Woods 1984). Figure 10 shows a typical voltammogram of a chalcocite electrode conditioned with oxygen at
pH 8.5 and 10.8. At pH 8.5, the first major peak at +0.08 V (vs. the standard calomel electrode, or SCE) in the anodic scan has been assigned to the oxidation of the chalcocite mineral to form covellite and also copper hydroxide (Roos, Celis, and Sudarsono 1990).
Cu 2 S + 2H 2 O ↔ CuS + Cu ( OH ) 2 + 2H + + 2e –
E h = +0.06 V (vs. SCE) (EQ 5)
The second major peak in the anodic scan is present at +0.21 V (vs. SCE) and has been
identified as the oxidation of covellite (Reaction 3). The covellite (CuS) is present in the
chalcocite sample but is also formed during the previous oxidation reaction (Equation 5)
(Roos, Celis, and Sudarsono 1990).
CuS + 2 H 2 O ↔ Cu ( OH ) 2 + S + 2H + + 2e –
E h = +0.12 V (vs. SCE)
(EQ 6)
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242
FLOTATION FUNDAMENTALS
Chalcopyrite
40
Bornite
20
0
–20
H2O2
Zeta Potential, mV
O2
–40
O2
N2
N2
Chalcocite
40
Covellite
20
0
O2
–20
O2
–40
N2
N2
–60
4
6
8
10
4
6
8
10
12
pH
NOTE: Filled and empty symbols refer to a pH change from high to
low pH values and from low to high pH values, respectively (the
arrows show the direction of pH change).
FIGURE 9 Zeta potential versus pH curves of copper sulfide minerals conditioned at pH 11.0
for 20 minutes in nitrogen (N2), for 60 minutes in oxygen (O2), and for 60 minutes with hydrogen
peroxide (H2O2)
On the cathodic scan, the double peaks at –0.10 and –0.16 V correspond to the reduction of the oxidation products formed at 0.21 and 0.08 V, respectively (Reactions 5 and 6).
A similar voltammogram is observed at pH 10.8. The only difference is the shift of all
the peaks by approximately –0.16 V compared with those at pH 8.5. Also, the –0.08 V
anodic peak is very broad, especially on its more positive side, and hides the smaller peaks
observed at +0.08 and +0.21 V at pH 8.5. A good understanding of sulfide mineral oxidation has been obtained by combining several analytical techniques such as cyclic voltammetry and XPS (Buckley, Hamilton, and Woods 1985).
XPS is, by far, the most used surface analytical technique for the study of sulfide mineral oxidation, because it can more directly measure the type and proportion of surface species.
It has been able to confirm the mechanism of oxidation through Reactions 1 to 4, as a function of pulp conditions such as pH and dissolved oxygen (DO) content (Buckley and
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PULP AND SOLUTION CHEMISTRY
243
1
3
0
Current, mA
pH = 10.8
3
2
1
0
1
1
pH = 8.5
3
–0.85
–0.65
–0.45
–0.25
–0.05
0.15
0.35
Potential, V (vs. SCE)
FIGURE 10 Voltammograms of chalcocite conditioned in oxygen at pH values of 8.5 and 10.8
(1, 2, and 3 refer to Reactions 3, 4, and 5, respectively)
Woods 1983, 1984; Buckley and Walker 1988; Buckley, Woods, and Wouterlood 1989;
Buckley and Riley 1991; Fornasiero et al. 1994, Fairthorne, Fornasiero, and Ralston 1997).
As an example, Figure 11 shows the XPS spectra of chalcopyrite conditioned at pH 5.0 or
9.5 and with nitrogen or oxygen gas. With this analytical technique, all the major species
involved in the sulfide mineral oxidation mechanism (Reactions 1 to 4) were identified, and
their proportions on the chalcopyrite surface were monitored with changes in pulp conditions. It was found that more iron dissolves from the chalcopyrite lattice than copper, and,
therefore, more ferric hydroxide was formed than copper hydroxide. In particular, the ratio
of ferric hydroxide (Fe(OH)3) to ferrous sulfide (FeS) doubled with oxygen purging,
whereas the ratio of copper hydroxide (Cu(OH)2) to CuS remained constant at pH 9.5
(Figure 11).
U S E O F D I AG N O S T I C P U L P A N D S O L U T I O N C H E M I S T RY
I N S E L E C T I V E F L O TAT I O N
This section discusses some practical case studies where pulp and solution chemical measurements are used as a first step toward process optimization. First, a typical plant survey
and key sample points for pulp chemical measurements are discussed. Next, the importance
of correlating laboratory conditions with plant performance are outlined. Finally, examples of
measurements taken from a plant and their interpretation are discussed in two studies.
Pulp Chemical Measurements
It is very important to measure both the primary ball mill discharge Eh and pH during a typical plant survey because this is usually the point in the process where the Eh is at its lowest
value ( Johnson 1988) and corresponds to the state of least oxidation (Figure 12).
This sample point is also important because it indicates the extent of media oxidation
(Grano et al. 1994), and the possibility of maintaining low Eh values during collector
conditioning and subsequent flotation ( Johnson, Jowett, and Heyes 1982). Measurements
conducted at the primary ball mill discharge point can also be very useful in demonstrating
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FLOTATION FUNDAMENTALS
5.0/N2
Fe(OH)3
Intensity, arbitrary units
9.5/N2
9.5/O2
FeS
704
708
712
716 704
708
712
716 704
708
712
716
Binding Energy, eV
Intensity, arbitrary units
S2p
Cu2p(3/2)
Sn2–
CuS
S2–
S0
158
162
Cu(OH)2
166
930
934
938
Binding Energy, eV
Source: Fairthorne, Fornasiero, and Ralston 1997.
FIGURE 11 Top: Fe2p(3/2) XPS spectra of chalcopyrite as a function of conditioning pH and type
of gas; Bottom: S2p and Cu 2p(3/2) XPS spectra of chalcopyrite conditioned at pH 9.5 in oxygen
(lines represent the calculated XPS spectra with their individual components)
Reagents
Primary
Cyclone
Overflow
Process
Water
Conditioned
Rougher Feed
Rougher Tailing
X
Rougher
Block
X
X
Process
Water
X
SAG
Feed
SAG Mill X
Primary
X Ball Mill
X Rougher Concentrate
X
Reagents
Regrind
Ball Mill
Reagents
Plant Condtioning
Includes Regrind
Pump Box and Cyclone
Cleaner
Block
X = Sampling Point
Reagents
X Regrind Cyclone Overflow
Reagents
Process Water
X
Cleaner Tailing
X Final Concentrate
FIGURE 12 Typical points (x) for measuring pulp chemistry values in a flotation plant,
showing typical reagent and process water addition points
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245
the difference between the semiautogenous grinding (SAG) and primary ball mill chemical
environments (Grano et al. 1994). This is critical for subsequent unit flotation stages in the
primary grinding circuit that may be adversely affected by low Eh values. The pulp chemical
change across the primary ball mill can be assessed by comparing the primary cyclone underflow with the primary ball mill discharge. The cyclone overflow should be measured separately from the “conditioned rougher feed” because of the reagents, as well as aeration in
cycloning and pumping from the primary ball mill circuit to the flotation feed. The “conditioned rougher feed,” of course, is the feed presented to rougher flotation with all reagents
added. The regrind ball mill discharge and regrind cyclone overflow should also be separately measured. If there are any conditioning stages, such as pulp heating, pH adjustment,
aeration, or reagent additions (e.g., CuSO4, SO2), the change in pulp chemistry across the
conditioning stage should also be assessed. An excellent example of the use of diagnostic Eh
and pH values to solve plant problems is discussed by Johnson and Munroe (1988). The
utility of pulp and solution chemistry measurements in solving processing problems is outlined in the following section. These measurements involve Eh and pH inorganic and
organic composition of the circuit water, extractable metal ions from mineral surfaces, and
temperature.
The circuit water should be assessed, including other key streams in the process water
circuit (e.g., tailings return water, makeup water, mine water, tailings thickener overflow
water, concentrate thickener overflow water), as required (Levay, Smart, and Skinner 2001).
The importance of process water chemistry is illustrated by an example where copper (II) in
solution, emanating from a thickener overflow stream (which contained 10 ppm Cu concentration at pH 7), inadvertently activated sphalerite in the lead circuit of a lead/zinc flotation plant. This caused sphalerite to report to the lead concentrate and, thus, dilute the
latter. With this analysis information available, steps were taken to minimize this inadvertent copper activation. Tailings return water, which contained much lower levels of copper
(<0.5 ppm), was used instead as the makeup water in the lead circuit. The importance of
water quality is again illustrated by the use of fresh water in the zinc-cleaning stages of a
lead/zinc flotation plant, where lime consumption was reduced to achieve a set-point pH
value and to improve selectivity between sphalerite and pyrite. The normal process water at
this plant is supersaturated in calcium sulfate, suggesting that calcium sulfate has and/or is
probably precipitated/precipitating in the process pulps (Grano et al. 1995). This has a pH
buffering effect, which means that higher lime additions are required to achieve a set pH
(~11) and that more calcium sulfate will precipitate because of lime addition, a source of
calcium.
Samples and measurements should be taken to compare the products of the unit separation steps (Pietrobon, Georgiev, and Grano 2001), including those across the rougher block
(i.e., rougher concentrate, rougher tailing reporting to final tailing, and rougher feed) and
cleaner block (i.e., final concentrate, cleaner block tailing reporting to final tailing, and
cleaner feed). Ethylenediaminetetraacetic (EDTA) extractions and surface analysis samples
contrast the oxidation state of mineral particles present in these streams. In general, the valuable metal in the tailing streams has a greater proportion in an oxidized state than does the
valuable metal in the concentrate streams, both measured on a mass of metal basis when
using the EDTA technique (Kant, Rao, and Finch 1994).
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Correlating Laboratory and Plant Conditions and Performance
Correlation of the conditions and performance from the laboratory to the plant is extremely
important when any optimization work is to be contemplated at laboratory scale. Laboratory experiments are often performed under conditions that are far removed from the plant
(e.g., adding lime in the conditioning stage in the laboratory in contrast to adding lime to
the mill at the plant). In order to successfully transfer the results of laboratory studies to
plant scale, it is necessary, at a minimum, to mimic the plant pulp chemical conditions at
laboratory scale. The advantage is that the initial stages of laboratory investigations will be
focused on the effect of simple parameters that may be relatively easy to manipulate at plant
scale and that may acutely affect plant performance. For example, the effect of water quality
or water source, grind pH, aeration, and standard plant reagent additions may be evaluated.
Future ores that, by definition, have not been processed through the plant may be tested
with greater confidence.
During a survey, a SAG mill feed sample (Figure 13) is treated in the usual way to produce samples suitable for flotation testing. Storage of the crushed and riffled sample should
aim to minimize oxidation; however, the extent of oxidation must be readily assessed. Flotation tests on the conditioned rougher feed should be undertaken in triplicate under known
physical conditions such as cell type and mechanism, gas flow rate, impeller speed, and bubble size, if possible. These same physical conditions are to be used in later testing of the SAG
feed sample to ensure that differences between the plant and the laboratory are only due to
differences in the feed (properties) presented to the laboratory flotation cell. A large sample
of circuit water should also be collected and used in laboratory flotation testing (Figure 13).
However, circuit water samples may not be stable because of the continuing oxidation of
thio-salts (Rolia and Tan 1985) and biological activity. In this case, a simulated circuit water
sample should be prepared, based on the inorganic composition of the circuit water, that
matches all of the major constituents of the circuit water during the survey, including the
total dissolved solids. This is particularly important for pulps containing very high concentrations of dissolved species (Chen et al. 1999). Unfortunately, it is very difficult to match
the organic components, which may include collector decomposition products, bacteria,
and residual polymer from dewatering. Comparing the flotation performance with wellcharacterized water types can be a very diagnostic measure of the effect of water quality on flotation.
After these samples have been collected and appropriately treated, the SAG feed, water
samples, and reagents may be used at laboratory scale. A special mill (e.g., Magotteaux Mill)
allows the pH to be continuously controlled during grinding (Greet, Small et al. 2004),
which may be important in matching the laboratory to the plant, because the pH of the feed
to most plant primary ball mills (i.e., the cyclone underflow) is usually close to the pH of the
plant ball mill discharge. In contrast, most batch mills exhibit wide variations in pH from
the beginning of grinding to its completion, with the mill discharge pH usually set at the target pH value. The Magotteaux Mill permits continuous purging with different gases (e.g.,
N2, O2) during grinding to ensure that the laboratory mill discharge Eh is at the same value
as the plant ball mill discharge. Purging during grinding and postgrinding may also be very
informative in terms of process optimization. After grinding is complete, the slurry is conditioned with the same reagents and at the same concentration as those in the plant during the
survey period. Therefore, laboratory flotation is conducted under the same physical conditions as those applied to the laboratory flotation tests conducted during the plant survey.
With care, it is possible to duplicate the flotation test performance in the laboratory and in
the plant survey (Pietrobon, Grano, and Greet 2004). However, differences can highlight
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247
Ore
(SAG Feed)
Reagents
Reagents
Process
Water
Plant Grinding
Circuit
Laboratory
Grinding
Reagents
Reagents
Laboratory
Condtioning
Plant Condtioning
Matching Points
Plant Flotation
Laboratory Flotation
Source: Greet, Small et al. 2004, reproduced with permission from Minerals Engineering.
FIGURE 13
Procedure for correlating the laboratory and plant pulp chemistry conditions
variables that affect plant performance, including grind pH, grind Eh, flotation pH, flotation Eh, water quality, pulp temperature, grinding media type, reagent quality, and so forth.
C A S E S T U DY 1 : L A B O R AT O R Y A N D P L A N T C O R R E L AT I O N
This case study involves laboratory conditions that were correlated with the plant’s chemical
conditions (Pietrobon, Grano, and Greet 2004). In this plant, the valuable minerals are
galena (9%) and sphalerite (12%), contained in a high-pyrite (47%) matrix in the feed.
Because of the high-pyrite feed, oxygenation was applied to the primary cyclone overflow in
two separate tanks (in which ~300–350 m3 of O2 hr–1 or 2.1–2.5 dm3 of O2 kg–1 of ore is
sparged) prior to galena flotation, in order to passivate the pyrite surface and enhance galena
flotation. The pH in the SAG mill discharge was 10.0, due to lime addition to the SAG feed
(Table 2). Lime addition, process water, and ore dissolution inputs all contributed to the dissolved and precipitated metal ion composition of the grinding circuit product. The proportion of each contribution is discussed later in this case study. However, it is clear that the
process water was a major contributor to the grinding circuit product composition, at least
in the case of the magnesium, sodium, chloride, and sulfate concentrations (Table 3). The
pH of the ball mill discharge was lower because of the oxidation of sulfide minerals and
grinding media during ball mill grinding, for which there was no separate lime addition
point. At this sample point, the Eh was at its lowest value in the circuit. The Eh increases
when transferring the pulp to the cyclone overflow and after oxygenation. Clearly, it is
important to match pulp chemical conditions in both grinding and postgrinding during the
oxygenation stage.
The ore was not oxidized in the SAG mill discharge, as shown by the very low EDTAextractable metal ions at this sample point (Table 3). The minerals oxidized during the subsequent oxygenation step, as indicated by the increase in the quantity of EDTA-extractable
metals at this sample point. Galena oxidizes more rapidly than either sphalerite or iron sulfide minerals (Table 3). It is usually accepted that galena oxidation is detrimental to recovery
of galena in flotation (Guy and Trahar 1984), shown here by the increase in EDTA-extractable lead for the flotation tailings streams relative to the concentrate streams (Pietrobon,
Grano, and Greet 2004). However, it is likely that the high oxygen demand of the pulp prior
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248
TABLE 3
FLOTATION FUNDAMENTALS
Pulp chemical measurements at sample points in a lead/zinc flotation plant
Sample
Eh, mV
Point
pH
SHE*
SAG
10.0
+50
discharge
Ball mill
7.5
–90
discharge
Primary
9.1
+110
cyclone
overflow
Oxygenation 8.8
+245
stage
Process
6.3
—
water
EDTAExtractable
Metal Ions, %
–3
Dissolved Ions, mg dm
Mg
Pb
Zn
Na
Cl
SO42–
HCO3–
Pb
Zn
Fe
1,350
95
0.5
0.4
1,300
1,250
1,200
55
0.1
0.0
0.0
1,300
165
0.5
0.5
1,400
1,250
1,450
150
0.1
0.0
0.0
1,150
135
0.4
0.3
1,300
1,150
1,000
150
4.4
0.2
0.5
1,250
145
1.7
1.4
1,300
1,250
1,350
38
9.7
0.4
0.7
905
145
4.6
25.0
1,340
1,150
1,200
38
—
—
—
Ca
*SHE = standard hydrogen electrode.
to oxygenation plays a determining role. In this case, the low DO content of the pulp, due to
its rapid removal by sulfide mineral and media oxidation, may inhibit collector adsorption
on unoxidized galena (Guy and Trahar 1984). Figure 14 shows the DO content of the sampled pulps as a function of time (i.e., DO demand)—the DO demand of the fresh pulp after
grinding (i.e., cyclone overflow) is much greater than the pulp after oxygenation in the
tanks. This demonstrates the importance of measuring the DO demand (Greet, Steinier, et
al. 2004) as well as the instantaneous DO concentration.
Laboratory experiments were then conducted in which the SAG feed sample, collected
during the plant survey, was ground in a Magotteaux Mill (at a controlled pH 9.1) and the
slurry purged with oxygen gas in a laboratory flotation cell for different periods of time and
at a fixed gas flow rate. This stage was designed to simulate the oxygenation that takes place
in the tanks of the plant. With increasing oxygenation, the laboratory-ground pulps exhibit
a reduced oxygen demand (Figure 15). In the absence of oxygenation, the mill discharge
shows very high DO demand, very similar to that of the plant cyclone overflow stream (i.e.,
prior to the tank; Figure 13). After 20 minutes of oxygenation (Figure 14), the DO demand
of the laboratory-ground sample approached that of the tank product (Figure 13). Twenty
minutes of oxygen purging under these conditions closely matched the DO demand of the
tank product. This oxygen purging time was subsequently used in flotation experiments on
the SAG feed sample, with the objective of matching the flotation performance of the laboratory flotation test on the tank product during the plant survey. With 60 minutes of oxygenation (Figure 15), it is clear that the minerals are very heavily oxidized and no longer
significantly consume oxygen.
Laboratory tests conducted after 20 minutes of oxygenation time with demineralized
water were analyzed to determine the dissolved species at pH 9.1, using sodium hydroxide as
the pH regulator in grinding (Table 4, Stream 2, ore dissolution). This pulp solution composition
does not use process water or lime and may be compared with the actual solution composition of the tank discharge solution (Table 4, Stream 6, tank solution). There is a relatively
minor concentration of ions in solution, suggesting only modest dissolution of the ore components under these conditions. It is clear that process water and lime addition are required
to adequately simulate the plant solution composition. This point is further highlighted by
mathematical addition of the process water (Table 4, Stream 1, process water), ore dissolution
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PULP AND SOLUTION CHEMISTRY
249
20
Cyclone Overflow
Tank 1 Product
Tank 2 Product
18
Dissolved Oxygen, mg dm –3
16
14
12
10
8
6
4
2
0
0
1
2
3
4
5
6
7
8
9
10
Time, min
Source: Pietrobon, Grano, and Greet 2004, reproduced with permission from Minerals Engineering.
FIGURE 14 DO concentration as a function of time (DO demand) for the plant cyclone
overflow, oxygen tank (1 and 2 in series) product streams after the discontinuation of oxygen
gas purging in a lead/zinc flotation plant
20
18
Dissolved Oxygen, mg dm –3
16
0 min Oxygenation Time
20 min Oxygenation Time
60 min Oxygenation Time
14
12
10
8
6
4
2
0
0
1
2
3
4
5
6
7
8
9
10
Time, min
Source: Pietrobon, Grano, and Greet 2004, reproduced with permission from Minerals Engineering.
FIGURE 15 DO concentration as a function of time (DO demand) for the SAG mill feed sample
ground in Magotteaux Mill (grind pH 9.1) using circuit water after the discontinuation of oxygen
gas purging
(Table 4, Stream 2), and lime (Table 4, Stream 3, reagents) inputs to the grinding stage to
produce a simulated solution composition prior to equilibration (Table 4, Stream 4, sum).
Most of the dissolved magnesium, sodium, chloride, and sulfate in this simulated grinding circuit product are derived from the circuit water. In the case of calcium, significant
quantities emanate from the lime addition to the grinding circuit, as well as the circuit water.
Chemical “equilibration” of this simulated solution shows that gypsum will precipitate and
the pH will decrease with time (Table 3, Stream 5). Comparison of this equilibrated, simulated
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FLOTATION FUNDAMENTALS
TABLE 4 Dissolved and precipitated species in the pulp during plant survey (tank solution)
and process water after laboratory grinding in demineralized water (ore dissolution) and in
circuit water (simulated pulp)
Dissolved Ions, mg dm–3
1.
2.
3.
4.
5.
6.
7.
8.
9.
10.
Stream
Process water
Ore dissolution
Reagents
(lime)
Sum (streams
1 + 2 + 3)
MinteqA2 in
solution
Tank solution
MinteqA2 for
tank
(Stream 5 –
stream 6) =
Simulated pulp
(Stream 5 –
stream 9) =
pH
6.3
9.1
Ca
905
35
Mg
145
8
Pb
4.6
0.2
Zn
25.0
0.1
Na
1,340
110
Cl
1,150
25
SO42–
1,200
115
HCO3–
38
36
—
710
—
—
—
—
—
—
—
153
4.8
25.1
1,450
1,175
1,315
74
1,650
4.9
1,412
153
4.8
25.1
1,450
1,175
745
74
8.4
1,250
145
1.7
1.4
1,300
1,250
1,350
38
4.5
1,045
145
1.7
1.4
1,300
1,250
1,130
38
162
8
3.1
23.7
150
–75
–605
36
9.1
1,000
120
0.4
0.1
1,400
2,300
2,100
18
412
33
4.4
50
1,125
–1,355
56
25
NOTE: Solution data were chemically balanced using a chemical speciation program called MinteqA2/Prodefa2 (EPA 1991)
and were used to predict species most likely to precipitate from the process water and pulp solutions.
solution to the actual composition of the tank discharge (Table 3, Stream 6) shows that significant (on a relative basis) additional calcium, lead, and zinc have been removed during
grinding beyond that predicted by simple solution equilibration. In the case of lead and
zinc, these will have precipitated, and possibly adsorbed onto mineral surfaces, at the higher
pH value of the former plant stream. In the case of calcium, adsorption onto minerals seems
likely. In the case of sulfate, the tank discharge shows a higher concentration than that
expected after solution equilibration, which is similar to that obtained from the simulated
solution without equilibration. This suggests that although gypsum precipitation is
expected under solution equilibrium conditions, equilibrium has not been obtained in the
grinding circuit pulps. Furthermore, calcium is removed from solution, most likely via
adsorption, without complete precipitation of gypsum. Chemical equilibration of the tank
discharge solution also shows that gypsum may precipitate and the pH will decrease with
time (Table 3, Stream 7).
The actual composition of the SAG feed sample after laboratory grinding at pH 9.1 in
circuit water, with lime addition for pH control and 20 minutes of postgrinding oxygenation (Table 4, Stream 9) can now be determined. As for the tank discharge, the SAG feed
sample under these conditions shows less calcium, lead, and zinc in solution relative to that
predicted on the basis of the mass balance of inputs and equilibration (Table 6, Stream 10).
It is likely that in the case of this particular laboratory condition, calcium, lead, and zinc are
being removed from the process water input because of precipitation and adsorption. In the
case of sulfate, the SAG feed sample under these conditions shows a higher concentration,
again suggesting that calcium is predominantly removed from solution, not as precipitated
gypsum but most likely by means of adsorption onto minerals. Finally, comparison of the
tank discharge (Table 4, Stream 6) with the laboratory value under similar conditions (Table
4, Stream 9) shows reasonable agreement with the exception of chloride and sulfate. Under
these conditions, the flotation results show a close correlation between the plant and laboratory, demonstrating that these differences may not be significant to flotation.
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PULP AND SOLUTION CHEMISTRY
251
800
Rougher Cell 9
Surveys 1 and 2
O2
H 2O
Eh, mV
400
CuO + S0
CuS + Cu2–xS
0
Plant Cyclone
Overflow Survey 1
Cu2S
Cu2–xS + CuO
H 2O
Plant Cyclone
Overflow Survey 2
Cu2S
H2
–400
Cu0 + H2O
Cu0 + HS
Cu + S2–
–800
0
2
4
6
8
10
12
14
pH
Source: Orwe, Grano, and Lauder 1998, reproduced with permission from Minerals Engineering.
FIGURE 16 Eh–pH predominance diagram for the Cu-S-H2O system showing experimental
Eh–pH points for the plant surveys
C A S E S T U DY 2 : O X I D AT I O N / R E D U C T I O N P O T E N T I A L S
A N D D I S S O LV E D O X Y G E N D E M A N D
To further demonstrate the value of undertaking simple pulp chemical measurements during
a plant survey, a second case study was considered. This example comes from a flotation
plant that normally treats porphyry copper ore (i.e., copper in chalcocite and chalcopyrite in
an igneous host rock), but for which the proportion of the run-of-mine (ROM) ore from
sulfide skarns is generally increasing throughout the remainder of the mine life. Sulfide
skarns arise from the igneous intrusion into the country rock and generally contain higher
copper and gold concentrations, as well as higher pyrite contents. When treating feeds with
high sulfide skarn content, the flotation plant encounters significant problems such as low
copper recovery (<75%) and a low final concentrate grade (<20% Cu). Ore blending is only
partially effective and cannot be used in the latter stages of mine life because of the preponderance of sulfide skarn. The low concentrate grade is a consequence of both high pyrite
recovery and low copper recovery. The normal conditions in the plant (lime addition to the
SAG feed to pH 11–11.5 and use of a dithiophosphate collector in the primary ball mill)
were inadequate for the selective separation of sulfide skarn ores. The copper-bearing minerals are adequately liberated from the gangue, that is, ~80% fully liberated across all size
ranges at the normal grind size of 150 μm in the flotation feed. The reason for the low copper concentrate grade and recovery was related to pulp chemistry differences between the
porphyry and sulfide skarn ores.
Consider the Eh–pH diagram for the Cu-S-H2O system (Figure 16). The Eh–pH measurements from two plant surveys (Survey 1 and Survey 2) on a porphyry ore are shown.
The ore’s principal copper-bearing sulfide minerals are chalcocite and chalcopyrite. The typical pH for this plant is 11.5. The Eh–pH diagram shows that chalcocite is only stable to an
Eh of approximately 0 mV at pH 11, whereas above this Eh value it may oxidize. In Survey 1,
the Eh of the pulp at the cyclone overflow is +150 mV, suggesting that chalcocite oxidation
may occur. Oxidation was confirmed both by EDTA extraction and surface analysis
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FLOTATION FUNDAMENTALS
(Putubu, Grano, and Morey 2001). In this system, further aeration of the pulp in flotation
does not appreciably increase the Eh, suggesting that the pulp is fully aerated under these
conditions (i.e., DO is high, and the DO demand of the ore does not outweigh oxygen
delivery to the pulp). This is evidenced by the small increase in Eh from the plant cyclone
overflow to rougher cell 9 in Survey 1 (Figure 16).
In Survey 2, sodium hydrosulfide, or NaHS (~100 g/t), was added to the plant cyclone
overflow to an Eh of –50 mV (Orwe, Grano, and Lauder 1998). This decreased the Eh of
the cyclone overflow (as shown in Figure 16, Survey 2) into the stability region of copper
sulfide (Cu2S). Oxidation would be prevented and oxidation products on chalcocite (copper hydroxide) would be reduced to a Cu-S phase (Orwe, Grano, and Lauder 1997). With
aeration in the plant flotation cells, the Eh increased to +150 mV (Figure 16, Survey 2, cell 9),
which is probably in an acceptable region for collector adsorption onto chalcocite (Richardson and Walker 1985). Copper recovery increased with this treatment though the extent of
this increase depended on copper mineralogy (chalcocite oxidizes more easily than chalcopyrite).
In general, sulfide skarn ores exhibit a much higher DO demand in the plant rougher
cells than do porphyry ores (Figure 17). The DO demand is a function of the surface oxidation state and exposure of sulfide minerals. The high DO demand of sulfide skarn pulps is
related to their high pyrite content. Through processing in the plant, the pyrite surfaces
become progressively more oxidized, consuming less oxygen. The copper rougher recovery
during surveys on the high pyrite-skarn blends varied widely, between 60% and 80%,
whereas the copper rougher recovery during surveys conducted on the porphyry ore ranged
from 80% to 85%. The rapid passivation of the sulfide minerals—most probably pyrite—at
alkaline pH values and under laboratory grinding and flotation conditions was demonstrated in other tests. High pyrite-skarn feeds require longer aeration times to achieve complete oxidation, though grinding under alkaline conditions is critical for pyrite passivation.
A
B
100
Cell 15 Tail
Dissolved Oxygen, %
80
Cell 9 Tail
Cell 3 Tail
60
Cell 15 Tail
Cell 9 Tail
Rougher
Feed
40
Cell 3 Tail
Rougher
Feed
20
0
0
30
60
90 120
0
30
60
90 120
Time, sec
FIGURE 17 DO concentration as a function of time (DO demand) after the discontinuation of
air purging for plant rougher streams in surveys on (a) porphyry ore blends (head grade: 0.6%
Cu, 1.2% S, Cu–S 0.5) and (b) sulfide skarn blends (head grade: 1.2% Cu, 4.0% S, Cu–S 0.3)
flotation0.book Page 253 Tuesday, January 2, 2007 7:36 PM
PULP AND SOLUTION CHEMISTRY
253
In terms of plant practice, it was recommended that lime be added to the primary ball
mill feed, as well as to the SAG feed, with collector addition after the grinding circuit
(Putubu, Grano, and Morey 2001). The objective was to passivate pyrite prior to collector
addition, so that most of the collector would be consumed by chalcopyrite rather than
pyrite. It was also recommended that stage-adding collector be tested, not only for coarse
composite recovery but also for pyrite control. The effect from using the first cell as an aerator
to passivate pyrite prior to collector addition needed to be investigated for high pyrite-skarn
feeds. When conditions were established that ensured effective pyrite depression, collector
additions could then be increased in accordance with the higher copper grades of the sulfide
skarn ores. DO demand and EDTA extraction should then be used to monitor pyrite oxidation and copper activation, and the data correlated with plant performance. Furthermore,
laboratory tests should continue to focus on correlating laboratory conditions with the
plant, and the effect of Eh that is controlled by gases in the mill and conditioning stages
should also be studied. Finally, the possibility of increasing the Eh from +170 mV to +230 mV
during the treatment of high pyrite-sulfide skarn blends was recommended as a means of
passivating pyrite.
S U M M A RY
Pulp and solution chemistry have a pivotal role in bubble–particle capture. Surface heterogeneity, degree of surface hydrophobicity, degree of oxidation, bubble nucleation, and
related phenomena influence the stability of the aqueous wetting film between a particle
and an approaching bubble. For metal sulfide minerals that oxidize relatively easily in many
cases, monitoring and control of oxidation is often the key to selective flotation. Surface
spectroscopy, electrical double-layer studies, and cyclic voltammetry collectively enable various mechanisms to be understood. The application of this information to several sulfide ore
operations has been discussed, demonstrating how rather straightforward tests can be used
to diagnose and correct problems in operating plants. Oxidation product dissolution and
re-precipitation, as well as metal ion hydrolysis formation and adsorption, are central issues.
The use of pulp and solution chemistry as diagnostic tools in selective flotation has been
examined in several case studies, thereby integrating theory and practice.
AC K N OW L E D G M E N T S
This work has been supported by the Australian Research Council Special Research Centre
Scheme and Amira International, as well as the University of South Australia (Adelaide).
Beneficial discussions with Stanislav Dukhin, Nataliya Mishchuk, Hans Schulze, Roger
Smart, Bill Johnson, Russell Schumann, and George Levay are warmly acknowledged.
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flotation0.book Page 259 Tuesday, January 2, 2007 7:36 PM
The Physics and Chemistry
of Frothers
Robert J. Pugh
A B S T R AC T
This chapter discusses several important physicochemical aspects of frothers. Initially, the relationship among hydrophilic/lyphophilic balance, molecular weight, foaming performance, and
solubility of polypropylene glycol (PPG) frothers is discussed. The decrease in frother performance, which occurs at relatively low concentrations for high-molecular-weight PPG frothers,
emphasizes the importance of dosage control as a key operating factor in flotation performance.
Recent important literature studies on surface tension, drainage, particle entrainment, and synergistic foaming of mixtures are presented, together with the effects of hydrophobic particles on
the structure and stability of three-phase froths. Finally, frothing and flotation in the absence of
frother and collector (collectorless flotation) and foaming in aqueous solution of electrolytes are
briefly reviewed.
INTRODUCTION
In the flotation process, particles are transferred from the collection zone of the cell to the
froth phase, and the efficiency of this transfer process is defined as the froth recovery. Froth
decay (bubble coalescence) reduces the bubble surface area flux across the froth phase and
increases the particle detachment rate. Overall, the recovery and grade of concentrate are
strongly influenced by the structure of the froth, which can be modified by the types and
concentrations of chemicals used in the processing. One of the important classes of chemicals used as a frother is usually a neutral molecule, which adsorbs at the water–air interface,
aids in the production of bubbles, and stabilizes the flotation froths. They have limited solubility in water, which decreases with the increases in hydrophobicity and molecular weight.
Frothers can be essentially divided into four chemical groups. The first group consists of
aromatic alcohols, such as α-cresol and 2,3-xylenol. A second group is the alkoxy types, such
as triethoxy butane (TEB). The third group consists of aliphatic alcohols, such as 2-ethyl
hexanol, diacetone, and methyl isobutyl carbinol (MIBC). Overall, MIBC is the most commonly used single frother today and can be preblended with other chemicals. It is a versatile
frother, relatively inexpensive, and gives good performance with a range of different ores.
However, there has been environmental concern with regard to MIBC’s low flash-point
temperature and high vaporization rate that produces an unpleasant odor in warmer climates. It is, therefore, frequently necessary to find a suitable replacement for MIBC, and
several types of aliphatic alcohols have shown promise.
In recent years, a fourth important group of synthetic frothers consisting of PEO (polyethylene oxide), PPO (polypropylene oxide), and PBO (polybutylene oxide) types has been
introduced into the market. These chemicals may be represented by the general equation:
R(X)n OH, where R = H or CnH2n+1 and X = EO, PO, or BO (ethylene oxide, propylene
259
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260
FLOTATION FUNDAMENTALS
oxide, or butylene oxide, respectively). They are essentially derivatives of ethylene or propylene oxide and were manufactured by Dow (Dowfroth products) and Union Carbide
(polypropylene glycol [PPG] frothers). The relative length of the hydrophobic to hydrophilic ends is controlled by changing the number of –CH2– groups in the alkyl ether and
EO –CH2CH2O–, PO –CH2 (CH3) CH O–, or BO –CH2CH2 (CH3) CH–O– groups
in the PEO chain. The propylene and butylene groups are hydrophobic, and the ether oxygen and hydroxyl represent hydrophilic groups. These surfactants represent the most flexible group of neutral frothers and are possibly the second major class of commercial frothers
used today. They are extremely stable throughout a wide range of pH and can be used in
alkali flotation. Together with MIBC, the polyglycol ethers account for about 80%–90% of
all frothers used in metallic ore flotation.
Different chemical structures give dissimilar flotation rates with variations in recoveries
of the mineral particles. The molecular weight of the frothers is very important with regard
to the flotation kinetics, and low-molecular-weight frothers have high diffusion coefficients
and increase the flotation kinetics. Usually, higher-molecular-weight frothers are referred to
as “more powerful frothers,” and the stronger frothers types are Dowfax 400 and Dowfax
1400. Research by Klimpel and co-workers (Klimpel and Hansen 1988; Klimpel and Isherwood 1991)showed that the main advantages of the polyglycols are that they exhibit a high
degree of selectivity with well-defined chemical structures and composition, which could
give more enhanced performance. Several new families with well-defined chemical structures were invented for coals and have now been commercialized. The relationship between
frother molecular structure, gas–solution interfacial properties, and foaming is in need of
further investigation.
T H E H Y D R O P H I L I C / LY P H O P H I L I C B A L A N C E
After discovery of the striking relationship between ethylene oxide content and emulsion
stability, the hydrophilic/lyphophilic balance (HLB) classification of surfactants, as developed by the Atlas chemical industries, was originally intended for nonionic ethoxylated surfactants with the HLB value defining their emulsification performance. Later, the concept
was extended to ionic surfactants, and when the composition of the surfactant became
known, the HLB could be calculated. In the original theory, the HLB is related to the
molecular weight percentage of the hydrophilic part of the surfactant molecule.
The central assumption is based on the idea of an optimum interfacial stability that can
be defined by critical HLB value, which is dependent on the degrees of hydrophilicity and
hydrophobicity of the two phases. This optimum value can lead to maximum performance,
but often this can only be achieved by mixing two surfactants in predetermined proportions.
Using the well-known equation derived by Davies, discussed by Becher (1984) and
shown below in Equation 1, the HLB values of the frothers have been frequently calculated.
HLB = 7 + ∑ ( hydrophilic group number ) – ∑ ( lipophilic group number )
(EQ 1)
By varying the ratio of the hydrophilic groups to hydrophobic groups and the molecular
weight during polymerization, the HLB number and solubility of the synthetic frothers in
water is changed. Recent studies with these synthetic frothers show that the performance of
frothers is extremely dependent on HLB and also chemical structure (Pugh 2000). Klimpel
and co-workers (Klimpel and Hansen 1988; Klimpel and Isherwood 1991) carried out
extensive development work in the plant to study the effect of frother structure on the
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261
selectivity and recovery of various minerals. They found that with increased branching, the
maximum particle size of recoverable mineral decreases while the selectivity increases. It has
also been reported in the literature that those frothers with the same HLB but with larger
molecular weights give froths that are more persistent (Aston et al. 1983).
However, more recent studies have shown that although members of the same series of
surfactants could be usefully classified chemically according to the HLB value, this is only
valid on an intraseries basis. In the case of a different series of surfactants, there was no relationship between HLB of one series with that of another series in terms of their physical or
chemical properties. For example, for the PPGs, the HLB values can be compared within
the same family, but these HLB values bear no relationship with the HLB values of another
series of frothers. To make useful comparisons of different frothers, the chemical structure of
the frothers in addition to the HLB need to be quantified.
T H E S O L U B I L I T Y A N D S U R FA C E T E N S I O N O F F R O T H E R S
Several researchers have classified polyglycol ether as completely water soluble or completely
miscible with water. When these synthetic frothers came onto the market, many mill metallurgists assumed that good frothers could be only slightly soluble in water. However, there
are few in-depth studies that have reported about the influence of frother solubility (phase
separation) and surface tension on performance. A recent study by Tan and co-workers
(2004) reported on the relationship among foaming, solubility, surface tension, and molecular weight (obtained by reverse-phase high-performance chromatography, or HPC) of four
members of homogeneous series of PPGs. This data indicated a fairly narrow and distinct
distribution for the low-molecular-weight polymers (PPG 192), but a broader distribution
for the higher-molecular-weight polymers (PPG 400, PPG 1000, and PPG 2000), as indicated in Figure 1.
1.0
PPG 192
0.9
0.8
Relative Amount
0.7
0.6
0.5
0.4
0.3
PPG 400
PPG 1000
PPG 2000
0.2
0.1
0.0
0
10
20
30
40
50
Chain Length
Source: Tan et al. 2004.
FIGURE 1
Molecular weight distribution of a series of PPG frothers as determined by HPC
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262
FLOTATION FUNDAMENTALS
In Figure 2a, the equilibrium surface tension (measured by the Wilhelmy method) versus concentration of the five PPG frothers is compared to MIBC. The plots clearly show
that the higher-molecular-weight polymers are more surface active, reducing the surface tension more effectively at lower concentrations and in following the order: PPG 2000 > PPG
1000 > PPG 725 > PPG 400 > PPG 192. Considering the difference in molecular weight,
75
PPG 192
PPG 400
PPG 725
PPG 1000
PPG 2000
MIBC
Equilibrium Surface Tension, mN/m
70
65
60
55
50
45
40
35
30
0.0001
0.01
1
100
10,000
Concentration, mM
A. Equilibrium surface tension vs. concentration of PPG and MIBC.
The arrows indicate where droplets were observed in the solution
(phase separation).
75
PPG 192
PPG 400
PPG 725
PPG 1000
PPG 2000
Equilibrium Surface Tension, mN/m
70
65
60
55
50
Droplet
Formation
45
40
35
30
1E-06
0.0001
0.01
1
100
10,000
C/C*
B. Equilibrium surface tension vs. concentration of PPG relative to the
phase separation concentration.
Source: Tan et al. 2005.
FIGURE 2
Surface tension vs. concentration
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THE PHYSICS AND CHEMISTRY OF FROTHERS
263
MIBC was slightly more surface active than PPG 192. For the PPG series, this increase in
surface activity follows the increasing number of hydrophobic groups on the polymers.
In the plots for Figure 2a, there was no evidence of a critical micelle concentration
(CMC), which would have resulted in a breaking point in the curve and a plateau. However,
in all solutions, phase separation occurred in the higher concentration ranges, causing the
solution to become turbid. The earliest stage of this phenomenon resulted in droplet formation, which was detected by dynamic light scattering on the prefiltered solution as previously reported (Tan et al. 2004). The critical concentration of the phase separation (CPS,
C*) is indicated by arrows in Figure 2a.
To enable a comparison between the different CPS values of the various polymers, the
results were normalized, and the surface tension versus concentration data were plotted in
the form of surface tension versus C/C* (Figure 2b). This plot shows that the C/C* values
coincide, indicating that phase separation occurs at the same relative concentration for the
different PPG polymers. This value corresponds to a solution surface tension value in the
range of 42 to 48 mN/m.
The average HLB and the range of HLB values are shown in Figure 3 for the PPG polymers together with the value of MIBC for comparison purposes. The results show a
decrease in the solubility limit and HLB value with an increase in molecular weight, with
the average HLB value falling from 10.5 to about 6.5 for the PPG polymers. The HLB number of MIBC is indicated.
The frothing performance of the chemicals was also evaluated at a 20-ppm concentration
level, and the intermediate molecular weight frother (PPG 400) with HLB performed better.
1E+0
PPG
MIBC
1E–1
C*(M)
1E–2
1E–3
1E–4
1E–5
2
4
6
8
10
12
Hydrophilic/Lyphophilic Balance
Source: Tan et al. 2005.
FIGURE 3 The phase separation concentration, C*, versus the average HLB and range of HLB
for the PPG. The HLB of MIBC is indicated.
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FLOTATION FUNDAMENTALS
R E L AT I O N S H I P B E T W E E N F R O T H I N G A N D P H A S E
S E PA R AT I O N O F F R O T H E R S
Additional studies of foamability measurements of the PPGs were conducted by Tan et al.
(2004, 2005) over a wide range of concentrations and gas flow rates. They reported that the
foam height (or foamability) followed a characteristic plot (Figure 4), which could be separated into three distinct regions. At low concentrations, foamability increases and reaches a
critical foam height at concentration C1. At this point, the foamability reaches a maximum
value and remains constant until another point, C2, is reached. At concentrations greater
than C2, the foamability decreases. The values of C1 and C2 decreased with increasing PPG
molecular weight. Clearly, these results demonstrate that each foamer has a well-defined
concentration range for foaming and defoaming. The results for the PPG series of frothers
are shown in Figure 5 and show characteristic plots over different concentration ranges.
In the low-concentration range (where foaming increased following an increase with
concentration), the amount of surfactant adsorbed at the air–solution interface increased
because of an increase in surface pressure. In the plateau concentration region, although the
concentration of surfactant in solution is increased, the foaming characteristics of the system
remain constant. Finally, a critical point is reached at a higher concentration where the
foamability begins to decrease.
The plateau region is likely to be related to the influence of bulk surfactant on the
Marangoni effect. Surface tension gradients build up at the air–solution interface, resulting from thinning of the film. This causes the transport of bulk liquid into the thin film,
restoring the thickness and preventing rupture from occurring. Generally, the plateau
concentration C1 is found to decrease for molecules with increasing surface activity,
because a more-surface-active molecule adsorbs more readily at a gas–liquid interface. At
concentrations beyond a critical limit, C2, the surfactant exceeds its solubility limit and
Constant
Decreasing
Foamability
Increasing
C2
C1
Concentration
FIGURE 4
Foamability versus concentration profile of PPG frothers
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THE PHYSICS AND CHEMISTRY OF FROTHERS
265
phase separation occurs as droplet formation. In this high-concentration region, the droplets act as an antifoamer.
Many years ago, this foam depression effect caused by excessive frother was recognized
as “over oiling.” In fact, increasing the frother dosage caused a rise, then a leveling-off, and
finally, a drop in recovery of minerals. Curiously, the effect seems to be almost forgotten
among researchers today, although Klimpel (1987) studied over-oiling by frothers on an
industrial scale. As demonstrated in the previous section, the wide variation in concentration,
which results in this depressant effect depending on the molecular weight of the PPG,
emphasizes the importance of frother dosage changes as a key factor in flotation operations.
Gas Flow Rates
33 mL/min
54 mL/min
43 mL/min
65 mL/min
Parts per Mllion
1.9
25
19
192
1,920
Parts per Million
19,200
A
20
15
10
5
C1
0
0.01
0.1
1
42
10
5
C2
10
100
0
0.01
1,000
0.1
Steady-State Foam Height, mm
Steady-State Foam Height, mm
1.9
25
15
10
5
Concentration, mM
19.4
PPG 2000
20
0.1
1.0
194
D
20
15
10
5
0
0.001
0.01
Concentration, mM
Source: Tan et al. 2004.
FIGURE 5
1,000
Parts per Million
940
C
0.01
10
Concentration, mM
94
PPG 1000
0
0.001
B
15
Parts per Million
9.4
420,000
20
Concentration, mM
0.94
25
4,200
PPG 400
Steady-State Foam Height, mm
Steady-State Foam Height, mm
PPG 192
0.42
25
Foamability versus concentration profile for PPG frothers
0.1
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FLOTATION FUNDAMENTALS
A D S O R P T I O N O F F R O T H E R S AT T H E M I N E R A L – S O L U T I O N
A N D A I R – S O L U T I O N I N T E R FA C E S
Ideally, in flotation circuits the frother preferentially adsorbs only at the air–solution interface. However, in the case of excessive overdosing of frother where phase separation occurs,
the droplets can adsorb at the mineral–solution interface. Interestingly, the adsorption of
MIBC and kerosene-type frothers on hydrophobic solids such as coal has been reported in
the literature (Aktas and Woodburn 1994), but this only appears to occur at high concentrations (greater than C2). It was suggested that the adsorption is influenced by the internal
pore structure of the mineral. In fact, the rival effects of pore penetration and surface spreading of the frother droplets on the surface of the hydrophobic solid have been a controversial
topic of debate. For coal particles, the flotation yield using short-chain hydrocarbon frothers
was retarded at high concentrations by the penetration of the surfactants, such as diacetone
alcohol and 2-ethyl hexanol, into the pores of coal.
Early studies by Leja and Nixon (1957) also suggested that the interaction between collector and frother could lead to a complex, which, upon adsorption onto the mineral, could
remove frother from solution. Some form of collector-frother adduct was suggested by Crozier and Klimpel (1989) for the xanthate-alcoholic frother on the surface of sulfide minerals. However, to date, there has been no follow-up, and this work has yet to be verified.
S U R FA C E T E N S I O N A N D B U B B L E S I Z E
Typically, frothers are used in the very low dosage range, and at these concentrations they
cause very limited reduction in the surface tension of the solution. Generally, bubble size is
much more sensitive to very low frother concentrations than the surface tension, but few
study results have been reported in these low frother concentration ranges.
The air-bubble size dependence on the surface tension of the liquid has been studied by
Rao and Stenius (1998) for a series of frothers. They measured the population size and distribution of bubbles (using a laser light-scattering technique) formed in solutions of nonionic alkyl PEO frothers and a long-chain anionic surfactant. These co-workers showed that
as the surface tension of the liquid decreased, the average maximum size of the bubble population decreased, and the distribution became narrower. In addition, larger numbers of
smaller bubbles were produced at increased surfactant concentrations. The experimental
data reported from these studies are shown in Figures 6 and 7. In Figure 8, the surface tension of the solution was related directly to the bubble size in agreement with the Laplace
equation for the nonionic frother systems.
The coalescence of bubbles is also highly dependent on the electrolytes in the solution.
The flotation efficiency increases as bubble size decreases. Because of streamlines around a
rising air bubble, small particles frequently do not collide with the bubble, which is much
larger than the particles. In addition, the stability of bubbles and foaming characteristics are
dependent on the combination of both equilibrium and dynamic effects, with the dynamic
effects dominating. For strong foaming, the literature concludes that the surfactant must be
capable of rapidly lowering the surface tension. However, a relatively slow rate process is also
required by which a freshly created liquid surface retains high, non-equilibrium surface tension long enough for surface flow to occur to stabilize the film. Mere rapid reduction in surface tension does not lead to the stabilization of the foam. What is necessary is the slow
attainment of equilibrium after a fresh surface is produced.
Surface elasticity arises from the variation of the surface tension during deformation of
a liquid film. This can be manifested under equilibrium conditions (when the surface layer
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267
300
Surface Tensions, mN/m
C11E7
C11E10
C12SO3
Water
28.44
250
Number of Bubbles
32.28
200
35.08
150
100
71.81
50
0
15
25
35
45
55
65
Diameter, μm
Source: Rao and Stenius 1998.
FIGURE 6
Bubble size distribution in water and nonionic surfactant solution close to CMC
20
Surface Tensions, mN/m
30.6
55.5
71.8
Bubble Population, %
16
12
8
4
0
20
30
40
50
60
70
Bubble Diameter, μm
Source: Rao and Stenius 1998.
FIGURE 7
Bubble size distribution in water containing C11E10 at different concentrations
50
Maximum Bubble Diameter, μm
46
42
38
34
30
30
40
50
60
70
80
Surface Tensions, mN/m
Source: Rao and Stenius 1998.
FIGURE 8
Relationship between bubble size and surface tension for C11E10
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268
FLOTATION FUNDAMENTALS
is under tension during equilibrium with the bulk phase), and this is defined by the Gibbs
elasticity. In non-equilibrium conditions, it is defined by the Marangoni elasticity. The
Marangoni elasticity of the monolayer can be determined from the dynamic (non-equilibrium)
surface tension as the surface is abruptly extended or pulsated. The Marangoni elasticity is
usually larger than the Gibbs elasticity for the same system and is a more important parameter in foam stability. The Marangoni elasticity occurs at a range of frequencies, but it can be
evaluated from the dynamic surface tension data, usually at low frequency of dilation
(compression). In practical foam systems, this could be as low as 1 cycle/min or as high as
900 cycles/sec. Essentially, it depends on the extension–contraction cycle in the foam. In
fact, measurements (using a suitable experimental technique) ideally need to be made to correspond with measurements of the change in surface area of the actual bubbles during foaming.
T H E D R A I N AG E O F F ROT H E R S O L U T I O N
Generally, it has become clear from model studies that the same factors which play a role in
foam stability (film thickness, elasticity, etc.) also have a decisive influence on the stability of
the isolated thin films. Hence, model film studies are very important, and the drainage of
thin foam films will be considered in the next section. Experimental thinning studies have
been reported with a microfilm formed from a concave liquid drop suspended in a short,
vertical capillary tube. The apparatus was originally developed by Schedludko (1966) and
has been well documented in the literature. Direct measurement of the thickness of the aqueous film versus time (the drainage process) can be determined by micro-reflectance methods.
The drainage time of a flat film is determined from the Reynolds equation:
ho
T =
∫ dh ⁄ VRe
(EQ 2)
ht
where
T = the time of drainage of the film, which has an initial thickness of ho and
drains tothe thickness ht
h = the distance between surfaces
VRe = the Reynolds thinning velocity
t = time
For the case of horizontal, fairly thick films (>100 nm), an expression has been derived
for the thinning between two disc surfaces under the influence of a uniform external pressure. The change in film thickness with drainage time can be expressed as the Reynolds drainage:
dh2h 3 ΔP
V Re = –-------= ---------------dt
3ηR 2
(EQ 3)
where
R = the radius of the disc
η = the viscosity of the liquid
ΔP = the pressure difference between the film and the bulk solution and is taken to
be equal the capillary suction Pc in the surrounding “plateau” border
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269
Figure 9 shows the drainage patterns of films produced with three different frothers: ethoxylated nonionic—C10(EO)5 and C12(EO)5—and Dowfroth 200, which is a PPG with a
molecular weight of about 200. In all cases, the results—expressed as the experiment-measured
drainage rates (V) as compared to VRe versus bulk concentration (see Figure 9)—indicate
that the films thin faster (i.e., drainage times are shorter) than those calculated from the
Reynolds equation (Equation 3). The drainage is of the same order of magnitude as in previous results and follows the order: Dowfroth 200 > C10(EO)5 > C12(EO)5. The increase in
frother concentration has a pronounced effect on drainage but only at higher concentrations where drainage is considerably reduced. Polyethylene glycols at low concentrations are
well-known drag-reducing agents, and high drainage velocities are expected.
2.0 2.0
2.0 2.0
V/VRe
V/VRe
1.5 1.5
1.5 1.5
Dowfroth
Dowfroth
C10(EO)5
C10(EO)
5
C12(EO)5
C (EO)
12
5
1.0 1.0
1.0 1.0
0.5 0.5
0.5 0.5
τ/ττ/τ
Re
0.0 0.0
10–610
–6
10–410
–4
10–210
–2
Re
0.0 0.0
Bulk
Concentration,
Bulk
Concentration,
M M
Source: Manev and Pugh 1992.
FIGURE 9 The relative thinning ratios of C10(EO)5 and C12(EO)5 and the ratio of measured
drainage times (V to the calculated VRe) are compared to Dowfroth 200. Below the CMC the
results are similar, but above the CMC the ethoxylated frothers drain slower than the Dowfroth
200.
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FLOTATION FUNDAMENTALS
25
25
B
20
Foam Height, mm
20
Foam Height, mm
A
PPG 1000
MIBC
15
10
15
10
5
5
0
0
Mixture PPG 1000/MIBC
0
10
20
30
40
50
0
Concentration, ppm
10
20
30
40
50
PPG 1000 Concentration in Mixture, ppm
Source: Tan et al. 2005.
FIGURE 10 Foam behavior of (a) PPG 1000 and MIBC from 0 to 5 ppm, and (b) PPG 1000/MIBC
mixture at a total concentration of 50 ppm
M I X E D C H E M I C A L F ROT H E R S Y S T E M S
In the flotation of minerals, a frother or combination of frothers is chosen depending on the
requirements of the mechanical and interfacial rheology characteristics of the froth. In addition to technical advantages, there are also economical advantages in using blends. Often in
sulfide flotation, two or more frothers will be blended to improve flotation efficiency. Frequently, a short-chain alcohol such as MIBC is chosen as a versatile frother together with a
higher-molecular-weight frother, such as pine oil or a PPG type, which leads to improvements of the froth or bubble size, especially when coarser particles must be floated.
Figure 10a shows the foam height versus concentration at a constant specific gas volume for a series of individual and mixtures of frothers. In this figure, the froth heights for
the individual frothers, PPG 1000 (HLB 8.4) and MIBC (HLB 6.1), are shown for a concentration level from 0 to 50 ppm. These plots clearly show that for the single-frother systems, PPG 1000 performs better than MIBC with the foam height increasing with frother
concentration, whereas with MIBC the foam height remains relatively low throughout the
concentration range. In Figure 10b, the foam height for a mixture of PPG and MIBC is shown
with a total frother concentration of 50 ppm. These results show that the mixed system with
lower amounts of PPG 400 (5 ppm) but higher amounts of MIBC (45 ppm) gives a better
foaming performance than that of MIBC or PPG 1000 alone at 50 ppm.
Additional experiments have shown that a mixed surfactant film, consisting of a lowmolecular-weight/high-HLB frother mixed with a high-molecular-weight/low-HLB
frother, results in improved foaming compared to either frother alone.
In Figure 11, configurations are suggested for the low-, high-, and mixed-PPG polymers
at the air–solution interface. In the mixed systems, these structures suggest that a closed,
packed, cohesive film at the air–solution interface can be achieved, which could explain the
higher foaming.
S TA B I L I Z AT I O N O F B U B B L E A N D F OA M B Y PA R T I C L E S
The terms foaming and frothing are used interchangeably, but it is more usual in mineral processing to refer to the gas–water macro-cluster systems (two-phase) where the broken structure leaves a homogeneous aqueous phase as a foam. In mineral processing, the froth
contains dispersed solid particles and is a three-phase system, so that the broken structure
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THE PHYSICS AND CHEMISTRY OF FROTHERS
A
271
B
Air
Air
Water
Water
Low-Molecular-Weight/
High-HLB Polymer
C
Medium-Molecular-Weight/
Medium-HLB Polymer
D
Air
Air
Water
Water
High-Molecular-Weight/
Low-HLB Polymer
Mixed Film Consisting of
Low-Molecular-Weight (High-HLB)
and High-Molecular-Weight (LowHLB) Polymers
Source: Tan et al. 2005.
FIGURE 11
Suggested structures of PPG polymers and mixtures at the air–solution interface
gives a two-phase system (aqueous solution and finely divided particles). In mineralized
froths, the important parameter would appear to be the wettability of the particles, but size
and shape also have some influence. Although it has been clearly established that foam stability can be increased or decreased by many types of particles, to some extent the mechanism is complex because, frequently, there are several mechanisms operating in the same
system. Usually, the particles have some critical degree of hydrophobicity, which plays a critical role in the dynamics of rupturing the thin foam films. Both the particle size and shape
have been shown to play an important role, and systems have been studied using particles
within a wide size range ( Johansson and Pugh 1992; Dippenaar 1982).
It must be also stressed that throughout frothing technology, the main problems are frequently related to both stabilization and destabilization of the system, and the role of the
particles is crucial in both functions. In fact, the use of particles as foam breakers is well
known throughout industry, and hydrophobic particles are important ingredients in many
foam-breaking formulations (Pugh 2002a).
F R O T H I N T H E F L O TAT I O N P R O C E S S
Overall, the dynamic behavior of particles and the interaction with frothers are critical, but
steps in industrial froth flotation that are poorly understood usually lead to loss of recovery
during the froth phase. Although foams are stabilized by adsorbed surfactant molecules
(Figure 12a), froths are usually stabilized by small particles with a critical degree of wetting
at the gas–liquid interface, which causes the bubbles to become “armored” (Figure 12b).
Several early ideas were based on the premise that coalescence is prevented because of a steric
interaction.
As reported by many flotation researchers, the stability and drainage of a three-phase
froth depends on hydrophobicity of the mineral particles present in the froth. Froth
becomes stabilized and draining of the liquid from the thin layer is restricted by hydrophobic solid particles. The more strongly the particles adhere to the bubble, the more stable the
froth becomes; thus, an increase in contact angle of particles to a certain critical value benefits froth stability. However, there is considerable evidence that the highly hydrophobic particles (e.g., with contact angle greater than 90°) will destabilize froth (Garrett 1993).
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FLOTATION FUNDAMENTALS
A
FIGURE 12
B
Different stabilizing actions of (a) frother molecules and (b) solid particles
A. Nonwetting particles cause receding of the liquid and premature rupture of
the film at liquid–solid interface, as indicated by arrows.
B. Wetting liquid allows particles to retain liquid within the film, delaying rupture.
Source: Ip, Wang, and Toguri 1999.
FIGURE 13 The influence of wettability on film stability as the film approaches the particle size
thickness
C A P I L L A R Y E F F E C T S O N F R O T H S TA B I L I T Y
As the liquid film drains down to a critical thickness, the nonwetted particles can reduce
froth stability by inducing the liquid to de-wet around the particle, causing the liquid to
recede from the particle at areas indicated by the arrows in Figure 13a. This leads to rapid
rupture, but in the case of partial wetting, the particles trap the liquid, making the film more
stable (Figure 13b).
However, the influence of particle concentration, density, and shape needs to be taken
into consideration. As shown in Figure 14, the effect of low and high concentrations of nonwetting particles and the effect of plate-like particles (for example, clay particles) on thin
film stability are illustrated.
Two mechanisms have been suggested to explain the froth stabilization effects caused
by hydrophobic particles adsorbed at the interface. The first effect results from a change in
capillary pressure. This is caused by the presence of adsorbed particles modifying the curvature of the gas–liquid interface, which reduces the pressure difference between the plateau
borders and the three films associated with it. The situation is illustrated in Figure 15.
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THE PHYSICS AND CHEMISTRY OF FROTHERS
273
Air
Water
Air
B. High Mineralization of Bubbles
A. Poor Mineralization of Bubbles
C. Effect of Plate-like Particles
Source: Warren 1981.
FIGURE 14
Effect of hydrophobic particles on bubble stability
Without Particles
A
With Particles
D
Film
Profile
View
Plateau
Border
B
E
Pgas
Pgas
Pfilm
Pfilm
Pgas
Pgas
Ppb
Ppb
Pgas = Pfilm > Ppb
Pgas > Pfilm ≈ Ppb
Near Particles
Film
C
F
Plateau
Border
Pressure
Source: Ip, Wang, and Toguri 1999.
FIGURE 15 The effect of hydrophobic particles on the pressure difference between the foam
film and the plateau borders
In the case of the foam with no particles (Figure 15a), the liquid can flow from the film
into the plateau border and then flow through the structure by gravity.
The flow rate would be proportional to the pressure difference, ΔP, expressed by
ΔP = P film – P pb
= γ ⁄ R pb
(EQ 4)
where Pfilm is the pressure in the films, Ppb is the pressure in the plateau border, γ is the surface tension of the liquid, and Rpb is the radius of curvature of the gas–liquid film interface
(Figures 15b and 15c). Therefore, when ΔP is high, the flow rate is increased, which causes
faster drainage, and the foam becomes less stable. If many hydrophobic particles are
attached to the gas–liquid interface (Figure 15d), the radius of curvature of the gas–film
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274
FLOTATION FUNDAMENTALS
interface would change and become almost equal to that of the gas–plateau border interface
(Figures 15e and 15f ). This will cause the pressure difference to decrease, resulting in a more
stable froth.
According to Lucassen (1992), capillary effects are also very important for small particles where gravitational forces are negligible. This results from small floating solid particles
at the fluid interface that can interact with each neighbor because any solid particle will virtually always cause deformation of interfaces. The extent of deformation should be altered
by the approach of the neighboring particle. This especially occurs in cases where particles
have an irregular wetting perimeter, which disturbs the smoothness of the interface. Calculations made for a model particle with a sinusoidal edge indicate that the disturbances can
become significant when induced by capillary interaction with neighboring particles.
Restricted Drainage Mechanism
It has also been suggested that attached particles can cause the overall drainage within the
thin film to be hindered as the film and the liquid passages become constricted and tortuous. To some extent, the volume of the particles stabilized by froth is, therefore, roughly proportional to the amount of hydrophobic solids present, but there is an upper limit. It
appears that if the size of the hydrophobic particle is small enough compared with the film
thickness, as previously discussed, then they can arrange themselves at the liquid–gas interface and stabilize the films by the capillary mechanism described. If the particles are large
(i.e., the diameter is larger than the film thickness), the particles can bridge and rupture the
foam film. Frothing experiments by Garrett (1993) using fairly large particles (1–100
microns) showed that these particles could easily destroy the foam.
This effect has been observed in the flotation process. For instance, it has been reported
that 0.1-mm galena particles can prolong the life of froths of isoamyl alcohol aqueous solutions from 17 seconds to several hours; whereas 0.3-mm galena particles can only increase it
to 60 seconds (Lu, Pugh, and Forssberg 2005).
Frothing Studies with Model Quartz Particles of Well-defined Size and
Hydrophobicity
This section studies the influence of size and hydrophobicity of particles on the stability of
froths using both the modified Bikermann column and the thin film balance ( Johansson
and Pugh 1992). In these experiments, surface-modified quartz particles were used as models. The hydrophobicity of the quartz surface was controlled by reacting the dry surface with
trimethylchlorosilane in cyclohexane under a dry environment according to standard procedures. In order to characterize the surfaces, a quartz plate was also placed in the reaction vessel together with the particles. After the reaction, the plate and particles were removed and
rinsed with cyclohexane and washed with water. Finally, the contact angle of a drop of water
on the plate was determined using the Ramé-Hart goniometer. The hydrophobicity of the
quartz particles was quantified from small-scale flotation experiments using a Hallimond
tube apparatus.
The froths (containing quartz particles) were characterized by both dynamic and static
frothing tests. The dynamic test (carried out during froth generation) quantified the equilibrium state of the froth, whereas the static tests (after the gas flow had ceased) determined the
rate of collapse of the froth. In this study, a modified Bikermann test was used consisting of a
glass column where the maximum equilibrium volume of the froth (Hmax) was determined
at a standard flow rate. A typical set of data is shown in Figure 16 for dynamic frothing with
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THE PHYSICS AND CHEMISTRY OF FROTHERS
275
Size Fraction (Frother Concentration)
26–44 μm (20 mg/L)
74–106 μm (20 mg/L)
26–44 μm (50 mg/L)
74–106 μm (50 mg/L)
12
PPGMME, n = 3
A
PPGMME, n = 4
B
12
Hmax, cm
Hmax, cm
8
8
4
4
0
0
0
20
40
60
80
0
100
20
Flotation Yield, %
40
60
80
100
Flotation Yield, %
10
α -Terpineol
MIBC
C
D
10
8
8
Hmax, cm
Hmax, cm
6
4
6
4
2
2
0
0
0
20
40
60
Flotation Yield, %
80
100
0
20
40
60
80
100
Flotation Yield, %
Source: Johansson and Pugh 1992.
FIGURE 16 Relationship between the dynamic stability expressed as the maximum froth
height (Hmax) at the flow rate of 60 L/hr vs. the hydrophobicity of the quartz particles (expressed
as the flotation yield)
four commercial mineral processing frothers: polypropylene glycol monomethyl ether
PPGMME with n = 3, PPGMME with n = 4, ∝-terpineol, and MIBC.
These results express the dynamic froth characteristics in terms of Hmax versus the
hydrophobicity of the particles expressed in terms of the flotation yield. From the results
obtained with the small particle fraction (26–44 microns), there appears to be trend for a
distinct maximum corresponding to a flotation yield of about 70%. This corresponds to a
critical degree of hydrophobicity of 60° for the small particle fraction. This value seems to be
reproducible in all the frother systems. Also, at higher flotation yield, the froth collapsed,
indicating that the particles were acting as foam breakers; therefore, it could be concluded
that below this critical degree of hydrophobicity, the particles appear to have less effect on
the system. These trends are observed at both high and low frother concentrations. However,
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FLOTATION FUNDAMENTALS
these effects were not observed with the larger size fraction. In fact, the particles do not
appear to influence the stability of the system. Similar trends have been observed at both low
(20 mg/L) and high (50 mg/L) frother concentrations. Additional experiments were carried
out at a range of frother concentrations where a similar trend was observed for Hmax values
versus concentration plots.
E N T R A I N M E N T O F S U B M I C R O N - S I Z E PA R T I C L E S I N T H E
F R O T H P H A S E A N D T R U E F L O TAT I O N
The process leading to the formation of three-phase froth layers involves not only the hydrophobic particles carried over in the froth but also the hydrophilic entrained gangue particles.
These particles are only feebly adhered to the gas bubbles or are situated in the water film in
the froth and are usually washed out during drainage (Figure 17). At the same time, the
ascending air bubbles carry the hydrophobic particles to the top of the froth. Thus, the top
of three-phase froth contains the more hydrophobic particles and has the higher grade; the
grades of the floated material decrease from top to bottom of the froth. This phenomenon is
referred to as a secondary concentration effect and is useful for upgrading the concentrate
quality and, therefore, has found application in column flotation.
In most cases, complete separation between gangue and valuable mineral in the pulp
zone is difficult to achieve and, nearly always, some gangue minerals are transported into the
froth with the entrained liquid. As the froth ages, some of the hydrophilic gangue returns to
the pulp because of slurry drainage, while the remainder is carried up with the concentrate,
reducing the quality of the product. The drainage of hydrophilic particles (or recovery of
gangue in the concentrate) is largely affected by the proportion of the gangue minerals, such
as density and particle size. To date, the effect of the gangue characteristics on the drainage
has been well established, but the influence of the froth characteristics on the gangue recovery has not been fully investigated. This is due to many factors that influence froth stability,
such as particle size, hydrophobicity, and reagents.
Recently, however, some interesting experiments have been carried out by Ata, Ahmed,
and Jameson (2004) that have helped to reveal some new aspects of this complex problem. A
specially designed flow cell was used in which the froth could be isolated from the pulp
zone; this enabled the collection of particles that had dropped out of the froth. In these
studies, hematite particles (82 microns in diameter) were floated forming the froth phase in
Air
Air
Air
Air
Source: Kelly and Spottiswood 1982.
FIGURE 17
Schematic illustration of froth drainage
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277
the conventional way, but glass particles of about the same diameter were added to the wash
water and fed to the froth to study the influence of hydrophobicity of the particles on the
recovery of gangue. The froth-fed glass particles had a range of contact angles of 0°, 50°, 66°,
and 82°. In other tests, hydrophilic and hydrophobic glass particles were mixed in various
ratios, and tests were conducted with the blended feed. This enabled studies to be conducted on the effect of the hydrophobicity of the particles attached to the bubbles and on
the drainage of gangue minerals in the froth phase.
From these studies, it was concluded that the hydrophobic particles had a strong effect
on the structure of the froth, and this had an influence on the drainage of the hydrophilic
particles. The hydrophobic particles significantly affected the drainage rate of hydrophilic
particles in the froth. The recovery of hydrophilic particles in the flotation product
increased as the hydrophobicity of the floatable particles increased, and the water recovery
rate decreased with increasing contact angle of the particles. It was concluded that the
hydrophobic particles had an effect on the structure with higher hydrophobicity, causing
increased drainage rate of water leading to drier froths in which the hydrophilic particles
become more easily entrapped. Gangue recovery in the concentrate also increased with the
concentration of the gangue mineral presented to the froth, even in the presence of wash
water, presumably as a result of the increased viscosity of the liquid in the froth.
C O L L E C T O R L E S S F L O TAT I O N
The frothing and flotation of hydrophilic metal hydroxides in the absence of frother and
collector, also known as collectorless flotation, could be considered as an area of special
interest and is sometimes referred to as contactless flotation. There are several reviews on the
microflotation of solids that occurs in the presence of hydrolyzing ions and in dissolved air
flotation circuits. The region of floatability corresponds closely with the regions where precipitation of the metal ion occurs. In fact, maximum coagulation corresponds to maximum
flotation. Hydrophilic solids, such as quartz, are usually readily coagulated by iron or aluminum, and various degrees of flotation occur with microbubbles. In these systems, it has been
suggested that entrapment of bubbles occurs in aggregates, and it has been reported that for
efficient frothing and flotation, small bubbles are required (Solari and Gochin 1992).
One theory suggests that naturally occurring organic compounds may be responsible
for particle hydrophobicity, which causes frothing in some industrial dissolved air flotation
operations. It has been well documented that natural water frequently contains biologically
derived surfactants which could stabilize microbubbles. In clean water systems, the flotation
of nonfloating suspension of ferric hydroxide flocs can occur fairly readily in the presence of
a few parts per million of collector. Also, coagulation with hydroxyl ions increases the effective
particle size and decreases the number of particles to be collected. General particle aggregation is needed, but in many cases, a collector may be also needed to improve the process efficiency. Experiments of Kitchener and Gochin (1981), have confirmed that the floatability
of metal hydroxides is very sensitive to the presence of organic impurities in the system.
They suggested natural water contains surface-active impurities that are adsorbed onto
metal precipitated hydroxides, thereby forming insoluble hydrophobic soaps that provide
sites for bubble adhesion. Because the flocs have low density, microbubble attachment to a
few hydrophobic spots on the flocs would be sufficient to ensure flotation.
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FLOTATION FUNDAMENTALS
S T R U C T U R E O F F R O T H S A N D F L O TAT I O N I N A Q U E O U S
S O L U T I O N S O F I N O R G A N I C E L E C T R O LY T E S
The frothing and flotation of naturally hydrophobic particles in inorganic electrolyte solutions was first documented in the former U.S.S.R. during the 1930s. This work was mostly
related to the flotation of coals in saline waters. Electrolytes also played a role in stabilizing
the froth. Following these early studies, many other cases have been reported in the general
area of mineral flotation. Wellham, Elber, and Yan (1992); Yoon (1982); and Yarar (1988)
have reviewed this specific area. Several surface chemical mechanisms have been proposed to
explain the flotation process. These include the action of the electrolytes in disruption of
hydration layers surrounding the particles and enhancing bubble–particle capture, reduction of the electrostatic interactions, and an increasing charge on the surface of the bubbles
to prevent primary bubble coalescence. However, none of these appear to satisfactorily
explain the experimental observations.
Yoon (1982) studied flotation coal cleaning using inorganic electrolyte, and it was concluded that the flotation efficiency increased with additional salt concentration. Flotation
rate and cleaning efficiency were also improved. More recently, Craig, Ninham, and Pashley
(1993) assessed the inhibition of bubbles to coalescence in electrolytes by the application of
a combining rule based on the nature of the cationic/anionic pair. This rule enables one to
predict whether or not the electrolyte would inhibit coalescence of gas bubbles in the electrolyte solutions. Viscosity and electrostatic repulsion were ruled out as possible explanations. In fact, following conventional electrostatic double-layer theory, an increase in salt
concentration would reduce the double-layer repulsion and should induce inhibition. It was
also suggested by Craig, Ninham, and Pashley (1993) that the coalescence in pure water was
caused by a strong hydrophobic attractive force, which opposed the hydrodynamic repulsion existing between the colliding bubbles.
Paulson and Pugh (1996) carried out flotation experiments with graphite (≈20 μm in
diameter) with a series of different inorganic electrolytes at a range of concentrations (in the
absence of an organic frother) in a small glass, cylindrical column cell. The flotation recovery of graphite as a function of the electrolyte concentration is shown in Figure 18.
The plot in Figure 18 shows that recovery generally increases with concentration and varies according to the cationic/anionic pair. In fact, it is possible to classify the electrolytes into
three groups according to their flotation performance. Group A, salts with divalent and trivalent cations or anions, include MgCl2, CaCl2, Na2SO4, MgSO4, and LaCl3, which give high
flotation response. In this group, flotation begins at about 0.02 M and reaches maximum
recovery at about 0.06 to 0.1 M concentrations. Group B includes NaCl, LiCl, KCl, CsCl, and
NH4Cl, which give medium flotation response with flotation beginning at about 0.05 to
0.1 M electrolyte. Finally, group C includes NaAc, NaClO4, HClO4, HCl, H2SO4, LiClO4,
and H3PO4, which give a very low flotation response, even up to concentrations as high as 0.3
M electrolyte. A plot of double-layer thickness versus flotation recovery (Figure 19) showed
that for group A and B electrolytes, a correlation exists between the flotation performance and
the Debye length (1/κ), which suggests that the electrostatic interaction plays a role in the process.
In this study, a relationship was also found between flotation recoveries and the surface
tension concentration gradient of the electrolyte solution. Additionally, a correlation showing a decrease in gas solubility occurred with increasing electrolyte concentration. Thus, the
increased flotation performance of the hydrophobic graphite appears to be linked with the
increase in stability of the gas bubbles and froth caused by a decrease in dissolved gas concentration in the electrolyte solutions.
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THE PHYSICS AND CHEMISTRY OF FROTHERS
279
100
90
80
A
B
Recovery, %
70
60
50
40
C
30
20
10
0
0.00
0.05
0.10
0.15
0.20
0.25
0.30
0.35
0.40
Salt Concentration, M
H 2O
MgCl2
NaCl
CaCl2
MgSO4
Na2SO4
LaCl3
LiCl
KCl
CsCl
NH4Cl
NaAc
LiClO4
NaClO4
HClO4
HCl
H2SO4
H3PO4
Source: Paulson and Pugh 1996.
FIGURE 18 The flotation recovery of graphite particles vs. the electrolyte concentration for a
series of different types of electrolytes. Group A = high flotation performance; Group B =
intermediate flotation performance; Group C = low flotation performance.
100
90
80
Recovery, %
70
60
50
40
30
20
10
0
0.0
0.5
1.0
1.5
2.0
2.5
3.0
3.5
Debye Length, mm
CaCl2
NaClO4
MgSO4
MgCl2
Na2SO4
NH4Cl
LaCl3
LiClO4
CsCl
H3PO4
KCl
H2SO4
LiCl
HCl
NaCl
HClO4
HaAc
Source: Paulson and Pugh 1996.
κ) and the flotation performance of
FIGURE 19 The relationship between the Debye length (1/κ
graphite in different inorganic electrolytes
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BIBLIOGRAPHY
Aktas, Z., and E.T. Woodburn. 1994. The adsorption behavior of nonionic reagents on two low rank
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Aston, J.R., C.J. Drummond, F.J. Scales, and T.W. Healy. 1983. Frother chemistry in fine coal
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Ata, S., N. Ahmed, and G.J. Jameson. 2004. The effect of hydrophobicity on the drainage of gangue
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Becher, P. 1984. Hydrophile-lipophile balance. J. Dispersion Sci. Technol. 5(1):81–96.
Craig, V.S.J., B. Ninham, and R.M. Pashley. 1993. The effect of electrolyte on bubble coalescence.
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Crozier, R.D., and R.R. Klimpel. 1989. Frothers: Plant practice. Pages 257–279 in Frothing in
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Dippenaar, A. 1982. The stabilization of froth by solids. Int. J. Miner. Process. 9:1–14.
Garrett, P.R. 1993. Defoaming. Surfactant Science Series, Volume 45. New York: Marcel Dekker.
Ip, S.W., Y. Wang, and J.M. Toguri. 1999. Aluminum foam stabilization solid particles. Can. Metall. Q.
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Johansson, G., and R.J. Pugh. 1992. The influence of particle size and hydrophobicity on the stability
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Kelly, E.G., and D.J. Spottiswood. 1982. Introduction to Mineral Processing. New York: WileyInterscience.
Kitchener, J., and R.J. Gochin. 1981. The mechanism of dissolved air flotation for potable water—
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Klimpel, R.R., and R.D. Hansen. 1988. Frothers. Pages 385–409 in Reagents in Mineral Technology.
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Klimpel, R.R., and S. Isherwood. 1991. Some industrial implications of changing frother chemical
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Leja, J., and J.C. Nixon. 1957. Ethylene oxide and propylene oxide compounds as flotation reagents.
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Lu, S., R.J. Pugh, and E. Forssberg. 2005. Interfacial Separation Processes. Studies in Interfacial Science,
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Lucassen, J. 1992. Capillary forces between solid particles in fluid interfaces. Colloids Surf. 65(2–3):131.
Manev, E., and R.J. Pugh. 1992. Study of drainage and equilibrium thickness of single foam films
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Moxton, N.T., C.N. Bensley, R. Keast Jones, and S.K. Nicol. 1987. Insoluble oils in coal flotation: The
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Pugh, R.J. 1996. Foaming, foam films, antifoaming and defoaming. Adv. Colloid Interface Sci. 64:67–142.
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BLANK LEFT PAGE
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Surface Characterization and New
Tools for Research
R.St.C. Smart, W.M. Skinner, A.R. Gerson, J. Mielczarski, S. Chryssoulis, A.R. Pratt,
R. Lastra, G.A. Hope, X. Wang, K. Fa, and J.D. Miller
INTRODUCTION
In the selective separation of mineral phases by flotation, surface chemistry is the principal
determinant of the average contact angle for a specific mineral phase in a flotation pulp. The
average contact angle is, in turn, the principal determinant of the bubble–particle attachment efficiency (Ea) in the overall collection efficiency (Ec) from which the flotation rate
constant can be determined (Ralston 1994a). The recovery and selectivity in sulfide flotation is ultimately dependent on the relative rate constants of the different mineral phases.
The average contact angle is not only mineral-specific, based on a statistical average of the
mineral particles in that phase, but also the contact angle for each particle is an average of
hydrophobic and hydrophilic areas across the particle surface. Determination of this hydrophobic/hydrophilic balance by particle therefore requires selection of the particular mineral
phase. Obtaining this information is not necessarily a simple task in a flotation pulp containing many different mineral phases, different particle sizes of individual phases, adsorbed
and precipitated species (often colloidal), and oxidized products. The hydrophobic/hydrophilic balance by particle and its statistical average by mineral phase require identification of
the major species contributing to each category in surface layers (Smart, Jasieniak et al.
2003). In addition to adsorbed collector molecules and their oxidized products (e.g.,
dimers), hydrophobicity can be imparted to sulfide mineral surfaces by oxidation to produce polysulfide Sn2– species resulting from loss of metal ions (usually Fe2+) from surface
layers. In acid solution, hydrophobic elemental sulfur can also be formed and is usually
imaged in patches on the sulfide mineral surface (Smart, Amarantidis et al. 2003). Almost
all other species found on sulfide mineral surfaces, such as oxide/oxyhydroxide/hydroxides,
oxy-sulfur (e.g., sulfate), carbonate, hydrous silica, and fine gangue particles, are essentially
hydrophilic but may be in the form of localized particles, colloids, and precipitates or continuous, reacted, or precipitated surface layers (Smart, Amarantidis et al. 2003).
The action of collector molecules in inducing hydrophobicity can be assisted by activating species such as copper and lead ions that complex the collector on the surface. Previous
research has shown that this activation can be inadvertently produced by dissolution and
transfer via solution of these ions to the surfaces of mineral phases not intended to float (e.g.,
Smart 1991; Lascelles and Finch 2002; Finkelstein 1997). Other complex mechanisms can
affect both hydrophilic and hydrophobic contributions. They include, for instance
a. The extent of liberation of individual mineral phases by grinding or, conversely, the
extent of remaining composite particles
b. Chemical alteration of the surface layers of sulfide minerals induced by oxidation
reactions in the pulp solution
283
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FLOTATION FUNDAMENTALS
c. The presence of a wide range of particle sizes in the ground sulfide ores
d. Galvanic interactions between different sulfide minerals to produce different reaction products on the mineral surfaces
e. Chemical interaction between particles in the form of aggregates and flocs
f. The presence of colloidal precipitates arising from dissolution of the minerals, particularly sulfides, and grinding media
g. The mechanism of adsorption of reagents on specific surface sites
h. Competitive adsorption between oxidation products, conditioning reagents, and
collector reagents
The primary purpose of surface characterization is, therefore, to identify these mechanisms with secondary application to the understanding and control of these mechanisms in
plant practice.
This chapter summarizes some of the advances in surface analytical techniques and
knowledge gained from surface characterization applied to flotation systems in the last two
decades. There have been major additions to both the armory of instruments and the methodologies available for these studies. The selection of examples from both techniques and
methods has been made as objectively as possible but necessarily has resulted in many excellent reports of research and plant surveys (or diagnoses) that have been missed or inadequately described. An attempt was not made to describe the theoretical or practical basis for
each technique but, rather, the type of new information that it can provide. In compensation, reference is made to technique descriptions, reviews, and more complete reports in the
references list. Some of the techniques surveyed in this chapter are now used in ore and
plant surveys to provide part of the metallurgical information on which plant control is
based but, more often, to diagnose reasons for losses in recovery or selectivity. Other techniques have been applied to basic studies of processes and mechanisms controlling flotation.
Some case studies, combining information from different techniques, have been included to
illustrate the now-established place of surface analysis and characterization in flotation
research and practice.
X - R AY P H O T O E L E C T R O N A N D A U G E R E L E C T R O N
SPECTROSCOPIES
X-ray photoelectron and Auger electron spectroscopy techniques (Briggs and Seah 1992;
O’Connor, Sexton, and Smart 2003), both analyzing the first 2–5 nm of the surface layers
used in static (spot) and imaging modes, have produced primary information on processes b,
c, e, f, and g previously described.
Sulfide Oxidation: Polysulfides
In sulfide flotation, the mechanisms of surface oxidation and the consequent physical and
chemical forms of oxidation products on the surface, which are derived from studies using
these surface analytical techniques, can be summarized as
• Metal-deficient (sulfur-rich) surfaces, polysulfides, and elemental sulfur
• Oxidized fine particles attached to larger sulfide particle surfaces
• Colloidal precipitates of metal hydroxide particles and flocs
• Continuous oxidized surface layers (e.g., oxide/hydroxide) of varying depth
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285
Adsorbed sulfate and carbonate species
Nonuniform spatial distribution with different oxidation rates (e.g., isolated, patchwise oxidation sites; face specificity)
Scanning Auger microscopy (SAM) can provide both electron (secondary and backscattered) microscopic imaging for topography and phase identification together with surface elemental mapping, line scans, spot (submicron) analysis, and depth profiles. This type
of information is illustrated in Figure 1 where a 50 wt % pyrite/50 wt % chalcopyrite mineral mixture was ground and conditioned in pH 9 nitrogen purged solution for >14 days. A
very wide range of particle sizes is evident. Oxidized fine particles 0.1–10 μm are imaged, as
well as flocs comprising clumps of loose aggregates with dimensions 1–3 μm, consisting of
smaller spheroidal particles each with approximate diameters 0.1–0.5 μm. A similar range of
particles is usually seen in samples taken from operating flotation plants. The flocs are not
necessarily observed on similar surfaces that have been laboratory-conditioned for shorter
periods (i.e., <1 hr) but have been seen in plant samples following fine grinding in which
accelerated dissolution of pyrrhotite and iron grinding media has occurred. They have been
precipitated from saturated ferric hydroxide solution as colloidal Fe(OH)3 particles.
Mechanical agitation or ultrasonication/decantation easily removes these flocs and some
oxidized fine particles from the sulfide surfaces, showing that they are only weakly bonded
to the oxidized sulfide minerals. The depth profiles for the oxygen signal in Figure 1c illustrate the variation in oxidized surface layers on different positions on the pyrite (left-hand
side, upper particle) and chalcopyrite (right-hand side, underlying particle) from 5 to 80 nm.
The >300-nm signal is from a ferric hydroxide floc. All of these surface species, which are
typically in islands or reacted patches of the particle surface, contribute to hydrophobicity.
•
•
1
MS + xH 2 O + --xO 2 ↔ M 1 – x S + xM ( OH ) 2
2
(EQ 1)
It is now well established that iron-containing sulfide minerals (e.g., pyrite, pyrrhotite,
chalcopyrite, pentlandite, arsenopyrite) essentially follow a reaction mechanism similar to
that in Equation 1, in that iron hydroxide products and an underlying metal-deficient or sulfur-rich sulfide surface are formed. The seminal work of Buckley, Woods, and their colleagues, using a combination of X-ray photoelectron spectroscopy (XPS) and
electrochemical techniques (reviewed in Smart, Amarantidis et al. 2003), has clearly demonstrated this mechanism in single mineral studies. In their work, oxidation of abraded pyrite
surfaces exposed to air for a few minutes produced a high binding energy (BE) doublet component of the S 2p spectrum in addition to ferric oxide/hydroxide reaction products (Buckley and Woods 1987). The sulfur product was attributed to an iron-deficient Fe1–xS2 surface
layer with the later proposition that polysulfide-like species Sn2– are formed. Specifically,
Mycroft et al. (1990) have correlated XPS and Raman spectra of electrochemically-oxidized
pyrite surfaces with polysulfide model compounds but only at Eh > 600 mV, pH 5. Recently,
monolayer-sensitive synchrotron radiation XPS (SRXPS) and time-of-flight secondary ion
mass spectrometry (TOF-SIMS) (discussed in the following sections) have been used to verify the polysulfide formation in surface oxidation under plant conditions.
The importance of the Sn2– n>2 polysulfides is that they contribute to hydrophobicity
independently of collector addition. Collectorless flotation of pyrite in alkaline solution,
correlated to electrochemical oxidation, can be explained by the production of a hydrophobic sulfur-rich surface together with hydrophilic iron hydroxide species (Ahlberg, Forssberg,
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FLOTATION FUNDAMENTALS
A
B
C
Source: Adapted from Smart, Amarantidis et al. 2003.
FIGURE 1 SAM images from pyrite and chalcopyrite particles at successively higher
magnifications (i.e., white bar = (a) 100 μm, (b) 10 μm, (c) 5 μm). Oxygen depth profiles for
Auger analysis gave approximately 5, 30, >300 (pyrite particle), 80, and 10 nm (chalcopyrite
particle), respectively, for each of the five points in (c).
and Wang 1990). After grinding, the surface becomes substantially covered by the hydrophilic species, and no significant flotation is observed without addition of collector. Collectorless flotation can, however, be easily obtained after complexing the iron with
ethylenediamine tetraacetic acid (EDTA) in solution, indicating that the underlying hydrophobic sulfur-rich layer is responsible for pyrite flotation under these conditions. Elemental
sulfur was not evident at pyrite surfaces exposed to air or neutral to alkaline solutions (Buckley and Woods 1987; Buckley, Hamilton, and Woods 1985). Thin layers of elemental sulfur
were, however, observed on pyrite surfaces exposed to aerated, dilute sodium sulfide solutions (Buckley and Woods 1987; McCarron, Walter, and Buckley 1990). The collectorless
flotation of chalcopyrite after air exposure or solution oxidation has been directly correlated
with the surface composition determined by XPS (Zachwieja et al. 1989). Removal of iron
hydroxide species during conditioning in alkaline solution to leave the hydrophobic sulfurrich sulfide surface showed strong flotation. Conversely, oxidized chalcopyrite surfaces
reduced in situ became copper deficient and were unfloatable.
A specific illustration of the XPS observation of polysulfide formation is found in
galena oxidation in pH 9 solution from fresh fracture to 30-minutes aeration (Figure 2)
(Smart et al. 2000). The growth of the high BE components of the sulfur S 2p spectra, due
to Sn2– species (B n=2; C n>3), is correlated with increasing contact angle and flotation
recovery (Prestidge and Ralston 1995). The evidence for assignment of the high BE components of S 2p XPS spectra to metal-deficient, polysulfide defect sites and elemental sulfur
has been reviewed (Smart, Skinner, and Gerson 1999).
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B. 10 Minutes Oxidation in pH 9 Solution
A 160.8 eV 100%
A
170
168
166
164
162
160
158
Intensity, arbitrary units
Intensity, arbitrary units
A. As Fractured
287
A 160.8 eV 77%
B 162.0 eV 12%
C 163.1 eV 11%
A
C
170
168
Binding Energy, eV
166
164
B
162
160
158
Binding Energy, eV
Intensity, arbitrary units
C. 30 Minutes Oxidation in pH 9 Solution
A 160.8 eV 69%
B 162.0 eV 16%
C 163.1 eV 15%
A
C
170
168
166
164
B
162
160
158
Binding Energy, eV
Source: Adapted from Smart et al. 2000.
FIGURE 2
XPS S 2p spectra from galena
Collector Actions
XPS and SAM have also done much to increase understanding of the actions of collector
molecules in flotation, which are considerably more complex than the earliest simplistic
model of adsorption of the head group and dangling hydrophobic tail. Their actions in several different modes have been studied using XPS and SAM in recent research (Smart, Amarantidis et al. 2003), namely
• Adsorption to specific surface sites
• Colloidal precipitation of metal–collector species from solution
• Detachment of oxidized fine sulfide particles from larger particle surfaces
• Detachment of colloidal metal oxide/hydroxide particles and flocs
• Removal of adsorbed, oxidized surface layers
• Inhibition of oxidation
• Aggregation and disaggregation of particles
• Patchwise or face-specific coverage
The presence of adsorbed xanthate on freshly fractured galena surfaces has been confirmed from both S 2p spectra and the more surface-sensitive X-ray induced Auger spectra
(i.e., S LMM and Pb NOO) signals. The work of Buckley and Woods (1991) has correlated
xanthate coverage (using voltammetry) with XPS spectra and flotation recovery showing
that only a fraction of the monolayer is adsorbed at maximum recovery. Sub-monolayer, perpendicularly-oriented, adsorbed lead ethyl xanthate was confirmed in combined XPS, Fourier transform infrared (FTIR) spectroscopy, and controlled potential studies (Suoninen
and Laajalehto 1993). There are now many examples of studies in the literature in which
uptake of the collector molecules on the sulfide mineral surface occurs through the formation of colloidal (precipitated) metal-xanthate or metal-hydroxyxanthate species in solution
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TABLE 1 Survey of the atomic concentrations (%) of elements for undeslimed pyrite particles
conditioned at pH 4, N2 purged
Initial
Element
Carbon (C)
Oxygen (O)
Iron (Fe)
Sulfur (S)
Calcium (Ca)
Phosphorus (P)
Etch
[DCDTP] = 0 M
28
7.8
28
24
10
27
27
33
3.6
4.6
0
0
Initial
Etch
[DCDTP] = 7.5 × 10–3 M
26
8.3
11
5.2
21
39
40
47
0.45
<0.05
0.28
0
or close to the surface. XPS and FTIR studies have shown that both copper(I) ethyl xanthate (XPS) and diethyl dixanthogen (FTIR) can be detected on the sphalerite surface as a
result of reaction (Prestidge et al. 1994).
It is generally agreed that if sufficient hydrophilic oxidized material is present on the
mineral surfaces, this will overcome any natural or self-induced floatability as well as modify
the collector-induced floatability of the sulfide particles (e.g., Shannon and Trahar 1986). It
has been suggested that collector molecules have the dual action of removing oxidized products from surfaces and providing a hydrophobic surface for bubble attachment and flotation. The surface cleaning action of xanthate has been reviewed by Senior and Trahar (1991)
in studies of the flotation of chalcopyrite in the presence of metal hydroxides. The “cleaning” mechanism does not appear to be simple dissolution of the oxidized metals as can be
achieved with other complexing agents (e.g., EDTA), because the amount of xanthate necessary to restore floatability is stoichiometrically orders of magnitude below that required for
the complete conversion of the metal hydroxides to dissolved metal xanthate species. This
fact is also well established in plant practice where EDTA additions are effective in giving
increased recovery but are prohibitively expensive.
XPS indications of the cleaning action in removing oxidized surface layers were previously
noted (Smart 1991), in which the initial high BE shoulder on the Cu 2p peaks, attributed to
charged hydroxide species, had been removed by the action of the xanthate from a ground
chalcopyrite sample. A similar action is found with the collector dicresyldithiophosphate
(DCDTP) (van der Steldt, Skinner, and Grano 1993). Table 1 shows that the surface oxygen percentage is dramatically reduced (~40%) by addition of 7.5 × 10–5 M at pH 4 (maximum adsorption) to undeslimed pyrite particles. Removing the fine particles from the
ground pyrite sample, by three successive ultrasonication/decantation steps with solution
replacement, also showed that the majority of the material removed by the collector addition is in the form of fine particles rather than oxidized surface layers, as the level of surface
oxygen concentration after fines removal (i.e., 13%) is similar to that after collector addition. After collector addition, both sulfate and ferric hydroxide signals are reduced in the
XPS spectra and the contamination by calcium, present on the surface before collector addition, is also removed. There has clearly been a process in which oxidized fine particles are
detached from the pyrite surface by the collector addition similar to that in mechanical
removal of these fine particles.
SAM results (Smart, Amarantidis et al. 2003) have demonstrated that at high xanthate
concentrations (i.e., 10–3 M), the formation of a surface layer, likely to consist of the oily,
hydrophobic dixanthogen species, can inhibit oxidation of pyrite surfaces that would otherwise produce iron hydroxide oxidation products on the surface and in solution at lower
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289
xanthate concentrations. These results bring in to question the interpretation and relevance
of mechanisms derived from studies at higher collector concentrations where the collector
has been added in the grinding stage or early in the conditioning stage.
Fine/coarse particle interactions and aggregation have been studied for the copper activation of sphalerite and the effect of xanthate addition using correlated microflotation and
on-line particle size distribution analysis coupled with surface analysis. Significant aggregation of <20-μm fine particles to 20–30-μm aggregates is already observed during initial conditioning at pH 5.5 and 8.5 during the first 20 minutes before any reagent addition.
Addition of Cu(II) at pH 5.5 produces some development of larger aggregates, 100–120 μm,
but flotation response of the fine particles remains poor. At pH 8.5, there is no evidence of
the larger aggregates on copper addition and flotation recovery of the fine particles is similarly poor. Addition of xanthate at pH 5.5 dramatically increases the extent of aggregation
>100 μm and flotation recovery (>95%), whereas at pH 8.5, only weak, larger aggregates
appear to form, and the recovery does not increase to the same extent (i.e., 60%–80%). In
the presence of coarser particles (38–75 μm) at pH 5.5 with Cu(II) addition alone, a very
high percentage of the fine particles were recovered and a fine/coarse particle aggregation
(i.e., “piggybacking”) mechanism was confirmed by on-line particle size analysis and optical
microscopy. At the higher pH 8.5, the interaction between fine and coarse sphalerite is
slower and much less complete with correspondingly poorer flotation of both the fine and
coarse fractions. XPS surface analysis confirms that this is due to partial coverage of the surface by colloidal hydroxides and overall hydrophilicity inhibiting strong hydrophobic interactions between the particles (Lange, Skinner, and Smart 1997).
There are now many case studies in which XPS surface analytical studies, combined
with flotation metallurgy and solution chemistry, have directly contributed to improvements in flotation recovery and grade. For example, naturally floatable iron sulfides with
graphitic surface layers have been identified and separately removed in copper flotation at
Mount Isa, Australia (Grano, Ralston, and Smart 1990). This study also identified the selective removal of ferric hydroxides and carbonates by collector addition. The effects of fine
grinding on flotation performance have been surveyed in a correlated XPS-flotation study
(Frew et al. 1994) and specific application to the zinc regrind at Cominco Alaska’s Red Dog
mine has been reported (Frew, Smart, and Manlapig 1994). The action of an extended
period of aerated conditioning before copper activation and collector conditioning in
increasing sphalerite flotation at the Murchison Zinc (Australia) concentrator (Kristall et al.
1994) was explained, using XPS, by the removal of zinc hydroxides from the sphalerite surface and the concomitant appearance of a metal-deficient sphalerite surface. XPS also demonstrated an increase in the oxidation state of pyrite after aerated preconditioning. The
presence of excessive surface oxidation in copper reflotation at Western Mining Corporation’s Olympic Dam operation (Australia) (Smart and Judd 1994), identified by XPS analysis, led to improved operation of Lasta filters. A low flotation rate of galena in lead roughers
at the Hilton Concentrator of Mount Isa Mines was analyzed (Grano et al. 1993, 1996)
using XPS. The presence of precipitated species and their removal by a change of conditioning reagents (i.e., lime to soda ash) and collector reagent (i.e., ethyl xanthate to DCDTP, a
collector that is stable in the presence of sulfite species over a wide pH range) has been used
to address this problem.
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S Y N C H R O T R O N R A D I AT I O N X P S
Spectroscopic Advantages
Conventionally, obtaining a profile of composition as a function of depth requires either
sequential ion beam sputtering and analysis, or the use of angle-resolved photoemission.
Sputter profiling is limited because the disruption of chemical bonds beyond the top few
monolayers in the surface detracts from any chemical environment information. Differential
sputtering can also occur in compounds, resulting in a progressive enrichment of one constituent over another. Angle-resolved XPS can yield nondestructive, surface-sensitive information but requires flat surfaces and depth accuracy requires knowledge of photoelectron
inelastic mean free paths in overlying reaction products. The majority of exposed sulfide
mineral surfaces are not flat. Indeed, only a few sulfide minerals exhibit good cleavage planes
(e.g., galena, sphalerite). It is also rare for the thickness distribution of reaction products to
be uniform.
The XPS analysis depth can be modulated by controlling the kinetic energy and, hence,
the escape depth of the emitted photoelectron of interest. This capability is afforded
through the use synchrotron radiation. Current soft X-ray synchrotron beamlines offer high
resolution monochromation in the 10–2,000 eV energy range with resolving powers in
excess of 104 (E/ΔE), and photon flux that is several orders of magnitude higher than conventional sources. Energy resolution and photon flux allow for full capability of the electron
analyzer to be realized, giving superior photoelectron spectral line widths and excellent
signal-to-noise ratio in short analysis times.
The universal curve of photoelectron escape depth as a function of kinetic energy is
reproduced in Figure 3. The maximum surface sensitivity is attained when the kinetic
hν = BE + KE + Φ
1,000
S 2p BE ~ 160 eV
λ, monolayers
100
hν = 1,487 eV
10
1
hν = 210 eV
1
10
100
1,000
Electron Energy, eV
Source: Adapted from Carlson 1975.
FIGURE 3 Universal curve of photoelectron escape depth as a function of kinetic energy.
Photoemission energy conservation equation shows the relationship between incident photon
ν), photoelectron BE, kinetic energy (KE), and instrumental work function (φ
φ).
energy (hν
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SURFACE CHARACTERIZATION AND NEW TOOLS FOR RESEARCH
291
energy of the emitted photoelectron is in the range of 40–50 eV. To achieve this for the S 2p
core-line (~160 eV BE), for example, an X-ray energy of approximately hν = 210 eV should
be used. These conditions yield an almost fivefold enhancement in surface sensitivity over a
conventional Al Kα source of hν = 1,487 eV for the S 2p signal.
Surface Sensitivity and Fracture Surfaces
The enhanced surface sensitivity and spectral resolution afforded by SRXPS is exemplified
by the analysis of the fracture surface of pyrite (FeS2). Figure 4a and 4b reproduce the S 2p
core-line measured using a conventional monochromated Al Kα source (1,487 eV) and a
synchrotron source (206 eV), respectively. The narrower line widths are immediately apparent in Figure 4(b). Two new S 2p doublet contributions (peaks a and b) appear to be located
at lower BE relative to the main peak. These two new peaks in the pyrite S 2p spectrum are
confirmed to derive from surface contributions as their intensities relative to the main peak
(bulk peak) decrease with increasing excitation energy (Nesbitt et al. 1998). The interpretation of these surface core-level shifts requires consideration of the molecular orbitals
involved in bonding, the mineral structure (e.g., metal and ligand coordination), and the
fracture mechanism. A detailed description is not possible within the constraints of this
review and the reader is directed to work by Nesbitt et al. (1998) for a comprehensive explanation. Put simply, during pyrite fracture, both Fe–S and S–S bonds may be ruptured. This
results in a reduction in the coordination of metal and ligand (in this case, sulfur) sites at the
surface and changes in electron density associated with these sites. Electrons involved in
bonding are usually retained at the ligand (sulfur) site. The rupture of Fe–S bonds leaves the
A
(i)
S 2p
B
S 2p
(iii)
hν = 206 eV
hν = 206 eV
2–
S2O3
hν = 1,487 eV
hν = 1487 eV
S2x(2x–n)–
SO32–
(ii)
C
(iv)
D
hν = 210 eV
c
hν = 206 eV
b
hν = 206 eV
III
II
hν = 210 eV
a
170
168
166
164
162
Binding Energy, eV
160
170
168
166
164
162
160
Binding Energy, eV
FIGURE 4 Sulfur 2p core-line spectra of fractured pyrite surfaces (a) using a conventional
α source, and (b) a 206-eV synchrotron source (Schaufuss et al. 1998).
monochromated Al Kα
Subsequently reacted pyrite surfaces are shown for (c) air oxidation for 14 hours (Schaufuss
et al. 1998) and (d) adsorption of mercaptobenzothiazole from solution (Szargan, Schaufuss,
and Rossbach 1999). The spectral contributions from bulk sulfur dimers (S22–) are indicated by
the solid circles ( ) above the primary S 2p3/2 component.
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FLOTATION FUNDAMENTALS
surface sulfur dimer site with increased electron density and, hence, will emit photoelectrons of lower BE than the fully coordinated bulk sulfur dimer (peak b in Figure 4b). The
formal oxidation state of the monomeric S1– produced by S–S bond rupture reduces to S2–
through the associated oxidation of Fe2+ to Fe3+, giving rise to even greater electron density
associated with this site and hence an even lower BE (peak a) (Nesbitt et al. 1998). This
level of detail is not possible using conventional, laboratory-based XPS instrumentation and
the level of surface sensitivity afforded by SRXPS approaches that of static secondary ion
mass spectrometry (static SIMS), as well as atomic force and scanning tunneling
microscopies (AFM and STM) and low-energy electron diffraction. Moreover, nondestructive and quantitative surface- and bulk-sensitive measurements may be made in the same
experiment, simply by varying the incident photon energy.
A range of important metal sulfide minerals has been examined by SRXPS over the last
decade, looking at surface states and initial reactions. These include, for example, galena,
PbS (Paulucci and Prince 1990; Leiro et al. 1998; Kartio et al. 1998); pyrite, FeS2 (Bronold,
Tomm, and Jaegermann 1994; Nesbitt et al. 1998; Schaufuss et al. 1998); marcasite, FeS2
(Uhlig et al. 2001); pyrrhotite, Fe1–xS; millerite, NiS (Nesbitt et al. 2001); arsenopyrite,
FeAsS (Schaufuss et al. 2000); loellingite, FeAs2 (Nesbitt, Uhlig, and Szargan 2002); gersdorfite,
NiAsS (Nesbitt, Schaufuss et al. 2003); covellite, CuS; chalcocite, Cu2S (Laajalehto et al.
1996); and chalcopyrite, CuFeS2 (Harmer et al. 2004). This is by no means an exhaustive
list; however, Harmer and Nesbitt (2004) have published a treatment of SRXPS interpretation showing how the sulfide mineral structure and composition determine whether surface
reconstruction occurs as a result of fracture and the type of species exposed at the surface
(e.g., metal oxidation, ligand polymerization). Recent work, combining ab initio calculations with spectroscopic analysis (von Oertzen, Skinner, and Nesbitt 2005), has confirmed
this interpretation for pyrite, chalcopyrite, and molybdenite, MoS2 (von Oertzen, Harmer,
and Skinner, in press). This provides a high level of support for the synchrotron XPS interpretation of other important sulfide mineral fracture surfaces.
Surface Reaction
The implications of using these kinds of measurements for minerals processing are manifold
as they enable the first surface exposed to solution to be probed (i.e., the fracture surface
immediately exposed during grinding). The subsequent initial reactions at this surface may
also be studied in depth using SRXPS in similar detail. Figures 4c and 4d illustrate further
SRXPS examples of oxidation and collector adsorption at pyrite fracture surfaces. From
these types of studies, it is possible to follow the relative reactivity of the various surface
states exposed on fracture and, in turn, relate this reactivity back to the structure and bonding within the mineral.
Other Aspects of SRXPS
In the soft X-ray region, photoionization cross-sections can vary strongly with photon
energy. This can be used to great advantage in SRXPS, particularly in valence band studies.
By collecting the valence band spectrum of a mineral at several photon excitation energies, it
is possible to enhance or diminish the spectral intensity contributions from metal and
ligand bonding and nonbonding orbital, thereby identifying them and monitoring their
involvement in surface reaction (Nesbitt et al. 2002; Nesbitt, Uhlig et al. 2003). Where possible within the constraints of surface roughness, angle-resolved measurements may also be
performed to further enhance surface sensitivity, particularly for adsorbate studies.
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SRXPS in conjunction with near-edge X-ray absorption fine structure (NEXAFS)
spectroscopy provides an extremely powerful combination for the study of sulfide minerals
(Goh, Buckley, Lamb, Skinner et al. 2006; Goh, Buckley, Lamb, Roseberg et al. 2006).
NEXAFS yields information on metal and ligand coordination, oxidation state and relative
location. It is particularly useful in monitoring organic reagent bonding mechanisms and
molecular orientation at the mineral surface.
The Future
Much of the methodology for the study of fracture surfaces has now been established,
despite the past scarcity of appropriate beamline/end-station combinations for these studies. Recent developments such as the soft X-ray facilities at the Canadian Light Source and
the new Australian XPS/NEXAFS end station (initially located at the National Synchrotron Radiation Research Center in Taiwan) have expanded the available capabilities for such
investigations. The investigation of the mechanisms of subsequent surface reaction will necessarily provide impetus into the future, together with further instrumental developments
(e.g., imaging photoemission, effective charge neutralization for insulating minerals).
T I M E O F F L I G H T S E C O N DA RY I O N M A S S S P E C T RO M E T RY
Diagnosis of the surface chemical factors playing a part in flotation separation of a value
mineral phase requires measurement of the species that are statistically different between the
concentrate and tail streams, together with an estimate (if possible) of the magnitude of the
differences. The recently developed statistical methods, based on TOF-SIMS, have moved
toward this ultimate aim. This technique used in static mode involves a very low flux of
heavy ions impacting surface layers with mass spectrometric analysis of the secondary ions
emitted from the surface. In the time of routine measurement, only 1–2 surface atoms in
1,000 are impacted. The secondary elemental and molecular fragment ions come from the
first two molecular layers of the surface and provide a very detailed set of positive and negative mass fragments from simple ions (e.g., Na+, OH–) through to molecular ions of specific
reagents (e.g., isobutyl xanthate (CH3)2CHOCS2–). Identification of molecular mass peaks
for collectors, activators, depressants, precipitates, and adsorbed species is possible with
comparative surface concentrations by particle and by phase between feed, concentrate, and
tail streams.
The pioneering work of Nagaraj and Brinen (Brinen et al. 1993) with TOF-SIMS and
the initial statistical analysis of air-dried particles using laser ionization mass spectrometry
by Chryssoulis and colleagues (Chryssoulis, Reich, and Stowe 1992) have greatly contributed to the approaches described here. The improvements in the methodology include
introduction of the mineral particles without exposure to air (Smart 1991), analysis of surface monolayer (or two), and full statistical analysis of all surface species. These analyses provide a statistical basis for assessment of surface chemical factors that have differentiated
particles of a particular mineral phase that have reported to a concentrate from those that
have reported to the tail.
Validation: Statistical Analysis
The use of TOF-SIMS to quantify changes in surface chemistry has been extensively validated in several ways. The amount of collector (e.g., xanthate, dithiophosphate) adsorbed
from solution and monitored by UV adsorption was calibrated against the normalized
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TOF-SIMS intensity for the molecular parent ion. A linear relationship up to monolayer
coverage was found with a transition to a plateau for multilayer coverage due to the extreme
surface sensitivity of the TOF-SIMS analysis (Piantadosi 2001). This is not a serious limitation because most minerals processing plant dosages are sub-monolayer. A second validation
of the TOF-SIMS representation of the surface chemistry was conducted in comparison of
spectra from polished surfaces of stoichiometric troilite, FeS; iron-deficient pyrrhotite, Fe1–xS;
and stoichiometric pyrite, FeS2. The products of air oxidation of troilite and pyrrhotite have
previously been shown to include disulfides and polysulfides (Smart, Amarantidis et al.
2003). The experiment was designed to test the representation of this surface chemistry in
TOF-SIMS spectra. For the iron-sulfur fragments, the FeS2/FeS ratios for troilite, pyrrhotite and pyrite were 0.59, 1.2, and 32, respectively. A similar sequence of Sn/S atomic fragment distributions confirmed the presence of polysulfides in these slightly reacted surface
layers (Smart et al. 2000). Correlation of TOF-SIMS with XPS spectra (Figure 2) for
freshly-cleaved galena (PbS) surfaces reacted in pH 9 solution for increasing periods of time
has also shown a systematic increase in Sn/S ratios with increasing components of S 2p XPS
spectra corresponding to polysulfide formation (Smart et al. 2000).
The basis for the methodology of sample preparation, mineral phase recognition,
TOF-SIMS analysis, and statistical evaluation has been described in the paper by Piantadosi
and Smart (2002). TOF-SIMS images of the particles in total ion yield mode are similar to
those shown in Figure 5. Scanning for specific signals (e.g., Pb, Zn, Cu, Fe) can then be used
to identify the particles of a specific mineral phase (e.g., galena, sphalerite, covellite, pyrite,
chalcopyrite) for specific analysis. The region-of-interest (ROI) facility in the software
allows definition of selected particles, as in Figure 5, corresponding to a specific mineral
phase with the boundary for analysis set at a fixed position inside the contrast edge. When
sufficient particles of the mineral phase have been identified for reliable statistics, a mass
spectrum from each particle is recorded and stored. The statistical analysis (Piantadosi et al.
Pulses/Pixel: 1
10 μm
In
G
In
Sp
Sp
In
In
In
G
Sp
Py
Sp
Sp
Ch
Py
Ch
Py
Py
Sp
Ch
Ch
Py
Py
Sp
Ch
Ch
Ch
Ch
Sp
G
Sp
Sp
Ch
Py
Py
Sp
Ch
Sp
Sp
Ch
Sp
Sp
Sp
sum of rest 2589522 285
Source: Hart et al. 2004.
FIGURE 5 Principal component analysis identification of mineral phases: pyrite (Py),
sphalerite (Sp), chalcopyrite (Ch), gangue materials (G), indium mounting material (In)
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SURFACE CHARACTERIZATION AND NEW TOOLS FOR RESEARCH
295
2000) then determines a mean value for each atomic and molecular species with 95% confidence intervals for each signal.
This analysis was first applied to the effects of calcium ion depression on galena flotation (Piantadosi et al. 2000). Correlation for 26 particle sets of Ca/Pb ratios for concentrate
and tail streams gave 0.022 and 0.048, respectively, indicating that there is statistically more
calcium (~2x) on galena particles surfaces in the tails compared with the concentrate. A second study (Piantadosi and Smart 2002) of the effect of iron hydroxides and collector, isobutyl xanthate (IBX), on galena flotation compared normalized intensities of IBX in feed,
concentrate, and tail streams. There is a clear separation between particles of galena reporting to concentrate (0.005) and tail (0.001). The IBX concentrations on particles in the tail
are not statistically different from those in the feed, due to increased hydrophilicity of
galena particles in the tail rather than to reduced hydrophobicity. An early attempt to derive
a hydrophobicity/hydrophilicity index, based on a ratio of signals from the (hydrophobic)
collector (IBX) to (hydrophilic) oxy-sulfur products (SO3) and iron hydroxide (FeOH),
gave a value for the concentrate of 44.7 ± 13.7 compared with 7.1 ± 2.4 for the tail, but it is
recognized that the index does not include all hydrophobic or all hydrophilic species contributing to the separation. Further comparisons for laboratory separation of galena and
pyrite using di-isobutyl dithiophosphate (DIBDTP) collector have also been reported
showing statistically ~12 times more collector on galena compared to pyrite. Galena particles
reporting to the concentrate show statistically less calcium, lead hydroxide, and oxy-sulfur species
on their surfaces compared to tail particles. The early flotation of galena was also considerably assisted by the presence of colloidal as well as adsorbed Pb DIBDTP.
Plant Diagnosis
This methodology has now been applied to full ore samples from plant operations including, as examples, Mount Isa Mines (MIM), Ok Tedi Mining Ltd. (OTML, Papua New
Guinea), Falconbridge (Strathcona, Canada), Anglo Platinum (South Africa), Mineracao
Caraiba (Brazil), and Inco Matte Concentrator (Sudbury, Ontario, Canada)—a total of 18
full statistical analyses to date (Smart, Jasieniak et al. 2003). Examples of results from samples
supplied by the client from rougher and rougher scavenger flotation are shown in Figure 6.
Poor flotation kinetics were exhibited by fine sphalerite (–10 μm) copper-activated down
the rougher-scavenger banks. The study was designed to determine whether the poor flotation response to fine sphalerite was due to differences in mineral surface chemistry rather
than hydrodynamic collision frequency factors alone. The process characteristics included
pH 10.5 adjusted with lime, collector addition of IBX, and copper sulfate activation.
Sphalerite particles in the –10-μm size range were selected using the ROI methodology so
that the surface chemistry of this mineral phase was examined selectively. The bars in Figure
6 show the median value of each positive and negative ion signal with the 95% confidence
intervals indicated by the smaller intervals at the top of each bar. Comparison between the
rougher concentrate, scavenger concentrate, and scavenger tail eliminates all species for
which confidence intervals overlap as not statistically significantly different (at least to the
first level of statistical analysis). Other signals are clearly statistically different with dissimilar
magnitudes in this comparison. In the selected positive ion species, discrimination into the
concentrate streams is indicated for Zn, Cu, and Na with discrimination into the tail for
increasing Fe, K, Si, Al, and particularly Mg. Low surface concentrations of Ca appear to
favor the rougher concentrate but are apparently depressant into the scavenger concentrate
and tail. In negative ion SIMS, the concentrates are statistically favored by high exposure of
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FLOTATION FUNDAMENTALS
(+) SIMS, Sphalerite Particles (Slimes)
0.40
Scavenger Tail
Scavenger Concentrate
Rougher Concentrate
0.35
Normalized Intensity
0.30
0.25
0.20
0.15
0.10
0.05
0.00
Na
Mg
Al
Si
K
Ca
Fe
Cu
Zn
FeOH
(–) SIMS, Sphalerite Particles (Slimes)
0.50
Scavenger Tail
Scavenger Concentrate
Rougher Concentrate
0.45
Normalized Intensity
0.40
0.35
0.30
0.25
0.20
0.15
0.10
0.05
0.00
C
CH
O
OH
F
S
SO3
IBX
Source: Smart, Jasieniak et al. 2003.
FIGURE 6 Statistical TOF-SIMS spectra for sphalerite particles in the scavenger tail,
scavenger concentrate, and rougher concentrate (bars = 95% confidence intervals)
S, CH, and by low surface exposure of F, OH, and O. More detailed examination of the collector IBX signal shows close correlation with the Cu signals. But the oxy-sulfur SO3 signals
are not statistically different between the three streams.
Comparisons show that the surface chemistry of the fine sphalerite is significantly different between concentrates and tails and even, for some species, between the rougher and
scavenger concentrates. The most important difference is the absence of copper exposure
and associated IBX on fine sphalerite in the tail stream, indicating low hydrophobicity of
these particles. This difference is exaggerated by the presence of high concentrations of Mg,
Ca, Al, OH, and F ions, apparently in the form of hydroxides and (alumino)silicates obscuring copper activation. Oxidation to oxy-sulfur species is not a major factor in the depression
of fine sphalerite. Calcium ions, in particular, appear to have a depressant role between the
rougher and scavenger flotation stages.
A second example of this statistical analysis for chalcopyrite in the OTML system has
recently been published (Piantadosi, Pyke, and Smart 2001). It has been possible to estimate
an average contact angle for this mineral phase by comparison with single mineral studies
using the same collector.
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297
Scores on PC# 2
250
0.4
0.3
0.2
0.1
0.0
–0.1
–0.2
–0.3
Scaled 95% Limits are 149.93 and 70.4232
Scores on PC# 2
Zn
PC2
Cu
0
20
40
60
80
Fe
100 120 140 160 180 200 220 240
m/z
2-PC2
2
1
0
1
5
Scaled 95% Limits are 162.531 and 86.1104
Fe
–1
0
20
40
60
80
100 120 140 160 180 200 220 240
m/z
Source: Hart, Biesinger, and Smart 2006.
FIGURE 7 Second PC image scores and factor loadings for positive ion TOF-SIMS image data
from the chalcopyrite/pyrite/sphalerite mixture. Top: autoscaled data; bottom: mean centered
data. Images are 100 × 100 microns. Positive factor loadings appear bright in image; negative
loadings appear dark.
Principal Component Analysis: Phase Recognition and Statistics
A recent improvement in the statistical analysis has been the introduction of principal component analysis (PCA) applied to the mass spectra. Reliable identification of specific mineral particles is central to this statistical analysis. A chalcopyrite/pyrite/sphalerite mineral
mixture conditioned at pH 9 for 20 minutes to study transfer of Cu from chalcopyrite via
solution to the other two mineral surfaces—because this mechanism can be responsible for
their inadvertent flotation in copper recovery—showed no statistical difference in the copper intensities on pyrite and sphalerite (selected from Fe and Zn images) after this conditioning. PCA identifies combinations of factors strongly correlated (positively or
negatively) in images or spectra from sets of data. In images, PCA selects these correlations
from the mass spectra recorded at each of 256 × 256 pixels in a selected area of particles. In
the image mode, PCA has proved to be a much better method of selecting particles by mineral phase with clearer definition of particle boundaries because of multivariable recognition. The first principal component, with factor loadings that are positive in weighting for
all masses, is representative of the largest variance in the data set: topography and matrix
(ion yield intensity) fluctuations. The second and subsequent principal components (PCs)
will then have this variance removed and, as such, are topography- and matrix-corrected.
Figure 7 illustrates the selection of sphalerite, chalcopyrite, and pyrite phases in the second
PC from each of autoscaled and mean centered calculation modes. The transfer of copper
ions from chalcopyrite dissolution to both pyrite (Smart 1991; Hart et al. 2004; Hart,
Biesinger, and Smart 2006) and sphalerite surfaces (Finkelstein 1997) is confirmed by the
surface analysis but it has also clearly separated a statistical difference in copper intensities
between the sphalerite and pyrite phases in favor of sphalerite.
The PCA method has been applied to concentrate and tail samples collected from the
Inco Matte Concentrator demonstrating extensive CuOH and NiOH transfer between the
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FLOTATION FUNDAMENTALS
chalcocite (Cc) and heazlewoodite (Hz) minerals (Hart et al. 2004). The PCs gave excellent
recognition of the two mineral phases with reliable statistics on the regions selected. The
close correspondence in these correlations between the isotopes of Cu (63, 65) and Ni (58,
60) gave further confidence in the interpretation. Statistical differences in normalized intensities (values are given in parentheses) illustrate the important discriminating depressant
action of NiOH in flotation despite the activation of Hz by Cu transfer. The inadvertent
flotation of Hz in the concentrate appears to be a result of Cu activation (0.16). There is also
abundant Cu on Hz particles in the tails (0.08), but this is about half that in the concentrate.
The Cu distribution between Cc and Hz particles in both concentrate and tails is the same
within statistical 95% confidence intervals. The large statistical difference is in the Ni distribution where there is much more (~5×) hydrophilic Ni ions on Cc particles in the tail compared with the concentrate. Hence, Cc in tails appears to be the result of high depressant
hydrophilic loadings rather than absence of hydrophobic Cu-collector surface species. The
exposure of Cu on Cc particles in the tails as compared with concentrate is ~0.5, corresponding to an increase in Ni exposure of ~7.5. Both Cu activation of Hz and Ni depression
of Cc are clearly operating in this system. There is also considerably more collector (>4×) on
Cc particles in the concentrate than in the tails. In the tail samples, there is no statistical difference in intensity of the collector signals between Hz and Cc. The possible depressant
action of Ca ions is not found to be selective in this surface analysis. Ca is found on both Cc
and Hz surfaces in the concentrate and tails in statistically inseparable signals. The reduced
chalcocite hydrophobic/hydrophilic ratio is, therefore, related to the presence of Ni ions on
the surface, with a consequent reduction in bubble attachment efficiency. Hence, the statistical analysis can be used to confirm some mechanisms and deny others proposed to control
recovery and selectivity, giving more focus on the control factors.
T I M E O F F L I G H T L A S E R I O N I Z AT I O N M A S S S P E C T R O M E T R Y
Laser probe microanalysis combining laser excitation of small samples with a time-of-flight
mass spectrometer dates back to the mid-1960s (Ruckman 1986). The application of the
time-of-flight laser ionization mass spectrometry (TOF-LIMS) to analyze solid surfaces was
introduced in 1986 by Clarke, Ruckman, and Davey. In 1988, while analyzing pyrite from
the Brunswick Cu-Pb-Zn concentrator, it was accidentally discovered that free pyrite particles,
devoid of galena inclusions, had significant levels of lead on their surfaces. The analyses
showed that lead was confined to the surface and that it was more abundant on floated, as
compared to rejected, free pyrites. Thus, the potential of TOF-LIMS to analyze the surface
of mineral particles from plant samples in order to explain phenomena such as loss in selectivity
during differential flotation, concentrate dilution, and rejection of free valuable minerals,
was recognized. Comparative surface microanalysis by TOF-LIMS did become an integral
part of several plant surveys but in most cases, this type of work is used for troubleshooting
in (Cu)Pb-Zn, (Au)Cu, and platinum-group element (PGE) flotation plants (54 in total
during the last 15 years).
Sample Preparation
Sample preparation for surface microanalysis with the laser microprobe was kept as simple as
possible, on the premise that surface compositional differences and, in particular, their magnitude as opposed to absolute values, dictate the distinct response to flotation of otherwise
similar particles. The only requirement is that all samples are treated in exactly the same way.
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299
The sample preparation protocol requires cutting representative samples, filtering, followed
by a brief rinse with deionized water to displace mill water, and then air drying. Cold storage
in vials under a nitrogen atmosphere is encouraged for samples with readily oxidizable minerals (e.g., galena, pyrrhotite). Gravity separation by panning is used only when required to
preconcentrate free particles of rare minerals, thereby facilitating picking under the stereoscope (e.g., gold, PGE minerals). Typically, 20 to 30 particles of the mineral of interest are
picked from each sample or from two size fractions of the same sample when compositional
differences between coarse and fine particles are being investigated.
Analysis
The commercially available laser microprobe instruments, the LIMA-2A, and its successor
the PRISM, use Cassegrain optics, which permit specimen illumination and viewing, laser
irradiation, and ion extraction—all normal to the sample surface (Figure 8). This unique
feature is ideal for rough surfaces such as mineral particles because it minimizes topographic
effects and increases elemental sensitivity. The use of a two-laser system (Figure 9) for the
nonresonant multiphoton ionization (NR-MPI) of neutrals ablated by the laser microprobe
technique (Schueler, Odom, and Evans 1986) allows for the decoupling of the sampling
from the ionization step. This, in turn, enables an increase in surface sensitivity to the point
where only copper is detected from a copper-activated sphalerite particle (i.e., monolayer
detection), with a concurrent increase in elemental sensitivity (yielding minimum detection
limits in the 1–50-ppm range (Figure 10). The analytical spot size (1 to 30 μm) and surface
sensitivity (0.001 to 0.025) are inversely related to the degree of focusing and the power of
the ablator laser. Laser probe microanalysis is very fast, with more than 100 analyses performed and processed per hour. Thus, large data sets, typically 200–400 strong, are collected and form the basis of t-test comparative statistical analysis, which is used to identify
and rank activators and depressants (Bolin, Chryssoulis, and Martin 1997). Rotational factor analysis, a multivariate statistical analysis program, is used to validate t-test findings by
analyzing data groupings based on factor loadings.
The TOF-LIMS technique in the NR-MPI mode is ideally suited for elemental surface
microanalysis. In the single-laser negative ion mode, simple radicals (OH, CO3, SO3, SO4,
AsO4, FeOHx) associated with surface oxidation can be easily detected. Collector identification and loading measurements, although possible by TOF-LIMS (Chryssoulis et al. 1995),
is preferably done by the complementary TOF-SIMS and vacuum UV surface analysis by
laser ionization (VUV-SALI) techniques, as they are more sensitive for organic surface
microanalysis because of better preservation of molecular ions (Chryssoulis, Weisener, and
Dimov 1995).
The use of TOF-LIMS to quantify changes in surface composition has been extensively
validated using several elements (e.g., Cu, Pb, Au, Ag) and minerals (e.g., pyrite, sphalerite,
carbonaceous matter). Two examples are the linear relationship of TOF-LIMS data on surface Cu on sphalerite with the milligrams of Cu consumed, determined from solution assays
(Chryssoulis, Kim, and Stowe 1994); and the surface concentration values of copper (in
atomic %) measured by XPS (Stowe et al. 1993). Quantification of TOF-LIMS surface data is
possible using minerals loaded to different degrees with the element of interest or with the
help of relative sensitivity factors determined under standardized conditions (Dimov and
Chryssoulis 1997). Surface characterization by TOF-LIMS and, in fact, any other surface
microanalytical technique, is usually not a standalone investigation. In most cases, it is a followup investigation on a liberation study, which identified liberated minerals of readily
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Ionizing Laser
Ablator Laser
and Light Path
L1 L2 L3
Target
Ions to TOF
x
y
z
Cassegrain Optics
FIGURE 8
Ion Lens
Schematic of TOF-LIMS sample analysis region
Ablator Laser
Ablated Ions
and Neutrons
Particle
Ion
Lens
Ion Deflector
Ion Trajectory
Sample
Holder
Focusing
Optics
Ion Reflectron
Ion Detector
Ionizing Laser
FIGURE 9
TOF-LIMS principle of operation
floatable size classes to be a significant fraction of losses to tails or a concentrate grade loss or
misplacement to the wrong concentrate. Several case studies are discussed in the follow sections to illustrate this point.
Brunswick Cu-Pb-Zn Concentrator
In the Brunswick Cu-Pb-Zn concentrator (British Columbia, Canada), two long-standing
issues have been the premature flotation of sphalerite in the primary Cu-Pb circuit, culminating in the production of a bulk Zn-Pb concentrate (Figure 11) and the dilution of the
final Zn concentrate by pyrite. Several detailed liberation studies on extensive plant surveys
documented that floated free particles are a significant part of both issues. TOF-LIMS studies did consistently show, firstly, significant amounts of Pb on the surfaces of the sphalerite
and pyrite particles and, secondly, that this Pb had a major effect on the flotation of both
minerals (Kim, Chryssoulis, and Stowe 1995). The surfaces of sphalerite and pyrite from
competent run-of-mine (lump-size ore from the semiautogenous-grinding-mill feed) are
clean of surface contaminants in contrast to conveyor belt fines (<100 μm), which are
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39
SURFACE CHARACTERIZATION AND NEW TOOLS FOR RESEARCH
K
301
23
Surface
Na
Cu
63
65
208
56
Fe
Pb
50
100
200
.5 V/div
56
Fe
Subsurface 1
K
208
88
56
23
64
66
68
39
Pb
50 Fe
100
200
Subsurface 2
0
50
208
88
66
68
39
23
64
Zn
100
a.m.u.
200
FIGURE 10 Successive TOF-LIMS spectra obtained from the uppermost surface layers of Cuactivated sphalerite grain by firing the ablator laser at the same site three times
Grinding
Cu-Pb Flotation
Zinc Flotation
Tailings
Zinc Concentrate
Copper Separation
Lead Upgrading
Bulk Flotation
Lead Concentrate
Copper Concentrate
Pyrite Concentrate
(to Tailings)
FIGURE 11
Bulk (Zn-Pb)
Concentrate
Brunswick mill flowsheet
inadvertently activated. Pre-activation by lead ions originating from galena begins as soon as
the minerals are contacted with water (Kim, Chryssoulis, and Stowe 1995). Most of the
lead-activated sphalerite and pyrite that floated in the primary Cu-Pb circuit followed
galena to the Cu-Pb concentrate separation circuit. On the other hand, the lead-activated
pyrite and sphalerite that was successfully removed in the Cu-Pb cleaners refloated in the
zinc roughers. The presence of lead on sphalerite surfaces complicated activation by copper
for zinc flotation, while sphalerite activated solely by lead was selectively rejected in the zinc
cleaners as a result of the lime conditioning. Enlightened by these studies and following successful pilot-plant testing, galena flash flotation was introduced in 1998 with remarkable
results both in the Cu-Pb and Zn circuits.
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Probably the most important family of flotation chemicals is pH modifiers. A plant
study of zinc rougher flotation at Brunswick was conducted using lime, soda ash, and a combination of both as the pH modifiers in order to establish the effects on sphalerite and pyrite
flotation kinetics and on mineral selectivity. TOF-LIMS surface microanalysis was one of
several microbeam techniques used in the investigation. The results revealed the important
role of carbonate ion (from the soda ash) in cleaning and more specifically in removing lead
from the sphalerite surfaces for effective activation by copper sulfate. The clear advantage of
the soda ash–lime combination was to maintain the benefit of the additional Cu activation
on sphalerite while at the same time increasing the effectiveness of the depressing action of
Ca on the pyrite by increasing coverage (Nesset et al. 2001). The Brunswick concentrator
now uses both soda ash and lime in their zinc circuit. The sequence of the soda ash, copper
sulfate, and lime additions can be important, because Ca may be blocking sites for Cu sorption on pyrite. Ca is added after Cu (Kim et al. 1997).
Boliden Cu-Pb-Zn Concentrator
When treating Petiknäs South ore at the Boliden concentrator (Sweden), pyrite floating to
the Cu-Pb rougher concentrate is largely liberated and is most abundant in the highly floatable 20–40-μm size range, indicative of true flotation. TOF-LIMS analysis revealed that its
flotation could in part be ascribed to a plethora of activating ions such as Cu and Pb on the
particle surfaces, together with a relative scarcity of depressant ions such as Ca and Fe. Several conditions were investigated in the laboratory with the intent of reducing the surface
activators while enhancing the surfaces with depressants. Of these, only the synergistic use of
lime and ferric sulfate showed any promise, and initial flotation tests have shown that these
may indeed control pyrite flotation (Bolin, Chryssoulis, and Martin 1997).
Fine Sulfide Particles
TOF-LIMS, because of its tiny 1–3-μm analytical spot size, was used in two cases to identify
differences in surface composition between coarser (>20 μm) and fine (<5 μm) particles of
pentlandite from Inco’s Clarabelle mill (Canada) (Chryssoulis et al. 1991) and of galena
from MIM’s lead-zinc mill that could be partly responsible for the slower flotation kinetics
of the finer particles. In both cases, depressants were found in higher concentration on the
finer particles.
Gold Flotation
Free gold particles of floatable size classes (7–150 μm) in general account for <10% of the
gold losses in final flotation tails (Chryssoulis, Dunne, and Coetzee 2004). Evidently, surface modifiers played a role in their rejection. Comparative statistical analysis of TOF-LIMS
data on floated versus rejected free gold from 10 flotation plants found that surface compositional changes compromised gold floatability in a number of ways (Figure 12). Excessive
sorption of hydroxyl and calcium ions has been the most common cause for free gold rejection;
this mostly occurring in the cleaner circuit. Ramping up the pH in the cleaners is one way to
moderate this detrimental effect, another being supplementary addition of collector in the
regrind at modest dosages.
The concentrator at Los Pelambres is a good example of a process where surface characterization of free gold particles by TOF-LIMS fostered an increase in gold recovery. The
porphyry copper ore assaying 1.0% Cu, 0.02% Mo, with 0.03–0.05g Au/t, is processed using
a standard Cu-Mo flotation circuit at a rate of 180,000 tpd. Gold deportments on 12-hour
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S
303
12
(10)
(7)
8
Cu
(5)
Activators
6
Pb
4
(2)
Enrichment
10
Ag
2
–2
–4
Depressants FeOx
(10)
Mg CO3 Ca
(10)
(4)
(9)
PO4
(3)
–6
–8
AsO4
(9)
Depletion
0
OH –10
(10)
FIGURE 12 Ranking of inorganic surface modifiers affecting gold floatability, based on
comparative analysis of floated and rejected free native gold and electrum particles from the
concentrators. Numbers in parentheses give frequency of occurrence.
composite samples of the rougher and cleaner scavenger tails assaying 0.010 and 0.021g Au/t,
respectively, established that approximately one-third of the gold losses were in the form of
free gold particles in floatable size classes. Comparative surface microanalysis of floated and
rejected free gold particles revealed that the surfaces of rejected gold grains had systematically less sulfur and silver, though they were enriched in lead and iron oxyhydroxides
(Chryssoulis 2001). Based on these findings, a short flotation program was designed to evaluate two reagents: NaHS and 3418A. Sodium hyposulfite was chosen to convert lead carbonate on the surface of rejected free gold grains to sulfide, whereas 3418A is known to be
an excellent galena and silver mineral collector. Following plant trials at modest dosages,
20g/t NaHS and 0.5g/t 3418A, overall gold recovery was improved by 7 percentage points.
The last example illustrates the complementary use of microscopy and detailed inorganic/organic surface microanalysis to characterize the factors that control the flotation
kinetics of free gold. Surface microanalysis of equant and large flaky gold particles from concentrates of staged rougher flotation tests on a pyritic free milling gold ore determined progressively lower concentrations of the collector monomer on gold from each followup
concentrate. Microanalysis also revealed a direct proportionality between the surface concentration of silver and the collector monomer (Figure 13). These two observations confirm
that silver activates gold flotation and that collector monomers play a significant, previously
not fully recognized, role in gold flotation. It was also determined that large gold flakes
require four times as much collector in order to float at the same rate as equant gold grains.
INFRARED SPECTROSCOPY
Selective separation of mineral components is achieved by addition of specific collectors,
activators, depressants, or modifiers and manipulation of solution conditions (i.e., pH, Eh,
aeration). All these changes in solution composition are performed to make the valuable
mineral components very hydrophobic, whereas gangue components should remain hydrophilic. The flotation behavior of ore components depends on the nature and structure of the
surface hydrophobic or hydrophilic species produced. The possibility of monitoring surface
phenomena at molecular and atomic levels at the interfaces of natural minerals contacted
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FLOTATION FUNDAMENTALS
Collector Loading on Gold Grains, arbitrary units
400
DIBDTP
300
200
PAX
100
0
1
3
5
7
Silver on Surface of Gold Grains, arbitrary units
FIGURE 13 Correlation of DIBDTP and potassium amyl xanthate (PAX) loadings with surface
concentration of silver (all in arbitrary units) on gold grains from sequential concentrates C1–C5
of a staged rougher test. DIBDTP (1 g/t) and PAX (12 g/t) were added in the mill.
with aqueous solution is vitally important for the understanding of the surface phenomena
that govern selective and efficient flotation. This requires application of the appropriate
experimental techniques.
Infrared spectroscopy is particularly well suited to determine the surface composition
of minerals with adsorbed collectors at a molecular level. Infrared absorption is functional
group selective, so it is particularly well adapted to detect small changes of the molecular
microenvironment properties emerging at the interfaces. Experiments can be performed in
situ at both gas–mineral and aqueous solution–mineral interfaces. There is only very gentle
interaction of the infrared beam with the sample examined to ensure nondestructive analysis.
Infrared External Reflection Technique
In recent years, an infrared external reflection technique was developed ( J.A. Mielczarski
and Yoon 1989; J.A. Mielczarski 1993; J.A. Mielczarski and Mielczarski 1995, 1999; E. Mielczarski,
Duval, and Mielczarski 2002) that offers the ability to overcome experimental problems and
collect reliable data to monitor and understand surface phenomena at any mineral interface
at a molecular level in close-to-real flotation conditions. The variety, precision, and reliability of information about interface phenomena delivered by this technique are superior to
any other single technique.
The developed technique is based on specific interaction of electromagnetic waves with
a multilayer system. A schematic diagram for a simple three-phase system is presented in Figure 14. For polarization perpendicular to the plane of the incident beam (s-polarization),
there is only one electric vector, E ⊥Y , parallel to the substrate plane. Hence, only molecular
groups of the adsorbed species with a dipole transition moment parallel to the interface in y
direction can interact with the incident radiation and produce an absorbance band, A ⊥Y .
For example, in the case of the adsorption of xanthate molecule involving both sulfur atoms
with the same distance from interface (Figure 14), it is expected that the absorbance, A ⊥Y ,
will show the highest value. No interaction and, obviously, no absorbance band are observed
flotation0.book Page 305 Tuesday, January 2, 2007 7:36 PM
SURFACE CHARACTERIZATION AND NEW TOOLS FOR RESEARCH
CH3
EII
y
Va
x
Vas
Y
EII Z
EII X
d
C
S
θ
E
T
O
s
CH2
z
305
n1, k1, Ambient
n2, k2, Adsorption Layer
n3, k3, Mineral
S
FIGURE 14 Schematic diagram of the interaction of electric field vectors in three directions
with a simple three-phase system. Also shown is a xanthate molecule with dipole transition
moments shown for the asymmetric stretching vibrations of the SCS group (parallel to
interface) and of the asymmetric stretching vibrations of the COC group (almost vertical to
interface).
when the sulfur-carbon-sulfur (SCS) group is turned 90° from the position presented in Figure 14, and the dipole transition moment of the asymmetric vibration of the SCS group
becomes vertical to interface. For parallel polarization (p-polarization), there are two electric field vector components at the interface: one parallel, EIIX , and one perpendicular, EIIZ ,
to the substrate plane. Using the example of xanthate molecules at the surface presented in
Figure 14, two absorbance bands—the AIIX for the SCS vibration group and the AIIZ
related to the carbon-oxygen-carbon (COC) group vibration—will be present in the
recorded spectrum. It is also possible to distinguish these two components because they
show reverse absorbance; the first one produces negative absorbance, and the second one is
positive.
The incident infrared beam reflected from a mineral surface carries all the information
about surface composition and structure. With proper manipulation of the experimental
optical conditions (incident angle and polarization), it is sufficient to record three spectra,
which together give a three-dimensional “picture” of the species present at the mineral surface. Importantly, simulation of the adsorbed layer is carried out before any experiment is
performed, allowing prediction of the best experimental conditions that give optimal spectral sensitivity and the maximum confidence in the interpretation of experimental results.
This also significantly speeds up the experimental procedure.
The developed technique has unique properties compared with other known infrared
techniques, such as transmission, diffuse reflectance, attenuated total reflection, and photoacoustic spectroscopies. This technique, supported by spectral simulation, allows almost all
the details about the mineral–aqueous solution interactions to be obtained, including the
• Nature of the adsorbed products, and by which molecular group adsorption takes place
• Adsorbed quantities of different surface products (starting from 20% of monolayer)
• Surface distribution of the adsorbed species (uniform layer or patches with determined thickness)
• Molecular orientation of the adsorbed species (through orientation of particular
functional groups)
• Molecular recognition (selective adsorption on specific surface sites)
• Lateral interaction between the adsorbed collector molecules
• Dynamic phenomena, such as kinetics of adsorption/desorption, stability of surface
products, and surface mobility of the adsorbed species
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306
FLOTATION FUNDAMENTALS
(a) adsorption
(b) conditioning
s-polarization
1125
1,034
1,024
70º
1,200 1,000
Wavenumber, cm
–1
1,200
1,126
1,050
1,040
0.001
1,190
1,127
1,049
1,194
1,225
1,198
1,126
1,049
1,200 1,000
1127
C
S
68°
85º
85º
85°
1049
12251126
43°
S
1,035
70°
70º
0.005
.005
CH2
O
1,050
1,042
1201
1128
1050
CH3
p-polarization
1,126
1,201
1,128
1,050
–log (R/Ro)
p-polarization
× 2.5
11981125
–log (R/Ro)
20º
20°
1028
1198
1033
1,198
1,028
1,033
× 2.5
.5
1,198
1,196
1,125
1,125
1,051
1,034
s-polarization
1,204 1,196
20º
× 2.5
Cu2S
1,000
Wavenumber, cm
–1
FIGURE 15 Results of infrared external reflection technique: (left) reflection spectra of
xanthate on chalcocite, solution of 5 × 10–5 M and pH 9.2: (a) after 3 minutes of adsorption and
(b) after 30 minutes of conditioning; (center) simulated reflection spectra of an isotropic 1-nm
layer for different polarization and incident angles; (right) determined molecular arrangement
of xanthate on chalcocite surface after conditioning.
This information leads to the determination of the mechanisms and dynamics of surface phenomena, which allow the design of selective and efficient flotation, and control of
this process.
The developed infrared external reflection technique allows the unique ability for study
of interface phenomena at a molecular level on heterogeneous substrates such as natural
minerals. There is no limit; all types of mineral samples can be investigated. The experiments are fast and nondestructive. High sensitivity, in-situ-collected information in a multiphase system, including the region of strong absorption of substrate, makes this technique
a very valuable experimental tool. The complexity of the recorded reflection spectra, their
sensitivity to any variations of the optical properties of all bulk and surface components, and
their spatial distribution in the system under investigation are, in fact, the major strengths of
the technique.
Examples of results are presented in Figure 15. Experimental spectra recorded at three
angles of incident beam and p-polarization (Figure 15, left) are compared with simulated
spectra (Figure 15, center), and the similarity and differences (for example at about 1,040 cm–1)
indicate that the hydrophobic product is cuprous xanthate; that the adsorbed molecules are
at first randomly oriented (Figure 15, left); and that after conditioning, they reorganize
themselves in an oriented layer (Figure 15, center) with the molecular arrangement presented in Figure 15 (right) ( J.A. Mielczarski et al. 1995) These results clearly explain the
increase in hydrophobicity during conditioning and rationalize the conditioning procedure.
Another application is the monitoring and interpretation of the galvanic effect between
two or more grains of different minerals. The detected changes are dramatic—from no
adsorption to several monolayers of collector coverage formation, with major consequences
for pyrite depression and galena flotation when they are together in collision in flotation
pulp. However, if these two minerals are separately contacted with xanthate solution, pyrite
is more hydrophobic than galena (Figure 16). These results clearly indicate that the observations
and conclusions about the collector action—hence, also the surface hydrophobization—when
flotation0.book Page 307 Tuesday, January 2, 2007 7:36 PM
67 mV
0.01
0.01
1,269
307
Single
Minerals
-log(R/
Potential
(OCP) of
Pyrite in
Xanthate
Solution
(SHE)
1,047
1,031
)
Ro
Potential
(OCP) of
Galena in
Xanthate
Solution
(SHE)
–log (R/Ro)
–log (R/Ro)
a
1,046
1,227
SURFACE CHARACTERIZATION AND NEW TOOLS FOR RESEARCH
)
a
Ro
a
195 mV
-log(R/
0.01
0.01
b
b
Minerals in
130 mV Collision
b
b
120 mV
X× 0.5
0.5
1,200
Wavenumber, cm
1,000
–1
1,200
Wavenumber, cm
1,000
–1
Source: E. Mielczarski and Mielczarski 2003a.
FIGURE 16 Reflection spectra of galena and pyrite after contact with amyl xanthate solution:
(a) single minerals; (b) minerals in galvanic contact by collision. Open circuit potential (OCP) is
reported vs. saturated hydrogen electrode (SHE).
made for single mineral systems do not describe the real situation in a multicomponent mineral system such as that in flotation pulp.
Numerous fundamental and practical questions have been answered with the help of
this technique. This multidiagnostic technique has been applied extensively to the study of
interaction of different sulfide minerals with various aqueous solutions including chalcocite
( J.A. Mielczarski et al. 1995; J.A. Mielczarski, Xu, and Cases 1996), chalcopyrite ( J.A. Mielczarski et al. 1995; J.A. Mielczarski, Mielczarski, and Cases 1997, 1998), tenantite, tetrahedrite ( J.A. Mielczarski et al. 1995; J.A. Mielczarski, Mielczarski, and Cases 1997), galena,
and pyrite (E. Mielczarski and Mielczarski 2003a, b). The collectors investigated were different xanthates, carbaminates, and phosphates. Major achievements are as follows:
1. Detailed description of solution conditions to produce selectively hydrophobic
mineral surfaces,
2. Observation of strong interactions between different minerals in multicomponent
systems and their influence on selective flotation,
3. Explanation of the role of particular reagents in the flotation system and the best
surface modification conditions, and
4. Understanding of the difference in hydrophobization action between different
xanthate-type collectors.
This information has allowed important full-industrial-scale technological improvements to be proposed that have already brought benefits to the user (e.g., J.A. Mielczarski
et al. 1999). Similarly, extensive research has been performed on selective hydrophobization
of semisoluble minerals such as fluorite (E. Mielczarski, Mielczarski, and Cases 1998; J.A.
Mielczarski, Mielczarski, and Cases 1999; E. Mielczarski et al. 2002), apatite ( J.A. Mielczarski and Mielczarski 1995), calcite ( J.A. Mielczarski and Mielczarski 1999), quartz ( J.A.
Mielczarski et al. 1995), and oxides (E. Mielczarski et al. 2004). A more general description
of the applications of infrared external reflection techniques focused on the understanding
and modification of surface phenomena (hydrophobicity) can be found in recent papers
(E. Mielczarski and Mielczarski 2003b, 2005).
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308
FLOTATION FUNDAMENTALS
A
B
FIGURE 17 Internal reflection configurations: (a) with mineral particles deposited on
transparent reflection element, and (b) with adsorbed layer on reflection element made of
transparent mineral
Infrared Internal Reflection Spectroscopy
Another infrared reflection technique, which ensures a better understanding of mineral surface phenomena that govern selective flotation, is internal reflection spectroscopy (IRS),
also known as the attenuated total reflection (ATR) technique. Two experimental
approaches have been explored. In the first approach, the mineral grains in contact with
solution are in close vicinity to the ATR element (Figure 17a). In this case, a reflection spectrum of the outermost mineral surface layer could be recorded. In the second configuration,
the investigated adsorption layer is produced directly on a reflection element made of natural or synthetic mineral (Figure 17b). The adsorption layer is also in contact with aqueous
solution during measurement.
The latter approach is limited to the minerals that are transparent for infrared radiation, such as fluorite (CaF2), sphalerite (ZnS), and so forth. The first approach can be used
for any type of mineral grains. Some examples of mineral grain surface characterization
could be found in reports by Mielczarski and colleagues ( J.A. Mielczarski 1986; J.A.
Mielczarski, Nowak, and Strojek 1980, 1983; Leppinen and Mielczarski 1986). The
recorded in-situ spectra carry the information about mineral surface composition and relative amounts of the determined surface species. The second approach is described in more
detail in the following paragraphs.
In-situ quantitative measurements of interfacial phenomena, such as surfactant adsorption at mineral surfaces using FTIR/IRS, have been achieved in the last two decades (Sperline, Muralidharan, and Freiser 1987; Jang and Miller 1993; Fa and Miller 2003; Kellar,
Cross, and Miller 1989). The results of these quantitative measurements greatly improved
the understanding of surfactant adsorption at various surfaces. The basic experimental setup
for FTIR/IRS sampling is shown in Figure 18. In the sampling region, three optically distinct phases are distinguished by their different refractive indices, ni, and absorption coefficients, ki, where i = 1, 2, 3 (Figure 18, left). The infrared (IR) light enters the interfacial
region from a dense phase (refractive index n1) to a rare phase, n3 (n1 > n3). The interfacial
region to be probed is the second phase (n2) sandwiched in between the first (incident) phase
and the third phase. When the incident angle is larger than the critical angle θc = sin–1 (n3/n1),
total internal reflection occurs, and an evanescent optical field is generated in phase 3 (Figure 18, right). The guided IR light inside the infrared element (IRE) experiences multiple
reflections at both sides of the IRE crystal. This ensures a high signal-to-noise ratio that is
important for quantitative analysis of adsorption reactions.
The simplest quantitative measurement is a comparison of the spectra of unknown
samples with those of known samples. The molecular packing angle (Simon-Kutscher, Gericke, and Huhnerfuss 1996; Jang and Miller 1995), adsorption density ( Jang and Miller
1993; Jang and Miller 1995; Tejedor-Tejedor and Anderson 1990), and film optical constants (Sperline, Muralidharan, and Freiser 1987; Jang and Miller 1993; Fa and Miller 2003;
flotation0.book Page 309 Tuesday, January 2, 2007 7:36 PM
SURFACE CHARACTERIZATION AND NEW TOOLS FOR RESEARCH
n1 > n3
θ
n3
Evanescent Field
Incident Phase, n1, k1
n1
d
Film, n 2, k 2
θ > θc
θc = sin–1 (n 3/n1)
n1 > n3
θ
Third Phase, n 3, k 3
θ
FIGURE 18
309
θ
Typical FTIR/IRS experimental setup
Kellar, Cross, and Miller 1989; Simon-Kutscher, Gericke, and Huhnerfuss 1996; Jang and
Miller 1995; Tejedor-Tejedor and Anderson 1990; Buffeteau et al. 1999; Lee and Sung
2001; Ren and Kato 2002; Buffeteau et al. 2001; Tickanen, Tejedor-Tejedor, and Anderson
1997) have been quantitatively measured from the FTIR/IRS spectra. One advantage of
FTIR/IRS over other sampling techniques is that surfactant adsorption at the IRE surface
can be studied by in-situ experiments. Generally, the advantage of in-situ investigation of
surfactant adsorption is that it offers both the adsorption kinetics and chemical information
regarding the adsorption state.
The adsorption density equation was originally derived by Sperline, Muralidharan, and
Freiser (1987) and later modified and applied to flotation chemistry research by Miller and
co-workers ( Jang and Miller 1993; Kellar, Cross, and Miller 1989). The adsorption density
equation in one form is
A – NC b ε d e
Γ = 10 7 ----------------------------------------------------N ε ( 2d e ⁄ d p + p ⁄ cos φ )
(EQ 2)
where
Γ=
A=
N=
Cb =
ε=
de =
dp =
p=
adsorption density, μmol/m2
integrated absorbance, cm–1
Avogadro’s number
bulk solution concentration, mol/L
absorptivity, L/mol·cm2
effective thickness of the sample, cm
penetration depth, cm
number of adsorbate-coated IRE entrance and exit faces
Adsorption density measurements using the FTIR/IRS technique (Figure 19) have shown
that at a fluorite surface, monolayer coverage can be achieved at very low initial oleate concentrations (below 1 × 10–6 M). Such behavior is characterized by a monolayer plateau in
the adsorption isotherm but is not the case for calcite. For fluorite, chemisorption is clearly
the dominant adsorption mechanism at low oleate concentrations with oleate bonding
directly with calcium sites at the fluorite surface. Chemisorption also occurs to a more limited extent (incomplete monolayer) at the surface of other calcium minerals, such as calcite
and apatite for low oleate concentration. On the other hand, at higher concentrations of oleate
and with calcium ion, the growth of multilayers and/or the physical adsorption of calcium
dioleate collector colloids are observed from the adsorption isotherm measured using
FTIR/IRS spectra.
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310
FLOTATION FUNDAMENTALS
Adsorption Density, μmol/m 2
1,000
100
10
1
Calcite
Fluorite
0.1
10–6
10–5
10
–4
10–3
10–2
Equilibrium Oleate Concentration, M
FIGURE 19 Adsorption isotherms of oleate at semisoluble mineral surfaces as determined by
direct FTIR/IRS analysis of the surface
Of course, in-situ FTIR/IRS has also been used extensively for qualitative studies of
surfactant adsorption at mineral surfaces (Young and Miller 2000; Free and Miller 1996; Lu
and Miller 2002). In these studies, high S/N ratios were obtained. These studies have also
provided considerable information regarding surfactant adsorption thermodynamics and
kinetics. However, it is difficult to obtain information regarding interfacial water structure
from in-situ FTIR/IRS experiments (Hancer, Sperline, and Miller 2000).
O T H E R V I B R AT I O N A L S P E C T R O S C O P Y T E C H N I Q U E S
Vibrational spectra, as well as infrared spectroscopy, can be obtained by other spectroscopic
techniques such as Raman and sum-frequency generation (SFG). Although it is difficult to
obtain detailed information about natural mineral surface composition at monolayer coverage by these experimental techniques, they present other advantages, which are described in
the following sections.
Raman Spectroelectrochemistry
Raman spectroscopy is an in-situ technique that is well suited to the investigation of a mineral surface. In Raman spectroscopy, the spectrum of the collected light comprises a series of
discrete bands at frequencies higher to (anti-Stokes) or lower than (Stokes) the elastically
scattered (Rayleigh) laser line. These lines are said to be Raman shifted from the Rayleigh
line, and the shift is equal to vibrational energy of the transition. Minerals may exhibit vibrational spectra because of the structure of the molecular units that they contain as in the case
of S–S in pyrite, or the spectra may be derived from the vibrational structure of the mineral
crystal (sphalerite) (Hope, Woods, and Munce 2001). Many common materials (e.g., glass
and water) are transparent to visible and near-infrared radiation, and they are weak Raman
scatterers in the wavelengths where the majority of laser excitation sources operate. The
dependence of the Raman shift on the vibrational and rotational transitions of the scattering
molecules means that some minerals may not exhibit a significant Raman spectrum, galena
and chalcocite being Raman-inactive sulfide minerals.
Raman scattering is an inefficient process with less than 1 in 106 photons undergoing a
Raman interaction. Typically, this requires a sample thickness of 5 to 50 nm for the spectra
flotation0.book Page 311 Tuesday, January 2, 2007 7:36 PM
SURFACE CHARACTERIZATION AND NEW TOOLS FOR RESEARCH
311
to be detectable with the most sensitive modern commercial instruments. In spectroelectrochemical studies of collector adsorption, electrodes of gold, silver, or copper have been used
for surface-enhanced Raman scattering (SERS). This technique enables the formation of a
monolayer of adsorbed collector molecules to be characterized. SERS has been used to
investigate the adsorption of a range of sulfide mineral flotation collectors: ethyl xanthate
(Woods and Hope 1998; Woods, Hope, and Brown 1998a, b); isopropyl, isobutyl, and
isoamyl xanthate (Hope, Watling, and Woods 2001b); O-isopropyl-N-ethylthionocarbamate
(Woods and Hope 1999); 2-mercaptobenzothiazole (Woods, Hope, and Watling 2000;
Hope, Watling, and Woods 2001a); di-isobutyldithiophosphinate (Hope et al. 2003; Hope,
Woods and Watling 2003); and butyl ethoxycarbonylthiourea (Hope , Woods and Watling
2001; Hope et al. 2004). In the case of the xanthates, the collector adsorption can be controlled using the electrode potential. Spectra can be obtained from surfaces prior to collector
adsorption, with partial monolayer coverage and for multilayer coatings. This behavior can
also be observed with thiocarbamates. The Raman spectra of these collectors were consistent with adsorption of the collector molecules on the surface and intact, through a collector sulfur group. Decomposition of a collector was only observed under applied potentials
greater than the solution potentials encountered in typical sulfide mineral flotation.
Mercaptobenzothiazole, di-isobutyldithiophosphinate and butyl ethoxycarbonylthiourea
flotation reagents adsorbed onto the metallic electrode surface throughout the accessible
potential range. The Raman spectra of the deposits formed on the electrode surface were
very close to the spectra of the relevant metal compounds prepared in bulk. Metal-collector
compounds could be extracted from the electrode surface after extended reaction times and,
when characterized by Raman, FTIR, and nuclear magnetic resonance, were found to be the
same as the bulk compounds.
Results indicate that sulfide mineral flotation reagents can act either through the
chemisorption of collectors (xanthates and thiocarbamates), or through the formation of a
metal compound on the mineral surface (mercaptobenzothiazole, di-isobutyldithiophosphinate and butyl ethoxycarbonylthiourea).
Sum-Frequency Generation
SFG was first introduced in 1987 by Shen’s research group (Zhu, Suhr, and Shen 1987).
Since then, SFG has been further developed into a surface-specific vibrational spectroscopy
and used to study the vibrational modes and orientations of molecules and monitor reactions at interfaces. SFG is a nonlinear optical process, where the signal is generated at a frequency that is the sum of the frequencies of two incident optical fields due to the nonlinear
interaction of infrared and visible lasers as shown in Figure 20 (Shen 1989). SFG is forbidden in a medium with inversion symmetry, but this nonlinear optical process occurs at surfaces where the inversion symmetry is broken. Most bulk materials have inversion
symmetry; thus, they do not generate SFG signals. This unique feature makes SFG a sensitive and powerful tool for the study of various interfaces and surfaces (Nickolov, Wang, and
Miller 2004).
The SFG spectrum of molecules with long hydrocarbon chains is very sensitive to
hydrocarbon chain order—loosely packed chains and disordered monolayers will, in general, have more random orientations of the methyl and methylene groups, and the intensity
of the SFG signal will be much smaller than in the case of an all-trans state. When an alkyl
chain is in an all-trans conformation, it is locally centrosymmetric around the C–C bond,
and the CH2 symmetric stretching mode at ~2,850 cm–1 is SFG inactive and does not
flotation0.book Page 312 Tuesday, January 2, 2007 7:36 PM
312
FLOTATION FUNDAMENTALS
ωvis
ωsum
ωIR
Sample Surface
FIGURE 20
SFG generation
appear in the spectra. When the chains are disordered because of gauche defects, this local
inversion symmetry is lifted, and a peak at 2,850 cm–1 will appear in the spectra to an extent
that depends on the degree of disorder. The ratio of intensity of the CH3 symmetric stretching to CH2 symmetric stretching can be used to provide a relative measurement of hydrocarbon chain ordering (Conboy et al. 1997). The peak ratio method, which directly depends
on intermolecular and intramolecular symmetries, therefore provides the most quantitative
measure of chain conformation in the monolayer (Smiley and Richmond 1999). In a closely
packed surfactant monolayer in all-trans conformation, there are only two dominant bands
at 2,875 cm–1 and ~2,940 cm–1 that correspond to the symmetric stretching vibration and
the symmetric stretching Fermi resonance with a bending overtone of the CH3 group, respectively.
In the following discussion, the acronyms SSP and SPS (where S and P represent polarization that is perpendicular and parallel, respectively, to the incident plane) designate the
state of polarization of the beams in the following order: output sum frequency beam, input
visible beam, and input IR signal beam. For example, Figure 21 shows the spectrum for an
oleic acid Langmuir–Blodgett (LB) film at the surface of fused silica (Wang 2004). The
spectrum suggests that the oleic acid forms a closely packed monolayer in all-trans conformation because there are only two dominant bands at 2,878 cm–1 and 2,940 cm–1 which
correspond to the symmetric stretching vibration and the symmetric stretching Fermi resonance with a bending overtone of the CH3 group, respectively, in the SSP polarization spectrum. Shown in Figure 22 are the spectra taken from the interface of CaF2 and a hydroxamic
acid D2O (10–3 M hydroxamic acid) solution (Wang 2004). The spectrum for the SSP
polarization state is similar to that for a monolayer of hydroxamic acid at a fused silica surface
(Wang 2004). The CH2 symmetric stretching is observed at 2,850 cm–1. The peak at
2,877 cm–1 is due to the CH3 symmetric stretching. The Fermi resonance of CH3 symmetric stretching appears at 2,940 cm–1. A small peak at 2,920 cm–1 is assigned to CH2 asymmetric stretching. For the SPS polarization state, the major peak is due to CH3 asymmetric
stretching at 2,961 cm–1. The peak from the CH3 symmetric stretching appearing at 2,880
cm–1 for the SPS polarization state indicates that the CH3 symmetric axis has a certain angle
with respect to the surface normal. The SFG intensity from the CH2 gauche structure is
greater than that for the LB-transferred monolayer. The strong CH3 symmetric stretching
peak that appears in the SPS spectrum means there is a significant tilt angle between the C–
CH3 axis and the surface normal.
flotation0.book Page 313 Tuesday, January 2, 2007 7:36 PM
4.5
9
4.0
8
SFG Intensity, arbitrary units
SFG Intensity, arbitrary units
SURFACE CHARACTERIZATION AND NEW TOOLS FOR RESEARCH
3.5
3.0
2.5
2.0
1.5
1.0
0.5
0.0
2,800
2,850
2,900
2,950
3,000
Wavenumber, cm –1
FIGURE 21 SFG spectra for oleic acid LB
film at the surface of fused silica under SSP
polarization conditions
313
7
6
SSP
5
4
3
2
SPS
1
0
2,750
2,850
2,950
3,050
Wavenumber, cm –1
FIGURE 22 SFG spectra taken from
hydroxamic acid at D2O solution/CaF2
interface SSP and SPS polarization
conditions
M I C R O S C O P Y A N D M I N E R A L O G Y : PA S T, P R E S E N T,
AND FUTURE
For many years, traditional descriptive mineralogical examination has been performed using
polished sections of samples. Determination of the mineral quantities (modal analysis) and
liberation analysis in ores and mineral processing samples has been a basic requirement. The
era of quantitative mineralogy (quantitative modal analysis and quantitative liberation analysis) began in 1848 when Delesse published a method to determine the mineral quantities
based on the areas of the minerals of interest in a polished section. The method was very
labor intensive: it involved tracing the outlines of the mineral grains onto a cloth, sticking
this cloth onto a tin foil, cutting the tin foil, sorting according to the different minerals, and
then measuring the relative weighting of each group of foil cutouts. The ratio of the weights
of the tin foil cutout minerals to the weight of the original tin foil was proportional to the
area of that mineral in the sample surface and was correlated to the volume percent and the
weight percent.
In 1898, Rosiwal published an improved method to determine the mineral quantities
in polished sections. Rosiwal’s method, which became known simply as “cord analysis,”
involved a grid of parallel lines covering the image of the specimen surface that were used to
determine the linear length of intercepts onto each mineral of interest. From the addition of
the linear intercept in the mineral of interest and the total length of the lines, the linear percent of the mineral was calculated, which was correlated to the volume percent and the
weight percent.
In 1934, Glagolev published an improved method for determining minerals quantities.
Glagolev’s method, which became known simply as “point counting,” involved manual
counting of the points in a grid that coincided with a mineral of interest in a polished section observed with an optical microscope. The ratio of the points on a mineral of interest to
all points on the examination grid was equivalent to the volume fraction of that mineral,
which was correlated to the weight percent.
The advent of affordable computers in the 1970s allowed the development of systems
that improved the speed for determining the mineral quantities. Basically, the methods
developed by Delesse, Rosival, and Glagolev (area grade, linear grade, and point counting)
are behind the computerized systems used to determine the mineral quantities. The advent
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314
FLOTATION FUNDAMENTALS
of computers also provided the required speed to quantitatively determine the mineral liberation, which basically involves the measurement of the particle mineral composition. A
review of the capabilities of early image analysis is given in works by Martens, Morton, and
McCarthy (1978) and Taylor et al. (1978). Many of the developments of early image analysis were spurred by the needs of metallography.
In 1974, the CANMET Mining and Mineral Sciences Laboratories (CANMETMMSL) in Canada adapted a Quantimet system to perform quantitative mineralogy.
Petruk (1976) described this system, which was interfaced to an optical reflected light
microscope with a black-and-white video camera, and performed measurements of areas of
mineral grains. Other systems interfaced to optical microscopes were used in other parts of
the world. However, system limitations were soon recognized in that many minerals could
not be properly discriminated.
Very early in the development of systems for quantitative mineralogy, it was realized
( Jones 1984) that minerals could be better identified automatically by using scanning electron microscopes (SEMs) or electron microprobes (EMPs) with X-ray detectors for elemental analysis. Si(Li) energy-dispersive spectroscopy (EDS) detectors or wavelength-dispersive
spectroscopy (WDS) detectors were used.
The system developed by Jones at Imperial College, London, United Kingdom ( Jones
1984), was based on an electron microprobe (Camebax) with WDS detectors. The system
identified the minerals using linear scans at a pixel size of ~2 μm and the elemental X-ray
signals from the WDS detectors. Measurement time was ~10 msec per pixel if the mineral
was identified from the presence or absence of an element, and 100 msec if more complete
elemental information was required. The identification of a single mineral grain requires
acquiring and processing information from10 to 30 pixel spots.
The QEM*SEM system was developed by the Commonwealth Scientific and Industrial
Research Organization (CSIRO) in Australia. It was initially presented by Grant and colleagues in 1976. The present configuration, commercially available, is called QEM*Scan.
The system is based on a SEM and identifies minerals mainly based on their elemental X-ray
counts obtained from EDS detectors. EDS detectors have a dead time that is a function of
the count rate. The first EDS detectors had analog pulse processors and could process a
maximum of 25,000 counts per second (cps). Although it was possible to purchase
QEM*SEM systems with one EDS detector, to increase its speed the first full configuration
of the QEM*SEM had four EDS detectors with analog pulse processing. In the 1990s, EDS
detectors with digital pulse processors became available and could process a maximum of
~50,000 cps (the actual count rate of an EDS with analog pulse processors is ~2,000 cps,
whereas the actual count rate for an EDS detector with digital pulse processor is ~20,000 cps).
The present configuration of the QEM*Scan has four EDS detectors with a digital pulse
processor. In general, the system uses a grid of points that are superimposed on the image of
the particles. From each of these stop points, the system acquires elemental X-ray information from the group of EDS detectors. The EDS information is compared with a database of
minerals, and the mineral at the point is identified. The acquisition time per stop point is
fast, ~10 msec. However, because a mineral grain may comprise 10–30 stop points, the overall speed of the system to process several thousands of mineral grains is slow. To partially
address speed concerns, the system has different operational modes to perform bulk mineralogical analysis or to perform a more detailed mineralogical analysis including the liberation analysis. In general, the main difference is the spacing of the stop points to acquire EDS
X-ray data.
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In 1984, a system was integrated at the CANMET-MMSL (Petruk 1988) to determine
mineral quantities and liberation analysis by measurements of areas of mineral grains and
particles. The system was based on a programmable generic image analyzer interfaced with
an electron microprobe ( JEOL 733). The highly stable electron beam current in the microprobe allows the system to identify the minerals using mainly the backscattered electron
(BSE) image. Acquisition of BSE images is much faster than acquisition of X-ray elemental
information. The first generation of the system (1984) was based on an IBAS image analyzer (Kontron) hosted in an MCP computer (Motorola); this system allowed identification
and measurement of ~5,000 particles per hour. The last generation of the system is based on
a KS400 image analyzer (Zeiss Vision) hosted in a computer that allows identification and
measurement of ~100,000 particles per hour. There are ~5,000 known minerals; among
those are many minerals with similar average atomic numbers (Zav) and, thus, similar BSE
gray levels. However, a given sample does not have 5,000 minerals but probably only ~20
minerals and of those, commonly, only five minerals are of interest. Thus, for many samples,
it is possible to use only the information from the BSE image to perform the image analysis.
For example, BSE images allow discrimination between pyrrhotite (Fe1–xS, x~0.2 with Zav =
22.1) and troililite (FeS, with Zav = 22.4), and discrimination between magnetite (Zav =
21.0) and hematite (Zav = 20.1). There are some cases where there are minerals of interest
with overlapping BSE gray level. For these cases, the system at CANMET-MMSL can
obtain elemental X-ray information from any of its X-ray detectors. The speed of the system
is reduced when the image analysis requires the use of X-ray data. The first generation of the
system had two WDS detectors and one EDS detector. The image analyzer automatically
controls the electron beam and obtains the elemental X-ray information from any of the X-ray
detectors. The system can acquire X-ray information from the whole field, or from rectangular windows around the mineral grains or from stop points similar to the QEM*Scan.
Thus, the system can identify mineral grains by a combination of BSE imaging and elemental X-ray data. The system could also be operated in a mode fully similar to the QEM*Scan,
although it would rarely be necessary (i.e., identifying all minerals based on elemental X-ray
information from stop points). In this latter mode, the first generation of image analyzer at
CANMET-MMSL was slower than the QEM*Scan, because it used X-ray data from a single X-ray detector (either EDS or WDS).
WDS detectors have better energy resolution than EDS detectors (~5 eV vs. ~140 eV,
respectively; e.g., Goldstein et al. 1994). Thus, considering the whole population of ~5,000
existent minerals, there will be minerals whose elements overlap in the EDS spectra but can
be well resolved using WDS detectors. However, WDS detectors have focusing limitations
(i.e., the signal intensity is not homogeneous for the entire field of view at magnifications
lower than ~300×). X-ray information can be acquired very quickly by deflecting the electron beam to a required position at the field of view. This can be done for any magnification
using an EDS detector. However, this cannot be done for magnifications lower than ~300×
using the WDS detectors, because the count rate for a given mineral will be different at the
extreme corners of the field of view. It is possible to automatically control the stage by moving it to a given point to the center of the viewing field and obtaining the X-ray data from a
WDS detector. This is known as stage-scanning and is a common feature of modern controllers
of electron microprobes. The image analyzer at CANMET-MMSL can be programmed for
stage-scanning and using WDS data at low magnifications for minerals that cannot be discriminated by EDS information. However, the speed of the system is further reduced.
Another possibility is to acquire BSE images and WDS information at magnifications of
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300× or higher and then zoom down the images to a lower magnification that would be better suited to the size of the particles in the sample (Matos, Lastra, and Petruk 1996).
Since 1984, the operational philosophy with the system at CANMET-MMSL has been
to discriminate as many minerals of interest as possible using only BSE information and
combine the information from EDS or WDS only when needed. Since 1984, having analyzed several thousand samples, the overall experience at CANMET-MMSL has been that
~80% of the sample cases can be analyzed using BSE information only, ~10% can be analyzed by adding single elemental information from the EDS detector, ~5% can be analyzed
by adding multiple elemental information from the EDS detector, and another 5% of the
sample cases require adding information from the WDS detectors.
There is another advantage of performing quantitative mineralogy based on BSE imaging. This is related to the excitation volume than can be demonstrated from first physical
principles. X-ray signals from electron-based instruments do not come from an area on the
sample surface; rather, they come from an excitation volume. This excitation volume is a
function of the accelerating voltage of the electron beam and the density of the sample
point. In general terms, the diameter of the excitation volume for X-ray signals from minerals is ~3 μm at 20 kV of accelerating voltage. On the other hand, the BSE signal provides a
spatial resolution of ~0.1–0.2 μm. Thus, for grains smaller than ~3–5 μm, the X-ray signals
will give information that may not be appropriate to properly identify the corresponding
mineral. Similarly, because the spatial resolution of X-ray signals is ~3 μm for minerals, borders of mineral particles and borders of minerals within particles may not be properly identified.
In the late 1990s, QEM*SEM became QEM*Scan based on a LEO (Zeiss) SEM. At
about the same time, Kontron imaging was acquired by Zeiss, whereby Kontron interfaces
for SEM instruments and electron microprobes became unavailable, and the Julius
Kruttschnitt Mineral Research Centre ( JKMRC, Australia) developed the Mineral Liberation Analyser (MLA) (Gu 2003). The MLA is based on a SEM (FEI, formerly Philips), and
is commercially available. The MLA is similar to the system at CANMET-MMSL. The
MLA discriminates many minerals of interest using only BSE information and combines
with information from EDS only when needed to discriminate certain specific minerals.
When analyzing only BSE images, the system is faster than the QEM*Scan. The MLA has
up to two EDS detectors; therefore, when identifying minerals based solely on X-ray information, it is slower than the QEM*Scan.
In 1997, a new type of X-ray energy-dispersive detector for SEM instruments and electron microprobes was introduced. This new technology is based on the silicon drift chamber (SDC) detector (Figure 23). The SDC detector does not require cooling by liquid
nitrogen, and its count rate capability is extremely high at ~400,000 to 1,000,000 cps. Thus,
one SDC detector at a count rate of 400,000 is equivalent to ~8 conventional EDS detectors with digital pulse processor (Figure 24). Initially, the SDC detectors had a beryllium
window and could not detect elements lighter than ~Na. In addition, the energy resolution
was not as good as that obtained by conventional EDS detectors. Thus, the electron
microscopy community was not very enthusiastic in adopting SDC detectors. Presently,
SDC detectors are available with a polymer window and can detect elements lighter than Na. In
addition, by compromising on the count rate, SDC detectors can achieve better energy resolution than conventional EDS detectors. It is also possible to obtain counts from a combination of four SDC detectors to acquire even higher count rates.
The present configuration of the image analyzer at CANMET-MMSL is still based on
an electron microprobe with two WDS detectors and one EDS detector, but it has an
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450
p+
400
Last Ring
n– S i
Back
Count Rate Capabilities, kcps
Clear
Anode
Field Effect
Transistor Ring #1
350
300
250
200
150
100
50
0
Analog
FIGURE 23
detector
General scheme of the SDC
Digital
SDC
FIGURE 24 Differences in count rate
capabilities of different EDS detectors at
similar energy resolutions (analog =
conventional EDS detector with analog
pulse processor; digital = conventional EDS
detector with digital pulse processor; SDC =
silicon drift chamber detector)
additional SDC detector (Röntec). The SDC detector allows acquisition times of 10 msec
per pixel and shorter times if electron beam currents of at least 15 nA are used. In full BSE
mode, the system at CANMET-MMSL processes ~100,000 particles per hour. In full X-ray
mode, it processes ~20,000 particles per hour; it is presently faster than any commercially
available systems. Speed is important, especially when considering the analysis of tailing
samples that contain very few grains of the mineral of interest, a common situation when
investigating the losses of valuable minerals.
Electron microprobes are instruments that are designed to provide very stable electron
beam currents. A stable electron beam current is a basic requirement for EMP analysis. The
electron beam current in an EMP is stable over a period of weeks and can be automatically
monitored and maintained constant every second. On the other hand, the SEM is basically
an imaging instrument and typically provides less-stable electron beam currents. In many
SEM instruments, the electron beam current varies over periods of 1 hour (Gu 2003). To
fully exploit the speed of quantitative mineralogy based on BSE imaging, the instrument
must provide a stable electron beam current over the full time of sample analysis. To analyze
a set of samples or to search and analyze precious minerals, the electron beam current must
be stable during a period of up to 24 hours. In addition, the EMP can yield higher beam currents than the SEM. Higher beam currents yield better contrast between minerals in the
BSE image. Also, EMP instruments commonly have four large ports and two additional
auxiliary ports. These six ports can be used to connect a range of X-ray detectors. Thus,
EMP instruments are better suited for quantitative mineralogy studies.
Without budget constrictions, the future instruments for quantitative mineralogy studies will be EMPs with two WDS detectors, each with four analyzing crystals and four SDC
detectors. The cost of these instruments could be similar to the cost of the present commercially
available systems for quantitative mineralogy studies. Of course, there will be some simpler
instruments for basic requirements and limited budgets.
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X-RAY ABSORPTION SPECTROSCOPY
Practically, X-ray absorption spectroscopy (XAS) must be carried out on a synchrotron radiation source to provide the wide X-ray wavelength range and high intensity required. XAS is
primarily of use in mineral flotation studies where information about the form of trace elements is required; that is, the form of the element that is of interest is not obscured by the
signal from the same element within the bulk sample. However, in certain modes, XAS can
also be used to probe surface phenomena where the element of interest is also present in the
underlying specimen. Surprisingly, XAS has been used relatively rarely for mineral flotation
studies, although there are a number of advantages to its use as compared to techniques that
require a vacuum environment. There is generally little sample preparation required, and the
sample turnaround time is rapid as most measurements can be carried out in air. XAS measurements lend themselves to the use of environmental cells and provide both spatial and
electronic structure information.
XAS can be divided into several subtechniques, which are more commonly referred to
in the literature (for a general review of XAS and other synchrotron techniques, see Gerson,
Halfpenny et al. 1999). In general, XAS is divided into two categories: XANES (X-ray
absorption near-edge structure), and EXAFS (extended X-ray absorption fine structure).
The former occurs up to approximately 40 eV above the absorption edge and is the result of
excitation from the valence to conduction bands. XANES is sensitive to the coordination
geometry and oxidation state. EXAFS results from scattering of excited photoelectrons off
neighboring atoms and occurs at higher incident energies than XANES. EXAFS is sensitive
to local structure out to approximately 5 Å.
XAS can be measured in several modes. Most commonly, the X-ray fluorescence (measured at 90º to the incident X-ray beam) or transmitted intensities are measured. Transmission measurements enable the bulk sample to be probed, whereas fluorescence
measurements are more surface sensitive with a measurement depth dependent on the
energy of the fluorescence yield (see, e.g., Kasrai et al. 1996). However, both measurement
modes can be used effectively to study surface-related phenomena where the element under
investigation is surface specific.
More surface-sensitive measurements can be obtained using total electron yield (TEY),
partial electron yield (PEY) or reflection EXAFS (REFLEXAFS). TEY is measured via the
drain current experienced on electron excitation to the continuum. The depth of analysis of
this form of XAS is dependent on the incident X-ray energy. Thus, for the Si L-edge (95–120 eV)
the depth of measurement is about 50 Å, whereas for the Si K-edge this would be approximately 700 Å (1,830–1,900 eV) (Kasrai et al. 1996). PEY is obtained via the measurement
of electron current induced on a conducting grid a short distance from the sample surface. A
voltage can be applied to the grid to repel low-energy electrons (i.e., those resulting from
deeper within the sample). REFLEXAFS requires the incident beam to intercept the surface
at a very low angle, thus enabling reflection rather than absorption. An angle of 100 millidegrees, half of the critical angle for chalcopyrite, has been adopted (England et al. 1999b). In
this instance, the depth of penetration was less than 50 Å. The data can be collected via measurement of the intensity of the reflected beam or the fluorescence yield at 90° to the sample
surface. This enables surface-sensitive measurements to be obtained at monolayer or even
sub-monolayer coverage (Greaves 1991). However, REFLEXAFS requires a flat, polished
specimen, whereas the other modes of measurement can be carried out on powders, slurries,
or liquids.
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XAS has been used for the study of Zn and Pb sorption onto chalcopyrite (CuFeS2)
(England et al. 1999a, b and references therein). Chalcopyrite tends to have natural hydrophobicity and thus self-floatability. This characteristic is attributable to the formation of a
sulfur-rich surface layer due to metal loss under conditions of low pH. The adsorption of
metal cations, as hydroxides, can result in a reduction in the flotation response. REFLEXAFS data revealed that Zn adsorbed at pH 5.5 is bonded to O (at 2.00 Å) only. There was no
evidence of bonding to S. On addition of xanthate (at pH 10.2), the Zn coordination
changed radically so that the nearest neighbor coordination sphere contained both O and S
(at 1.97 Å and 2.33 Å, respectively). For Pb adsorption, both O and S coordination is
present prior to the addition of xanthate. After xanthate addition, only S coordination is
observed. It appears, therefore, that the initial adsorption site of the Zn species is onto the
oxidized Fe(OH)x species present on the surface, whereas the Pb is adsorbed onto the sulfurrich surface regions. This interpretation provides an explanation for the observation that
only small amounts of Pb in solution are required to remove chalcopyrite self-flotation.
Adsorption of Pb (probably as a hydroxide) would render the hydrophobic sulfur-rich
regions hydrophilic. The loss of self-floatability would be much less affected by adsorption
of Zn species onto the already hydrophilic, oxidized Fe surface regions.
Another study, originating from the same research team, examined the adsorption of
Cu and Pb onto sphalerite (ZnS) (England et al. 1999a; Pattrick et al. 1998, 1999). Cu
adsorption activates the sphalerite surface for enhanced collector adsorption. ZnSe was also
examined, as an isostructural analogue to sphalerite so that the sulfur XAS data could be
obtained from the adsorbed xanthate without the spectra being swamped by bulk S contributions. On the basis of analysis of the REFLEXAFS data, it was proposed that 3 S atoms
and 1 O atom are bonded to the adsorbed Cu atoms (at 2.25 Å and 2.07 Å, respectively).
On addition of xanthate, the O were replaced by an S atom. The fourth S atom has a considerably longer bond distance to the Cu atom as compared to the three initial S atoms, 2.75 Å
as compared to 2.22 Å. Pb adsorbed onto sphalerite was proposed to be coordinated solely
to O atoms. On addition of xanthate, the Pb coordination consisted of two bonds to S and
two to O. It would appear, therefore, that, at least under the conditions used for this study,
Pb does not specifically adsorb onto sphalerite as does Cu.
A similar study of the adsorption of Cu onto sphalerite has also been carried out by
Gerson, Lange et al. (1999). Analysis of the EXAFS data indicated Cu coordination to S
(2.26–2.30 Å) but no evidence of Cu coordination to O. A data set plus Fourier transform
for Cu adsorbed onto sphalerite sample are shown in Figures 25a and 25b. The Fourier
transform is similar to a one-dimensional electron density distribution surrounding but offset by a phase shift. The large central amplitude is the result of the surrounding S atoms. The
other two maxima centered at approximately 1 Å and 2.7 Å are the result of the truncation
of the Fourier transform series from the experimental data. The exact location of the maxima will depend on the data range that is fitted by the simulated model data. In order to fit,
the analytical program must truncate the model data in a manner similar to the experimental data. It is unclear in the studies by Pattrick et al. (1999) whether the interpretation of O
coordination is, in fact, the result of this artefact induced by Fourier series truncation.
On the basis of the interpretation of the EXAFS and XANES data, together with other
experimental evidence and previous knowledge, a mechanism for the Cu activation of
sphalerite was presented by Gerson, Lange et al. (1999). This mechanism proposes the
replacement of Zn by Cu on the sphalerite surface to form a distorted trigonal planar structure, similar to the Cu structure found for half the Cu atoms within covellite (CuS). In-situ
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A
B
0.01
3.5
–0.01
4.5
5.5
6.5
–0.03
k, Å –1
7.5
8.5
Transform Amplitude,
arbitrary units
k3Xobs(k),
arbitrary units
0.03
3.0
2.5
2.0
1.5
1.0
0.5
0.0
0.5
1.0
1.5
2.0
2.5
3.5
r, Å
C
NOTE: (a) EXAFS data collected for Cu adsorbed
–4
onto sphalerite conditioned in 10 M Cu(NO3)2
for 15 minutes at pH 5.5. The regular sinusoidal
nature of the data is indicative of a single
coordination shell. (b) The Fourier transform of
the model and experimental data sets. Although
three maxima are shown, only one atomic shell
has been used to fit this data (Cu–S 2.26 Å).
(c) The proposed structure of Cu adsorbed onto a
sphalerite (110) surface. For both (a) and (b),
(x) shows the experimental data, and (–) shows
the data simulated from the model.
FIGURE 25
Experimental EXAFS data sets with correspondence to proposed copper structure
reduction of Cu(II) to Cu(I) occurs, resulting in the oxidation of the surrounding S atoms.
This smearing out of the surface charge results in the increased degree of hydrophobicity for
Cu-activated sphalerite as compared to unactivated sphalerite that has been observed and,
hence, results in increased floatability even without collector addition. The conditions
under which Cu activations were carried out during this study were carefully chosen to
inhibit copper(II) hydroxide (Cu(OH)2) formation.
The Cu activation of pyrite (FeS2) has also been investigated by Weisener and Gerson
(2000). On the basis of EXAFS data, a similar structure for adsorbed Cu to that projected
for sphalerite was proposed (i.e., distorted trigonal planar) and near-identical Cu–S bond
lengths were determined. Again, there was no evidence of Cu–O coordination except in the
instance where Cu(OH)2 was purposefully precipitated within the activating solution (pH
8.5, 2.84 × 10–4 mol m–2 Cu). In this instance, Cu–S bond lengths of 2.29 Å and Cu–O
bond lengths of 2.00 Å were derived from the EXAFS data. This also provides a possible
explanation for the observations by Pattrick et al. (1999) as to the coordination of adsorbed
Cu to O.
Todd and Sherman (2003) used XAS to probe the surface oxidation mechanism of
chalcocite (Cu2S). This leads to reduced flotation of chalcocite in minerals processing circuits. In this instance, TEY collection mode was used. For the Cu L-edge measurements
undertaken, there is likely to be a contribution from the bulk mineral. However, the O K-edge
measurements will be surface specific only because of the absence of O in the bulk. Oxidation of chalcocite in different pH solutions indicated that under acidic pH, cuprous oxide
(Cu2O) dominates the oxidation products, whereas under alkaline pH, cupric oxide (CuO)
is the major oxidation product.
XANES has been employed in a study of selected elements (Ti, V, Cr, Mn, and As) in
deep-cleaned Kentucky No. 9 coal (Huggins et al. 1997). Tail and float samples were prepared
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321
using both a Denver cell and column flotation. The tails and float fractions were examined
in order to gain a better understanding of the distribution and form of elements that may be
released during combustion and therefore form a pollutant risk and, hence, to identify
appropriate coal-cleaning strategies. XANES data has been interpreted through comparison
with data obtained from standards. V, Ti, Cr, and Mn were all found to be different in form
in the tail and float fractions. The tailings fraction was interpreted as containing Ti and V in
interlayer positions within illite, whereas the Ti and V present in the float fraction was proposed to be in an organic association. The pre-edge features from the V spectra indicate the
oxidation state of V to be higher in the float fraction than in the tails fractions. The difference between the data collected for Cr in the float and tailings fraction was found to be subtler with Cr present in the tails as illite and in float fractions as a hydroxide species. Mn in
the tails was associated with calcite (CaCO3) and illite, whereas in the float fraction, two
different organic forms were identified. The XANES data for the float fractions for these
elements suggest the principal coordination is to O anions. In contrast, As was primarily
associated with pyrite and oxidation products thereof, and the relative proportions of these
forms was not float or tail dependent.
Additional opportunities exist to further utilize XAS measurements to advance the
understanding of mineral flotation studies. In particular, an opportunity exists to exploit
energy-dispersive XAS for mineral processing applications. This type of measurement can
be carried out rapidly as it uses a broadband incident X-ray beam rather than the monochromatic beam traditionally employed. The latter requires a monochromator sweep in order to
scan the energy range required for the spectra. Energy-dispersive XAS can be carried out on
in-situ materials and in real time. Data acquisition can be carried out in the order of 10 seconds.
SCANNED PROBE MICROSCOPIES
Though the scanned probe microscopies—scanning tunneling microscopy (STM), and
atomic force microscopy (AFM)—are essentially research tools applied to model systems,
they have added much in the last two decades to the understanding of surface reactions and
adsorption mechanisms of minerals under flotation-related conditions. Both techniques can
image surfaces close to the atomic level. STM requires reasonably conductive samples
(which has made galena a mineral of choice) with some chemical identification in scanning
tunneling spectroscopic mode. AFM can image insulating surfaces, cannot chemically identify atoms, but can give astonishing information on particle–particle and even particle–bubble
interactions and changes with reaction and adsorption in force–distance (approach–retract)
mode. Hochella (1995) has reviewed STM and AFM studies of mineral surfaces and their
oxidation. Smart, Amarantidis et al. (2003) have reviewed applications to oxidation and collector additions in flotation. Some examples of particular insights, related to surface processes described previously, will illustrate the unique types of information that these
techniques can provide.
The development of isolated, patchwise oxidation in air and solution has been very well
illustrated by STM studies of galena surfaces. Eggleston and Hochella (1990, 1991, 1992)
have imaged (001) surfaces of galena at atomic scale after exposure to water for 1 minute.
Apparent vacancies at the sulfur sites are correlated with oxidation in their model of this
process. The oxidized regions do not initiate randomly, but after oxidation has begun at a
site, these regions tend to nucleate and grow without initiation of new sites. The boundaries
of the oxidized regions tend to lie along the [110] directions, apparently due to S atoms
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across this front having only one nearest-neighbor-oxidized sulfur whereas an unoxidized S
across a [100] boundary would have two nearest-neighbor-oxidized sulfurs. As with crystallization processes, [100] fronts move fast and disappear, leaving the slow-moving [110]
dominant.
At lower magnification, the process of galena oxidation in air has also demonstrated
random sites of oxidation and growth on (001) galena surfaces with no clear preference for
initiation at step edges or corners (Laajalehto et al. 1993). This process is illustrated in Figure 26
from that work and correlated with XPS spectra, showing that the initial oxidation products
are peroxide, hydroxide, and carbonate species successively. With time up to 270 minutes in
air, the oxidation products grow from surface features with lateral dimensions <0.6 nm to
overlapped regions >9 nm diameter with “holes” in the overlayer that still allow access to the
underlying sulfide surface. Further studies of galena oxidation in air (Kim et al. 1994), comparing synthetic and natural galena samples, confirmed that the growth mechanism on natural galena with the oxidation initiation sites correlated with impurity atoms in the surface
layer. The very much slower oxidation of synthetic galena did occur preferentially on edges,
dislocations, and lattice defect sites on the (001) faces of the galena crystal. The XPS spectra
in this case show predominantly lead hydroxide and sulfate with a smaller contribution from
carbonate in the oxidation products.
In solution, STM (and AFM) imaging showed the development of subnanometer pits
with increasing reaction time in air-purged water at pH 7 (Kim et al. 1995). The boundaries
of the pits lie in the (100) and (010) directions in the galena surface with depths corresponding to unit cell dimensions of galena (i.e., 0.3 and 0.6 nm). The process occurring in
solution is congruent dissolution, confirmed by XPS spectra showing unaltered Pb4f and
S 2p signals. The x- and y-dimensions of the pits and their rates of formation depend
strongly on the pH and purging gas (i.e., O2, air, N2) used. Dissolution rates, determined
directly from STM images of monolayers removed, decrease with increasing pH in agreement
with the reported dissolution studies on galena (Fornasiero, Ralston, and Smart 1994;
A
B
C
FIGURE 26 STM images from a 70 × 70 nm area of (a) freshly cleaned galena surface; (b) the
same surface after 70 minutes standing in air; (c) after 270 minutes in air. The upper row are
grayscale images; the lower row are 3-D (rotated) images with the vertical scales 1.8, 2.0, and
3.8 nm, respectively. Constant current mode (0.2–0.25 nA), bias ~0.35 V.
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323
Hsieh and Huang 1989). For all pH values studied, the growth of the dissolution pits is significantly greater in the x- and y-directions than in the z-direction, suggesting that the edges
are more active toward dissolution than the faces (Kim et al. 1995). At the relatively low surface-area-to volume ratios used in the STM studies, there is no evidence for the growth of
surface oxidation products similar to those observed in air or to adsorption/precipitation of
lead hydroxide colloids from solution. Increasing the lead ion concentration to 10–3 M in
solution resulted in surface product formation as elongated, oval colloidal projections with
dimensions of ~50 nm × 20 nm and average heights of 14 nm, with a distinct [110] directionality. XPS analysis has confirmed that these species are predominantly lead hydroxide,
presumably formed from the hydrolysis of lead ions followed by surface attachment. It is not
yet clear whether the mechanism of surface attachment involves the formation of lead
hydroxide colloids in solution and their precipitation onto the galena surface or adsorption
of Pb2+/Pb(OH)2 molecular species at specific sites on the galena surface before in-situ
growth. However, the formation of these patchy surface layers shows that the galena surface
is heterogeneous and that its overall hydrophobicity and flotation response will be controlled not only by the surface chemistry but by the surface arrangement of hydrophilic and
hydrophobic patches. Atomic level imaging of the (001) surface of galena, reviewed by
Hochella (1995), has been achieved, including observation of the oxidation of a single S site
at which the tunneling current has been effectively extinguished.
In-situ STM images of a freshly cleaved galena crystal in contact with an air-equilibrated 10–4 M ethyl xanthate solution show colloidal particles of lead ethyl xanthate (as confirmed by XPS and FTIR) formed at the surface corresponding to multilayer surface
coverage. In-situ STM studies of ethyl xanthate treated preoxidized galena surfaces have also
shown the removal of oxidized lead species and the formation of colloidal lead ethyl xanthate particles as flattened spheres with diameters of 10–20 nm and average heights of 6 nm
(Ralston 1994b; Kim et al. 1995).
Combined XPS and AFM studies of galena oxidation in acetate buffer (pH 4.9) by
Wittstock et al. (1996) produced dramatic imaging of elemental sulfur protrusions 10–
200 nm after initial roughening of the galena surface. These protrusions are separated by
several hundred nanometers and appear to result from a process of diffusion in the aqueous
phase. XPS shows the formation of elemental sulfur starting at potentials more anodic than
160 mV SHE (saturated hydrogen electrode). AFM imaging first detects the protrusions at
+236 mV SHE. The authors, therefore, propose that the process causing surface roughening
is dissolution of PbS to Pb(II) ions and HS– ions, whereas the deposition reaction is the
electrochemical oxidation of HS– ions to elemental sulfur. It seems likely that sulfur formation starts at impurity locations leading to different rates and sizes of protrusion development.
In nickel sulfide processing, magnesium silicate (MgO) gangue minerals often report to
the concentrate, causing downstream processing problems as well as increased smelting
costs. In addition, these hydrophilic MgO minerals may interfere with the flotation of valuable sulfide minerals such as pentlandite [(Fe, Ni)9S8]. Flotation of the MgO particles may
be via composite particles or through attachment to the valuable minerals as slime coatings.
A coating of hydrophilic slime particles will decrease the hydrophobicity of the sulfide particle and may also reduce collector adsorption (Learmont and Iwasaki 1984). Either of these
flotation mechanisms will reduce both the flotation rate and recovery, and will, therefore,
result in lower recoveries of the valuable sulfide minerals (Trahar 1981; Senior and Trahar
1991; Wellham, Elber, and Yan 1992). Slime coatings of lizardite and chrysotile have been
found to adhere to the surface of pentlandite, reducing its flotation rate (Edwards, Kipkie,
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324
FLOTATION FUNDAMENTALS
0.04
Force–Distance, mN/m
0.02
0.0
0
20
40
60
70
100
–0.02
–0.04
–0.06
Apparent Separation, nm
FIGURE 27 Direct interaction forces as a function of apparent separation measured on
approach between a pentlandite surface ( ) and lizardite particle ( ) in 10–3 M KNO3 (potassium
nitrate) at pH 9.4 and in the presence of 20 ppm CMC
and Agar 1980; Li 1993; Chen at al. 1999a, b; McQuie 1999). The interaction between lizardite and pentlandite has been directly investigated using the atomic force microscope in
force–distance mode and electrokinetic zeta potential determinations as a function of pH
and in the presence of the polymeric dispersant, carboxymethyl cellulose (CMC) (Bremmell, Fornasiero, and Ralston 2005). The lizardite mineral was positively charged with the
zeta potential independent of pH. The magnitude and sign of the pentlandite particles were
pH dependent and were negative for pH values above 4.5. At pH values greater than 9,
where flotation of nickel sulfide ores is routinely performed, the two minerals are oppositely
charged and, therefore, attract through an electrostatic mechanism. Direct interaction force
measurements between pentlandite and lizardite surfaces as a function of pH demonstrate
this attractive interaction. Adsorption of CMC at the lizardite–solution interface overcompensates the positive charge on the lizardite particle, and its zeta potential is rendered negative. In the presence of CMC, a repulsive interaction force between lizardite and
pentlandite, which was concluded to be of electrosteric origin, was measured in the AFM
(Figure 27). The results explain the flotation behavior of the minerals performed in this and
previous studies.
P L A N T C A S E S T U DY : U S I N G S U R FA C E A N A LY S I S T O
EXAMINE ZINC CIRCUIT MINERAL LOSSES
Introduction
The Matagami mine sulfide flotation plant in Northern Quebec is the focus operation for
this study. The Matagami plant produces copper and zinc concentrates. This study investigated zinc losses in the zinc circuit using surface analysis. For this work, the rougher feed,
rougher concentrate, rougher tails, cleaner tails, and zinc concentrate streams were sampled
and examined in the CANMET-MMSL surface science laboratories located in Ottawa,
Canada. Pulp samples were collected using established and modified protocols designed to
preserve the surface chemistry on particles from the time of sampling until the time of analysis. The surface chemistry of particles in each pulp was examined using the surface analysis
methods of X-ray photoelectron spectroscopy (XPS) and Auger electron spectroscopy
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SURFACE CHARACTERIZATION AND NEW TOOLS FOR RESEARCH
325
(AES). The bulk mineralogy of each pulp sample was determined from X-ray diffraction
(XRD) powder patterns. XRD was used because the materials are crystalline, and the
method has a detection range that is readily comparable to that of XPS. Standard XPS provides excellent chemical state information, but the spatial resolution of the data is limited to
hundreds of micrometers. For the acquisition of surface chemical information from individual particles, AES was used. The spatial resolution of the AES surface chemical data collected using a scanning Auger microprobe is limited essentially by the diameter of the
primary electron beam; therefore, AES spectra can be collected from exceedingly small particles.
Methods
To prevent the exposure of individual particle surfaces to air and preserve the chemistry of
the solid–liquid interface present at the time of collection, the sampling steps outlined in
the next section were followed. These procedures are in a continual state of development
and are the product of discussion, consultation, and experimentation with surface science
groups in academia, government, and industry in Africa, Australia, Europe, and North
America. The protocols are based on the premise that each particle in a pulp will have a
hydrated boundary layer that will prevent direct contact between the particle surface and
the atmosphere. When frozen, it is a passivating and protecting layer of ice that isolates individual particles from the air (Smart 1991; Love, Cayless, and Hazell 1993; Pratt, Nesbitt,
and Muir 1994).
Pulp samples were prepared for XRD analysis by grinding 2 g of material to less than
10-μm particle diameters using a custom micromill. XRD analyses were conducted on the
samples using an automated Rigaku diffractometer equipped with a rotating copper anode
X-ray source. XRD powder patterns were collected using monochromatic radiation; they
were then processed, and the mineral phases were identified using the Materials Data Inc.
powder diffraction pattern analysis program JADE (Release 6.1) and the ICDD Powder
Diffraction Database (Release 2001).
XPS spectra were collected using a PerkinElmer Corporation (Physical Electronics
Division) PHI-5600 spectrometer equipped with an OMNI V lens system. XPS data were
collected using 400-W achromatic Mg X-rays (Ex-ray = 1,353.6 eV) and Al X-rays (Ex-ray =
1,486.6 eV). The two X-ray sources were utilized as a means of resolving contributions from
coincident photoelectron and X-ray-induced Auger electron emissions. The energy scale of
the spectrometer was calibrated to the metallic Au(4f7/2) line at 84.0 eV and was to give an
energy dispersion of 857.8 eV between the metallic copper 2p3/2 and 3p lines. The analyzer
pass energy was 187.0 eV for broad-energy-range “survey” scans and 29.35 eV for narrowenergy-range “multiplex” scans. The binding energy scale of the spectra is referenced to the
C1s peak from adventitious hydrocarbon (static charge referencing) fixed at 285.0 eV (Swift
1982). XPS information was collected from spots measuring nearly 400 μm in diameter,
and the vacuum in the analytical chamber was approximately 8.0 × 10–10 Torr during analysis.
Details on the analysis of XPS spectra are provided in work by Pratt, Nesbitt, and Muir
(1994).
AES spectra were collected using a PerkinElmer Corporation (Physical Electronics
Division) PHI-600 scanning Auger microprobe equipped with a LaB6 thermionic electron
emitter. Analyses were obtained using an electron beam accelerated to a potential of 3.0 kV
at a current of 90 nA and an analyzer energy resolution set to 0.6%. Under these instrument
conditions, exceedingly small particles could be examined without any interferences from
the sample mount matrix. This is because the volume of analysis is defined by the primary
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FLOTATION FUNDAMENTALS
beam diameter (~300 nm) and the escape depth (λ) of Auger electrons, which is approximately 1 to 3 nm beneath a particle surface. Lambda (λ) is dependent on the kinetic energy of
the Auger electron and the solid’s density. Semiquantitative surface compositions were calculated using peak-to-peak heights and manufacturer-supplied, empirically-derived sensitivity
factors. The vacuum in the analytical chamber was about 6.0 × 10–10 Torr during analysis.
Metallurgical Processing Surface Science Protocols for Sampling and Analysis
The experimental protocol for preparation of surface samples is as follows:
1. Sample the pulp stream using a beaker.
2. Place several milliliters of pulp into a vial.
3. Purge with N2(g).
4. Cap the vial and seal the cap with silicon.
5. Freeze the sample as quickly as possible, and maintain in a frozen state until time for
analysis.
6. Thaw the sample within 30 minutes of the scheduled analysis.
7. Using a pipette, remove 1.5 mL of the process water and deposit it into a micro test tube.
8. Using a micro spatula, remove a minute amount of pulp and place it into the micro
test tube. For XPS-destined samples, the solution should be slightly clouded; this
amount provides about a monolayer of particle coverage on the filter. For AESdestined samples, the solution should remain clear; this amount provides the dispersed particle coverage needed for Auger analysis.
9. Separate the solids from the solution onto a 2.5-cm-diameter 0.45-μm nitrocellulose membrane filter using a vacuum filtration system. (NOTE : For these experiments, the membrane filter was sputter-coated with gold to increase conductivity
and fixed to the appropriate sample platen.)
10. Prepare the damp samples for surface analysis using an ultrahigh vacuum (UHV)
conditioning chamber that is attached to the XPS instrument. (NOTE : The conditioning chamber used in these experiments was designed and built by one of the
authors. One function of the conditioning chamber is to prepare damp samples for
the UHV conditions required for spectroscopic surface analysis. The chamber
design is attached to the XPS via a series of UHV gate valves and stainless-steel conduits to the analytical chamber of the XPS. This permits all manipulations to be
undertaken at high to ultrahigh vacuum conditions. Conditioning takes up to 15 hours;
therefore, UHV conditions should be maintained throughout the experiment.)
11. For collection of AES data, transfer the conditioned samples to the scanning Auger
microscope instrument from the XPS using a high-vacuum transfer vessel. (NOTE :
The high-vacuum transfer vessel used in these experiments was designed and built at
CANMET by Dr. Jim Brown.)
Results and Discussion
The XRD powder pattern and XPS survey scan data collected from the selected zinc circuit
feeds and the concentrates show that there is a progressive lessening in complexity with the
evolution of the pulps within the circuit. This discussion focuses on the results obtained
from the sphalerite component in the pulps. Contributions from sphalerite can be identified
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SURFACE CHARACTERIZATION AND NEW TOOLS FOR RESEARCH
327
in XRD patterns collected from each of the pulps examined. XPS survey scans collected
from each of the pulps show contributions that can be attributed to sphalerite as well. They
also show distinct Cu contributions. Qualitative evaluation of the XRD data shows that
only minor to trace amounts of chalcopyrite (CuFeS2) can be found in the pulps. The source
for the Cu detected by XPS is most likely the Cu added for activation of sphalerite. Because
the vast majority of the particles in the Zn concentrate is sphalerite, the approximate surface
Zn/Cu ratio of 3:1, obtained from this pulp, can be used as a model ratio for the qualitative
evaluation of sphalerite activation in each of the pulps examined. The approximate Zn/Cu
ratios for the pulps examined are as follows: rougher feed >30:1, rougher concentrate 4:1,
rougher tails 10:1, and cleaner scavenger tails 6:1.
As inadvertent activation of the sphalerite in the feed should not be very high, a high
Zn/Cu ratio is expected. Activation appears to have occurred within the rougher, and the
rougher concentrate has the appropriate Cu activation ratio. For the rougher tails, the high
Zn/Cu ratio obtained may indicate that the small amount of sphalerite present in the pulp
has not been sufficiently activated. For the cleaner scavenger tails, the Zn/Cu ratio is about
50% greater than the model ratio. These results indicate that there is a problem with the
activation of the sphalerite in the circuit.
Information into Cu chemistry at particle surfaces can be obtained through evaluation
of Cu 2p spectra collected from the rougher concentrate, rougher tails, cleaner scavenger
tails, and Zn concentrate (Figure 28). The spectra have been normalized such that the intensities near 932 eV are nearly coincident. The overlain spectra are characterized by Cu 2p3/2
and 2p1/2 peaks, respectively, at 931.8 eV and 951.7 eV. The mineralogy of the Zn concentrate is mainly sphalerite, and the Cu 2p collected from this pulp is interpreted to originate
from activated sphalerite surfaces. The shapes and positions of the Cu 2p peaks are similar
to those reported for Cu(I) (Chawala, Sankarraman, and Payer 1992), and, in agreement,
the signals are interpreted to originate from Cu(I) ions. The pulp with activated surfaces
most closely resembling those of the Zn concentrate data is the rougher concentrate data.
The Cu 2p spectra collected from the rougher tails and cleaner scavenger tails have an additional contribution near 942 eV. The position and shape resemble those reported for Cu(II)
ions (Chawala, Sankarraman, and Payer 1992) and are interpreted to be from Cu(II) species. These results show that a portion of the Cu on the sphalerite surfaces in the tails is
found as Cu(II).
AES Analyses of Individual Sphalerite Particles
AES spectra were collected from individual sphalerite particles in the process stream pulps.
A minimum of 10 particles in each pulp was analyzed. The AES spectra showed contributions from S, C, O, and the transition metals Fe, Cu, and Zn. On many of the sphalerite particles examined, contributions from Ca were detected. Two spectra representative of the
AES data collected in the study are shown in Figure 29.
The Ca detected is interpreted to be associated with a precipitated Ca sulfate species,
possibly gypsum. Although the sampling size is small, the AES results show that Ca concentrations are consistently the lowest on sphalerite particles in concentrates. Conversely, Ca
concentrations are consistently higher on sphalerite particles in tails.
Using O concentrations as a guide to the degree of surface oxidation, the AES data
shows that the tails have sphalerite particles that are the more oxidized (Figure 29a), and the
two concentrate pulps have sphalerite particles that are the least oxidized (Figure 29b).
These trends appear to apply to both the coarse and fine particles examined within the pulps.
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FLOTATION FUNDAMENTALS
Cu 2p3/2
Cu 2p1/2
Zn Rougher Concentrate
Zn Rougher Tails
Zn Cleaner Scavenger Tails
Zn Concentrate
1 Cu(I)
Normalized Intensity
2 Cu(II)
2
1
965
960
955
1
950
945
940
935
930
925
Binding Energy, eV
FIGURE 28 Overlain narrow region Cu 2p spectra collected from four zinc circuit pulps. The
peak intensities shown have been normalized. The dotted traces are the concentrates and the
solid traces are the tails.
dN(E)
A
Cu 0.3%
Ca 3.4%
Si
3.0%
Fe 9.2%
Zn 7.3%
O 10.5%
C 46.7%
S 17.7%
dN(E)
B
O 1.3%
C 18.5%
Fe 7.9%
Cu 9.1%
Zn 25.4%
S 37.8%
40
180
320
460
600
740
880 1,020 1,160 1,300 1,440
Kinetic Energy, eV
FIGURE 29 Representative AES scans collected from sphalerite particles in the (a) cleaner
scavenger tails and (b) zinc concentrate. Values are atomic %.
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SURFACE CHARACTERIZATION AND NEW TOOLS FOR RESEARCH
329
XRD shows that an appreciable amount of sphalerite is found in tailings pulps within
the circuit, and the XPS results show that the sphalerite is found in the tails because of inadequate Cu activation. Examination of the Auger spectra collected from individual sphalerite
particles in the circuit pulps corroborates the XPS interpretation. The sphalerite particles
examined in the two concentrate pulps show well-defined Cu peaks (Figure 29b). Those in
the tails have no clear and unambiguous contributions from Cu (Figure 29a). The number
of particles examined is small and a larger sample population would greatly increase the confidence in the interpretations put forward for the Auger data.
S U M M A RY
AES examination of sphalerite in the two concentrates shows that these particles are less oxidized and cleaner (less calcium species) than those in the two tails. The XPS and AES
results show that sphalerite component in the tails has not been sufficiently activated, and
the XPS results show oxidized copper is found on the surface of the particles.
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———. 1999. A SERS spectroelectrochemical investigation of the interaction of O-isopropyl-Nethylthionocarbamate with copper surfaces. Colloids Surf. A 146:63–74.
Woods, R., G.A. Hope, and G.M. Brown. 1998a. Spectroelectrochemical investigations of the
interaction of ethyl xanthate with copper, silver and gold: II. SERS of xanthate adsorbed on silver
and copper surfaces. Colloids Surf. A 137:329–337.
———. 1998b. Spectroelectrochemical investigations of the interaction of ethyl xanthate with
copper, silver and gold: III. SERS of xanthate adsorbed on gold surfaces. Colloids Surf. A
137:339–344.
Woods, R., G.A. Hope, and K. Watling. 2000. A SERS spectroelectrochemical investigation of the
interaction of 2 mercaptobenzothiazole with copper, silver and gold surfaces. J. Appl.
Electrochem. 30:1209–1222.
Young, C.A., and J.D. Miller. 2000. Effect of temperature on oleate adsorption at a calcite surface: An
FT-NIR/IRS study and review. Int. J. Miner. Process. 58:331.
Zachwieja, J.B., J.J. McCarron, G.W. Walker, and A.N. Buckley. 1989. Correlation between the
surface composition and collectorless flotation of chalcopyrite. J. Colloid Interface Sci.
132(2):462–468.
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The Flotation of Fine and Coarse
Particles
Graeme J. Jameson, Anh V. Nguyen, and Seher Ata
INTRODUCTION
The flotation process is used to separate or recover particles over a very wide range of sizes.
In the minerals industry, it is not unusual to grind an ore to a top size of 6–7 μm in order to
liberate the valuable material, and the feed to flotation will presumably include a certain
proportion of material that is below 1 μm. At the other end of the scale, the top size is limited by the performance of the flotation circuit; above a certain point, the recovery drops off
as particle size increases. Thus, in mineral flotation, there is a range of particle size that is
typically between 10 and 70 μm for base metals, where high recoveries can be obtained with
acceptable cell residence times. Outside this range, however, the recoveries decrease for various
reasons.
This review will focus on the two regions where high recoveries are more difficult to
achieve: the ultrafine (<10 μm) and the coarse (>70 μm). The physical factors that affect
each case and tend to reduce the recovery rate will be discussed. Special measures for increasing the recovery rates will be described. Emphasis will be placed on the physical effects that
arise in the extremes of particle size, which helps to explain the difficulties encountered and
paves the way for future action.
In order to provide suitable bounds for this chapter, a number of relevant topics will be
excluded, despite their obvious importance. It will be assumed that the particles to be recovered by flotation have been properly conditioned with appropriate reagents, so flotation
chemistry will not be considered a factor in the flotation of fine and coarse particles, except
where the hydrophobicity as reflected in the contact angle is specifically involved. Similarly,
the general principles of particle capture, hydrodynamics of flotation cells, the bulk flow of
froths, and the phenomena of liquid drainage within froths will not be discussed unless
there is some relevance to the particular problems of the extremes of particle size in flotation. There are some otherwise interesting papers in which the Hallimond tube was used to
study the effects of particle size on flotation, but these papers will not be discussed because
of the difficulty in characterizing the hydrodynamics of this apparatus. In the same vein,
although there have been numerous papers describing operational procedures in plants to
improve the recovery of adventitious coarse particles, which are present only in small concentrations, such discussions will not be included.
C A P T U R E O F PA R T I C L E S
The classical diagram showing the effect of particle size on recovery was given by Jowett
(1980) using data from Trahar (Figure 1). When lead or zinc sulfides were floated for a fixed
time in a batch cell, the recovery approached 100% for particles in the range of 20–70 μm.
However, for particle sizes smaller than 20 μm, the recovery appears to diminish uniformly
339
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340
FLOTATION FUNDAMENTALS
100
% Recovery from Size Fraction
90
80
70
Galena
Sphalerite
Pyrite
Pentlandite
60
50
40
30
20
10
0
1
2
5
10
20
50
100
200
500
1,000
Average Particle Size, μm
Source: Jowett 1980 (using data from W.J. Trahar).
FIGURE 1
The recovery of sulfide mineral particles after 1 minute of flotation in a batch cell
as the particle size is reduced. Similarly, for particles larger than 70 μm, the recovery
decreases as the size increases. This phenomenon is well known and is reproducible with
many different systems (see, for example, Yianatos, Bergh, and Aguilar 2000).
The size range of particles with maximum floatability depends on the density of the
particles, so that for coal, high recoveries can be achieved at particle sizes up to 350 μm, and
the recovery starts to decrease as the size drops to smaller than about 70 μm. Data, such as
those shown in Figure 1, represent the gross recovery from a given cell, including material
entrained in the froth. Special problems of fine particle flotation have been discussed in
many publications, such as those by Trahar and Warren (1976), Fuerstenau (1980), Somasundaran (1984), Jameson (1984), and Sivarohan (1990). To seek the reasons for the influence of particle size, it will be assumed that the rate-determining step will be the
consequence of a number of phenomena that take place in the liquid phase—the pulp—in
the flotation cell. It is quite possible that the froth will play a role. It is well known that the
froth can be manipulated to reject particles of low grade, which presumably will be attached
less strongly to the bubble surfaces in the froth. However, in this chapter, it is assumed that
behavior in the liquid phase dominates the kinetics of flotation, and that once a particle
enters the froth, it is in effect removed from the cell.
PA R T I C L E C A P T U R E I N Q U I E S C E N T L I Q U I D S
Models for the Capture Efficiency
The overall process of particle capture in the liquid phase is a balance between two competing effects: those of particle collection or attachment, and those of detachment. Two distinct methods have been used to model the flotation process. In one, flotation is modeled as
a sequence of events in which the probabilities of each event are calculated, and the overall
probability of capture efficiency is the product of these probabilities. The attachment process is further broken down into a number of steps, such as induction and sliding, film thinning, and rupture, each with a probability. While this is a valid way of decomposing a very
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THE FLOTATION OF FINE AND COARSE PARTICLES
341
complex phenomenon, an alternative is to regard the capture process in holistic fashion,
solving the equations of motion to find an overall capture efficiency and thereby subsuming
the various mechanisms within an overall model.
For convenience, the probability of collection for the various authors will be expressed
in terms of a collection efficiency E, defined as the ratio of the area of the tube of radius R,
that is coaxial with a bubble whose radius is rb, such that all particles of radius rp that lie
within the collection circle will be captured by the bubble. Thus,
2
2
E = R ⁄ rb
(EQ 1)
Gaudin (1932, 1957) considered the shape of the streamlines around a moving sphere
in the limits of creeping or viscous flow where the Reynolds number, Re, approaches zero,
and what he called turbulent flow, where Re continues to infinity. For the latter case, Gaudin
used the stream function for an inviscid fluid. For the probability of collection, he obtained
Viscous fluid:
E = ( 3 ⁄ 2 ) ( rp ⁄ rb )
Inviscid fluid:
E = 3 ( rp ⁄ rb )
2
(EQ 2)
(EQ 3)
In Sutherland’s (1948) model, a bubble is assumed to be rising in a liquid with a velocity
U. The flow around the bubble is inviscid, so the streamlines can be calculated by irrotational flow theory, assuming that the particles had no effect on the flow field. Sutherland
assumed that if a particle lay on a streamline that would bring it within one particle radius of
the surface of the bubble, collision or capture would occur. He derived a theory for the “collision radius” for a particle of radius rp being captured by a bubble of radius rb. All particles
lying within a circle of this radius, perpendicular to the axis of motion of the bubble, will
collide with it. In the limit as rp << rb, Sutherland showed that the collision radius is given by
R = ( 3r p r b )
0.5
(EQ 4)
which yields a collection efficiency E = 3(rp/rb), which is the same as that of Gaudin, as
would be expected.
These simple models assumed that the presence of the particles did not have an effect
on the streamlines of flow around the sphere. When a particle approaches closely to the
sphere, several phenomena come into play that can be ignored at larger separation distances.
Thus, viscous forces in the thin liquid film become important, because contact between
bubble and particle cannot occur until the film has thinned to the point where it can rupture. In addition, electrical and intermolecular forces become important at small separation
distances. In water, the surfaces of both the bubble and the particle will carry a charge that
may be of the same or opposite sign, so the particle may be attracted to, or repelled from, the
bubble surface. During the final stages of capture, when the film is probably <100 nm thick,
intermolecular forces will also be involved.
In early work, all these effects were lumped together into a single concept described as
the induction time. The induction time was a quantity that was to be measured experimentally, rather than predicted. Sutherland (1948), for example, assumed that if the time of contact of a particle as it traversed the surface of a bubble with which it had made contact was
less than the induction time, true contact or coalescence between the particle and the bubble
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342
FLOTATION FUNDAMENTALS
would not happen. Based on earlier experimental work, Sutherland took the induction time
to be in the range 0.005 to 0.1 sec. As before, a first consideration for collection was that a
particle should lie within the collision circle ahead of the bubble. However, because of the
induction time requirement, there was an additional stipulation that the particle would lie
within a smaller circle, such that the time of contact was at least equal to the induction time.
He derived an expression for the probability of attachment that can be put in terms of the
collection efficiency for a single bubble. Thus,
rp
2
3V λ
E = 3 ⎛ ----⎞ sech ⎛ --------------------------------------⎞
⎝ r b⎠
⎝ 4r b ( 1 + 2r p ⁄ r b )⎠
(EQ 5)
This equation is similar to Equation 3, with the inclusion of a coefficient that is a function of the velocity of approach, V, the bubble radius, rb, and the induction time, λ. The relation between collection efficiency and particle size is quite complex. Since Sutherland’s
time, there have been numerous investigations into the induction time and its effects,
including those of Jowett (1980); Dobby and Finch (1987); Ye and Miller (1988); Crawford and Ralston (1988); Li, Fitzpatrick, and Slattery (1990); Yoon and Yordan (1991);
Nguyen-Van (1993); Vinogradova (1994); Hewitt, Fornasiero, and Ralston (1995);
Nguyen, Kmet, and Schulze (1995); Nguyen, Ralston, and Schulze (1998); Dai, Fornasiero,
and Ralston (1999); Wang et al. (2003, 2004); and Gu, Nandakumar, and Masliyah (2003).
Sutherland’s work was developed further by modelers who relaxed some of his assumptions. Flint and Howarth (1971) solved the equation of motion of a particle approaching a
sphere, assuming that the particle was acted on by viscosity, with a drag calculated by Stokes’
Law. They showed the importance of the dimensionless group K, also known as the Stokes’
number:
2
2ρ p r p U
K = -------------------9μr b
(EQ 6)
where ρp is the density of the particle, U is the relative velocity between the particle and the
liquid, and μ is the viscosity of the liquid. This group represents the ratio of the inertia of
the particle approaching the bubble to the viscous forces acting on it. K is generally very
small for the sizes of bubbles and particles normally found in flotation, suggesting that inertial forces are seldom important. This point was further discussed by Reay and Ratcliff
(1973), who again solved the equations of motion and showed that when the particle–liquid
density ratio varied from 1 to 2.5, the collection efficiency E varied as (rp/rb)2, in agreement
with Gaudin’s result for a viscous liquid.
In works up to this point in time, the motion about the bubble had been assumed to be
either in Stokes’ flow or in inviscid flow. An improvement was made by Weber (1981) and
Weber and Paddock (1983), who calculated the collection efficiencies for bubbles at intermediate Reynolds numbers. Assuming the bubble behaved as a solid sphere, the collection
efficiency was given by
3 rp 2
( 3 ⁄ 16 )Re
E = -- ⎛ ----⎞ 1 + -----------------------------------0.56
2 ⎝ r b⎠
1 + 0.249Re
(EQ 7)
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THE FLOTATION OF FINE AND COARSE PARTICLES
343
1,000
800
0.20
700
600
0.15
500
400
Reynolds Number
Weber and Paddock (1983)
Yoon and Luttrell (1989)
300
200
0.10
0.05
Rise Velocity of Bubble, m/sec
Reynolds Number; Correction Factor
900
0.25
Rise Velocity of Bubble
100
0
0.0
0.5
1.0
1.5
2.0
2.5
3.0
3.5
0.0
4.0
Equivalent Bubble Diameter, mm
FIGURE 2 Bubble rise velocity and corresponding Reynolds number, and the correction
factors calculated by Weber and Paddock (1983) in Equation 7, and Yoon and Luttrell (1989) in
Equation 8, as a function of equivalent bubble diameter
for 0 < Re < 300. Because the bubble Reynolds number Re = 2rbρU/μ depends on the bubble radius and the rise velocity of the bubble (which in turn is a function of the bubble
radius), the correction term is very significant for bubble diameters >150 μm. Yoon and
Luttrell (1989) derived a similar correction using an empirical stream function to represent
the flow, resulting in
r p 2 3 4Re 0.72
E = ⎛ ----⎞ -- + ----------------⎝ r b⎠ 2
15
(EQ 8)
Figure 2 shows a plot of the coefficients of (rp/rb)2 as predicted by Weber and Paddock
(1983), and Yoon and Luttrell (1989), together with the rise velocity of air bubbles in water
and the corresponding Reynolds numbers. It can be seen that the coefficients are quite
important, and the numerical values of each would suggest that the collection efficiency is
very strongly related to the bubble diameter, db, as well as to the radius ratio, rp/rb. A
detailed commentary on the various models described here has been presented by Dai, Fornasiero, and Ralston (2000).
The second term in the brackets on the right-hand side of Equation 7, which describes
the correction to Equation 2 in the Stokes’-flow limit as Re → 0 , is due to the fore-and-aft
asymmetry of liquid flows at intermediate bubble Reynolds numbers. The asymmetry arises
from the nonlinear terms in the Navier–Stokes equations when ∞ > Re > 0. Because of the
asymmetry, the liquid streamlines and the particle trajectories are compressed toward the
front surface of rising bubbles and more relaxed in the bubble rear. The asymmetry also has
important effects on the attachment interactions. Dobby and Finch (1987) showed that the
correction to Equation 7 gives
2
⎧ sin ϕ a ⎫
3 ⁄ 16 Re
3-- ⎛ r p⎞ 2 1 + ----------------------------------- ---------------E =
0.56 ⎨ sin ϕ ⎬
2 ⎝ r b⎠
c⎭
⎩
1 + 0.249Re
(EQ 9)
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344
FLOTATION FUNDAMENTALS
90
80
Collision Angle, degrees
70
60
50
40
Dukhin (1983)
Exact
Bubble Radius rb = 385 μm
Bubble Velocity U = 0.196 m/sec
Particle Density ρp = 2,500 kg/m3
Liquid Density ρL = 1,000 kg/m3
30
20
10
0
0
5
10
15
20
25
30
35
40
45
50
Particle Radius, μm
FIGURE 3 Comparison between results of Dukhin (1983) and the exact numerical solution
(unpublished data) for the collision angle, ϕc, on the mobile surface of rising bubbles
The last term on the right-hand side of Equation 9 describes the attachment efficiency.
The collision angle ϕc depends on the bubble Reynolds number. The attachment angle ϕa is
a function of the induction time, similar to Equation 5. The first two terms on the righthand side of Equation 9 describe the collision efficiency and can be modified to include the
particle inertia, gravitational forces, and other forces (Dobby and Finch 1987). Further analysis shows that the fore-and-aft asymmetry of the particle trajectories around air bubbles is
also influenced by the surface mobility of rising bubbles and inertial forces. The surfactant
molecules adsorbed at the surface of rising bubbles are swept to the bubble rear by the liquid, causing the front surface of rising air bubbles to become mobile while the bubble rear
with the stagnation cap of adsorbed surfactants becomes immobile. The tangential component of the liquid flow on the mobile bubble surface is nonzero and significantly magnifies
the effect of inertial forces governing the particle attachment and collection processes
(Nguyen 1999; Dai et al. 1998). The effect of centrifugal force on the attachment in the
limit of inviscid (potential) liquid flows was analyzed by Dai, Fornasiero, and Ralston
(1999), who employed the approximate results of Dukhin’s analysis (Dukhin 1983). Note
that Dukhin’s results for the collision angle (or the angle of tangency) follow the exact
numerical solution of the particle motion equation only for very fine particles (Figure 3).
For the particle size range (30–100 μm) often encountered in flotation, Dukhin’s results significantly deviate from the exact numerical solution. Further analysis of the particle attachment onto the mobile surface of rising bubbles in flotation is required.
Collection Efficiency and Rate Constant
The collection efficiency, E, can be related to the rate constant, k, for batch flotation.
Assuming first-order kinetics, the equation for the rate of removal of particles from the pulp
( Jameson, Nam, and Young 1977) can be written as
dN p
3QEh
--------- = – ⎛ --------------⎞ N p
⎝ 4d p ⎠
dt
(EQ 10)
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THE FLOTATION OF FINE AND COARSE PARTICLES
345
or
dN p
1 6J g
--------- = – -- E ⎛ -------⎞ N p
4 ⎝ db ⎠
dt
(EQ 11)
where Np is the number concentration of particles in the flotation cell (m–3); Q is the gas
flow rate (m3/sec) into a batch flotation cell of height h (m); dp is the diameter of the particles being floated; and Jg is the superficial gas rate in the cell (m/sec). Jg = Q/A, where A is
the area of cross section of the cell. It is easy to show that 6Jg/db is equal to the surface area
flux, Sb, that is, the rate of flow of bubble surface area per unit area of cross section of the flotation cell. Thus, a rate constant, k, can be written as follows:
3EJ g
ES
k = ----------- = --------b2d b
4
(EQ 12)
Thus, if experiments are carried out in a batch flotation cell and the concentration of
particles is followed with time, it is relatively simple to calculate the rate constant and,
hence, the collection efficiency, E, and determine the effect of particle and bubble size.
The bubble surface area flux in flotation cells depends on the bubble size, which is, of
course, dependent on the way in which the bubbles are made and the surface chemistry of
the flotation pulp, as well as the influence of the particles undergoing flotation. Equation 12
shows that it is pointless to perform batch flotation tests to obtain kinetic data for scale-up
unless the gas superficial velocity, Jg , and the Sauter mean bubble diameter, d32, are measured.
The importance of Equation 12 cannot be underestimated. The rate constant has
dimensions of time–1. The relevant characteristic time for a continuous flotation process is
the residence time, τ, while for a batch process, one could take the half-life, τ50, that is, the
time taken for the concentration of floatable particles in the batch cell to drop to one-half of
the initial value. Thus, an appropriate dimensionless rate constant can be written
ES
k continuous = --------b4τ
(EQ 13)
for a continuous flotation cell or bank, and
ES
k batch = ---------b4τ 50
(EQ 14)
for a batch flotation. Note that the two dimensionless rate constants, in general, will not be
the same. However, batch tests on an ore will enable E to be determined for a given set of
hydrodynamic and interfacial conditions. Scale-up to a larger cell will only be possible if the
hydrodynamic conditions are the same, especially the power per unit volume dissipated in
the impeller region, and if appropriate scaling is applied for differences in Sb and τ.
Measurements of the bubble surface area flux in a wide range of industrial flotation cells
have been reported by Jameson and Allum (1984).
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FLOTATION FUNDAMENTALS
Experiments on Fine Particle Flotation
Reay and Ratcliff (1975) conducted an experimental investigation into the effect of particle
size on the collection efficiency, using glass beads and latex particles, to test their finding
that the rate of flotation should vary as dp2, at constant gas flow rate and bubble size. Using
glass beads of diameters dp between 1 and 30 μm, and polystyrene latex particles of diameters
between 4 and 10 μm. The results for the latex particles showed that E varied approximately
as dp0.44, which completely varies from the expected outcome. It was suggested that the reason for the unexpected result could have been electrostatic double-layer effects, which may
have been important with such small particles of almost neutral buoyancy. With the glass
beads, however, E varied approximately as dp1.5, which is closer to the predicted value for the
exponent. With respect to the dependence of E on bubble diameter, the experimental technique was inconclusive. Bubbles were generated from two frits, giving distributions of mean
size 42 and 71 μm. The measured change in the flotation rate constants for these frits supported the prediction that collection efficiency should vary similar to db2.05.
Collins and Jameson (1976) measured the flotation rate of fine polystyrene particles,
whose diameters ranged from 4 to 20 μm. They studied the rate of flotation of particles in
this range using a size-by-size analysis with a Coulter counter. The bubble size was kept constant at 53 μm with a standard deviation of 9 μm. They found that the collection efficiency
varies as dp to the 1.5 power. Collins (1975) developed a theory based on a method suggested by the work of Spielman and Fitzpatrick (1973), and the results were in good agreement regarding the particle size dependence. For particles greater than about 50 μm,
Collins’ theory predicted that E varies as dp2, essentially the same outcome as that of Reay
and Ratcliff (1975).
Anfruns and Kitchener (1977) generated single bubbles in a suspension of particles of
quartz or glass beads. They measured the collection efficiency of the particles, which were
made hydrophobic by surface methylation. For particles in the range of 10–50 μm and bubbles in the range of 500 μm–1 mm, their results indicate that E varied as dp2/db1.69.
The flotation of silica, pyrite, and galena in a flotation column was studied by DiazPenafiel and Dobby (1994). With silica of d80 35 μm, they found that over the bubble size
– 1.54
) at constant gas
range of 0.8–2.0 mm, the collection efficiency varied as db–0.54( k ∝ d b
flow rate. The exponent is rather smaller than found by others, especially that of Anfruns
and Kitchener (1977).
Recalling that the rate constant is proportional to E/rb, the experimental results can be
expressed as shown in Table 1.
In the work of Diaz-Penafiel and Dobby (1994), the collection efficiency of bubbles in
a flotation column was measured. The bubble sizes were similar to those used in industrial
practice and were significantly larger than those used by previous workers. They argued that
their experiments were conducted in columns with a high gas holdup, as compared with the
single-bubble experiments of others, and that this was the reason for the differences in the
observed collection efficiencies. They provided a graph that showed how their results compared with those of others. The effect of the gas holdup on the collection efficiency follows
the prediction by Nguyen-Van and Kmet (1994).
In Figure 4, data from various authors are plotted, together with the theoretical collection efficiencies predicted by Weber and Paddock (1983) and Yoon and Luttrell (1989).
The bubble terminal velocities required for calculation purposes were taken from Motarjemi and Jameson (1978). The reasons for the small irregularities in the theoretical curves
relate to the uneven behavior of the bubble rise velocity with bubble size, and the complicated
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THE FLOTATION OF FINE AND COARSE PARTICLES
347
TABLE 1 Flotation in a quiescent liquid: Experimental results for variation of rate constant
with particle and bubble diameter
Source
Particle/Bubble Size Range
Rate-Constant Dependence
Reay and Ratcliff (1975)
Latex particles, ρp 1,050 kg/m
3
db = 42 μm
db = 71 μm
k independent of dp
k ∝ dp0.14
42 < db < 71 μm
k ∝ 1/dp2.9
Glass beads, ρp = 2,500 kg/m3
db = 42 μm
db = 71 μm
k ∝ db1.57
k ∝ db2.9
42 < db < 71 μm
k ∝ 1/db2.44
db = 42 μm
k independent of dp
Collins and Jameson (1976)
Latex particles, ρp = 1,050 kg/m3
4 < dp < 30 μm
db < 100 μm
k ∝ dp1.65/db3
Anfruns and Kitchener (1977)
Quartz, ρp = 2,500 kg/m3
4 < dp < 50 μm
600 < db < 1,000 μm
k ∝ dp2/db2.69
Fine coal
11.4 dp < 40.1 μm
60 < db < 560 μm
Refer to Equation 8.
Quartz, d80 35 μm
k ∝ db–1.54
Yoon and Luttrell (1989)
Diaz-Penafiel and Dobby (1994)
0.1
Collection Efficiency
dp 11.4 μm Yoon and Luttrell 1989 (coal)
dp 12.0 μm Anfruns and Kitchener 1977 (silica)
dp 13.0 μm Diaz-Penafiel and Dobby 1994 (silica)
0.01
Weber and Paddock (1983)
0.001
Yoon and
Luttrell (1989)
0.0001
100
1,000
10,000
Bubble Diameter, μm
FIGURE 4 Comparison of experimental results and predictions. Note that the predicted values
do not include corrections for the induction time.
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348
FLOTATION FUNDAMENTALS
dependence on the Reynolds number. The data appear to be consistent although there is
some spread. The values predicted from theory tend to be too high, but the theoretical collection efficiency here does not include any correction for the induction time, so they are
actually collision efficiencies, whereas the experimental values are true collection efficiencies.
The data demonstrate the enormous influence of bubble size—there is an approximately
eightyfold decrease in the collection efficiency when the bubble size increases from 100 to
2,000 μm. Taking into account the dependence shown in Equation 12, the rate constant will
decrease by a factor of 1,600 over the same bubble size range. Thus, in flotation columns, it
is essential to create bubbles that are as small as practicable so as to obtain high recoveries.
The bubbles must not be too small, however, or they run the risk of passing out in the tailings stream from the column. In the design of the Jameson flotation cell, creating bubbles in
the range of 250–300 μm was found to be useful ( Jameson 1988).
A conclusion drawn from these studies is that for fine particles undergoing flotation in
a quiescent liquid or a column, there is a strong influence of particle size and bubble size. At
a constant gas rate and particle size, higher flotation recoveries are achieved with smaller
bubbles.
C O L L E C T I O N O F PA R T I C L E S I N S T I R R E D V E S S E L S
Experiments on the Effect of Particle and Bubble Size
There are very few studies in which the particle size and the bubble size were varied under
conditions of controlled agitation. Ahmed and Jameson (1985, 1989), using polymer latex,
quartz, and zircon particles, investigated the effect of bubble size on flotation kinetics. They
used a series of interchangeable glass frits to generate bubbles of known size: 75, 165, 360,
and 655 μm in diameter. The frit was mounted in the bottom of a baffled tank and agitated
by a Rushton turbine impeller. The particles used and their sizes were as follows: styrenedivinylbenzene copolymer latex, 4–26 μm; quartz, 5–42 μm; and zircon, 5–32 μm. The
performance of the particles was followed on a size-by-size basis using a Coulter counter.
There was a very strong effect of bubble size—the flotation rate increased up to one hundredfold when the bubble size was reduced from 655 μm to 75 μm. However, the results
were much less regular than those in a quiescent liquid. With the polymer latex, it was possible to conduct experiments in a quiescent liquid, but agitation was required in the case of
the quartz and zircon in order to keep the particles in suspension.
The results of the flotation tests differed widely depending on the particle density.
With the polymer latex, the rate constant was found to vary with the particle size with exponents in the range of 1.29 to 1.88, similar to those for the quiescent liquid shown in Table 1,
and with the bubble size db to exponents to the negative powers 1.83 (for dp = 6.5 μm) and
1.13 (dp = 26 μm). If the collection had followed the same mechanism as that in a quiescent
liquid, it would have been expected that k should vary approximately with db–3, but this was
not observed. In the presence of agitation, the rate constants increased uniformly as the
impeller speed increased, suggesting that the agitation led to higher rates of collision but
that detachment forces were insignificant. At the highest speed, the dependence of the rate
constant on the particle diameter had reduced to between dp0.54 and dp0.84. With the quartz
particles, there was clear evidence of particle detachment as the impeller speed increased.
The rate constants reached a maximum in the range 300–500 rpm, although with the largest
bubbles, no maximum was seen. With zircon, the rate constants for the largest bubbles
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THE FLOTATION OF FINE AND COARSE PARTICLES
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increased steadily with increasing speed, but for bubble sizes of 165–655 μm, the flotation
rate reached a maximum at about 200–400 rpm, decreasing when the speed increased to
600 rpm. The authors concluded that, in the presence of agitation, the following were true:
1. Smaller bubbles are more efficient in the flotation of fine particles, but the rate constant is never as strongly dependent as the theory for quiescent liquids would suggest ( k ∝ d b–3 ).
2. At constant agitation, the dependence of the flotation rate on the bubble size is a
complex function of the particle size and density.
3. The effect of the particle size on the rate constant diminishes as the agitation speed
is increased, and the exponent in the relation k ∝ d pn in situations with quartz and
zircon particles never exceeds unity, in agreement with results of Trahar (1981) in
industrial flotation cells.
4. The effect of bubble size is strongly dependent on the density of the particles. Thus,
for polystyrene latex (density 1,050 kg/m3), the exponent m in k ∝ 1 ⁄ d bm varied
from 1.66 to 0.82 when the impeller was changed from low speed (200 rpm) to high
speed (600 rpm). Over the same speed range, with quartz (density 2,650 kg/m3), m
changed from 1.29 to 0.76, and with zircon (density 4,560 kg/m3), m changed from
1.86 to 0.90. In all cases, the rate constant increased with decreasing bubble size.
5. Impeller speed produces strong effects. When operated at zero or low rotational
speed, the flotation cell functioned as a low-height column. With the latex, there
was a tenfold increase in flotation rate when the speed increased to 600 rpm. With
the denser min