Module - 6
Learning Objectives
• Complex nature of sulphides and sulphide
metallurgy
• Pyrometallurgical extraction process for
copper, zinc, lead, nickel etc.
• Hydrometallurgical extraction processes
• Concept of Process Fuel Equivalent ( PFE)
Methods of metal recovery
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Thermal decomposition
Roasting and subsequent reduction
Controlled roasting  matte  metal
Flash smelting  matte  metal
Metallothermic reduction of sulphide
Hydrometallurgical processing
Chlorination  chlorides  metal
Electrolytic refining of matte to pure metal
(Commercial process for Ni only)
Extraction of copper
Things to Note
During roasting iron sulphides are first oxidized in preference
to copper sulphides.If Fe2O3 forms then this cannot be
slagged. Hence roasting is left incomplete, no copper oxide is
formed. Even some FeS is left behind to ensure that Fe3O4
and Fe2O3 are not formed.
During smelting, residual Fe2O3, if any, is reduced.
10Fe2O3 + FeS
 7Fe3O4 +SO2
3Fe3O4 + FeS
 10FeO +SO2
The aim is to produce only a mixture of copper and iron
sulphides. FeO is removed in slag.
During converting copper metal is produced only through
oxidation of sulphides.
Cu2S(l) + 2Cu2O (l)  6Cu(l) +SO2 (g)
Roasting reactions (~5500 C)
2CuFeS2 + 6.5O2
 2CuO + Fe2O3 + 4SO2
CuFeS2 + 4O2  CuSO4 + FeSO4
2CuFeS2 + O2  Cu2S +2FeS+SO2
2CuFeS2 + 4.5O2  Cu2S + Fe2O3 + 2SO2
2CuFeS2 + 7.5O2  CuO.CuSO4 +2FeSO4 + 2FeSO4 +SO2
3CuFeS2 +9.5O2  3Cu2O + Fe3O4 + 6SO2
6CuFeS2 +13O2  3Cu2S +2Fe3O4 + 9SO2
Smelting Reactions ( ~12500 C) with fluxes
Product 2 layers ( clear cut separation)
Upper
Lower
Slag Layer
‘Matte’ i.e. metallic sulphides
6CuO+4FeS  3Cu2S+4FeO+SO2
2CuSO4+2FeS  Cu2S+2FeO+3SO2
Everything is molten
Cu2O + FeS  Cu2S + FeO
10Fe2O3+FeS  7Fe3O4+SO2
3Fe3O4+FeS  10FeO+SO2
Converting (In side blown converters)
To remove Fe, S and other impurities from matte
Slagging – Exothermic
2FeS+3O2  2FeO+2SO2
2FeO+SiO2  2FeO,SiO2(Fayalite)
Slag – 1-5% (CaO + MgO), Fe – 40–50%
Cu – 2–9%, SiO2 – 20–30%
Blister formation
2Cu2S (l) + 3O2 (g)  2Cu2O + 2SO2 (g)
Cu2S (l) + 2Cu2O (l)  6Cu(l) + SO2 (g)
Overall: 3Cu2S + 3O2  6Cu + 3 SO2
Conventionally, a smelting operation is carried out in reverberatory furnaces fired with either coal or
oil. A typical reverberatory furnace is shown in Fig. Smelting has also been carried out in electric
furnaces. A typical electric furnace is shown in Fig. An electric furnace is more advantageous than a
reverberatory furnace if hydroelectric power is available freely and inexpensively because the
generation of a large volume of combustion gases is avoided. This facilitates both the recovery of
SO2 and the cleaning of the furnace gases, which is generally carried out by an electrostatic
precipitator in order to recover the copper-bearing dust. However, an electric furnace consumes a
large amount of energy, when fossil fuel is burnt especially to generate electricity. It has now given
way to more energy-efficient process, namely, flash smelting and continuous smelting, which are
described later .
In the early days of copper concentrate smelting, the average capacity of the reverberatory furnace was 100 tons per
day, but, at present, the reverberatory or electric furnace can smelt over 1000tons a day.
Converting
The purpose of converting is to remove iron, sulphur and other impurities from matte. For this, the molten matte
produced as a result of smelting is charged into a side-blown converter which is a cylindrical vessel with a capacity of
100-200 tons of matte. A typical vessel is 4 m in diameter and 9 m in length and is lined with a layer of chromemagnesite refractory (about 40 cm thick). A typical side-blown converter is shown in Fig.
In the converter, the atmosphere is highly oxidizing compared with the neutral or mildly oxidizing
atmosphere during smelting. Air or oxygen-enriched air (up to a maximum limit of 32 vol % oxygen
in the blast) is injected into the molten matte through tuyeres. Each tuyere is about 5 cm in
diameter, and there are about 40 tuyeres in a converter. The total volume of gas flowing through
these tuyeres is about 600 m3 / min. The products of the converter are slag and blister copper.
The relative volumes of the two layers can be determined by the lever rule. When
the sulphur level eventually level reaches 1.2 per cent, only the metallic copper
phase remains. At this stage, care to be exercised to ensure that the metal is not
overixidised to Cu2O.
The completion of blow can be
determined by casting a small
sample of the copper and examining
the fracture of this sample. The
blistery appearance of this sample
lends the name blister copper to this
product. In Industrial practice, the
blister produced contains 0.02 – 0.05
per cent S along with 0.2 – 0.5 per
cent dissolved oxygen.
Attempts made in the early days to produce blister copper in the bottom-blown converter used in the
steel industry ended in failure. This is because , after point b is reached, a layer of copper rich liquid will
be formed at the bottom in contact with the tuyers. There will be very little heat generation due to
Cu2S oxidation in this layer, though heat is generated due to oxidation of copper.
4Cu+02 = 2 Cu2O
Although there is not much difference in the heat generated per mole of oxygen for copper oxidation
and that for Cu2S oxidation, the efficiency of copper oxidation is much lower. Consequently, the
temperature drops rapidly in the tuyer region of the Bessemer converter. This leads to the clogging of
the tuyers with solid and the stoppage of the conversion of the matte to metal. It was only after the
introduction of the side-blown Bessemer converter that the conversion of white metal to blister copper
in a converter became possible.
The oxidation taking place during roasting and that resulting from the
leakage of air into the furnace during smelting determine the extent
of oxidation of the iron sulphide in the charge to the slag. Generally ,
the object is to produce a matt that contains 35-45 per cent Cu, 20-22
per cent S, and 25-35 per cent Fe. This not only minimizes the loss of
copper to the slag but also provides a matt with a sufficient quantity
of iron sulphide for use in the next stage, i.e. converting, where iron
sulphide oxidation provides all the heat required to ensure an
autogenous converting operation. The relationship between the
percentage Cu in the slag and that in the matte is given in Fig. for
various reverberatory furnace operations around the world.
Fire refining of blister copper
Sulphur is removed from liquid copper by slow oxidation.
Residual oxygen is eliminated by hydrocarbons using poling
i.e. stirring by green tree branches. These are done in 400 t.
capacity reverberatory furnaces over 12-16 hours, door kept
open to mild air blast
Slow oxidation removes S, Fe, Se and Zn, solid oxides being
skimmed off poling is done at last. Though the method is
crude it is still the most common method. Finally Copper
produced is 99.7 per cent pure.
Electrolytic refining is done in concrete or wooden tanks (B5
x 1.1 x 1 m) using 250-320 Kg copper anode Electrolyte is
copper sulphate (35 g/L) in H2SO4 (2009 g/l) with some
additives (glue, alcohol)
Newer Process
Flash smelting
(combines roasting and
smelting in one unit)
Produces Matte
WORCA Process: Derived by combining the first three alphabets of invertor’s name
(H.K. Worner) and CRA (Conzino Riotinto of Australia Ltd.)
The process seeks to maximize the conservation of energy obtained from smeling
and converting by integrating to a high degree continuous unit operations.
Process Fuel Equivalent (PFE)
The concept was first introduced by Kellogg (1974)
while comparing energy consumption of different
copper smelting processes.
PFE = F + E + S – B
F = Quantity of fuel directly consumed by process
E = The fuel equivalent of electricity (2650 kcal / kwh)
which is the normal energy need to generate power
from fossil fuel
S = The fuel equivalent of major supplies used in
smelting as reagents, oxygen and flukes
B = The fuel equivalent of by products and useful
surplus heat
Hydrometallurgy of copper
~ 85 per cent of world copper production is by pyrometallurgy.
Hydrometallurgy can be employed for oxidized ores or low grade
sulphide ores.
Ferric chloride is an ideal leaching agent.
CuFeS2 + 3FeCl  CuCl + 4FeCl2 + 2S
Not attractive commercially
because energy requirment is
high compared to
pyrometallurgy.
Fig. Ferric Chloride Leaching of
Copper Concentrate
Sulphation roasting
Figure shows stability regions in the Cu-O-S system. In a fluid bed roaster a roasting
temperature higher than 6500C is required to achieve good roasting kinetics. 7000C is optimum
when partial pressure of SO2 lies between 0.04 – 0.09 and Po2 lies between 0.04 – 0.10
During the subsequent electrowinning of CuSO4 solution (obtained from the dilute acid leaching of the roasted
concentrate), H2SO4 is generated. The reaction is
Cu2+SO42-+H2O  Cu + H2SO4+1/2O2
The sulphuric acid in the spent electrolyte is normally used for the vat leaching of copper oxide ores. If such
copper ores are not available, H2SO4 is neutralized with either lime or limestone and rejected as gypsum.
Extraction of Lead
Uses : Antiknock compounds ( ~11%), batteries ( ~35%), sheets
and pipes ( ~11%), sheathing cables ( ~17%), Miscellaneous ( rest)
Common Ores : Galena (PbS) associated with ZnS, FeS, CuS ,
PbCO3 and several precious metals
Process : Roast below 8000 C ( to avoid fusion) in Dwight- Lloyd
sintering machine - sinters for blast furnace smelting using fluxes
( Limestone + Quartz0
PbS + 3/2 O2 = PbO + SO2
SiO2 + 2PbO = 2PbO.SiO2
2PbSO4 + SiO2 = 2 PbO.SiO2 +2SO2
In lead blast furnace ( Temp <12000 C)
PbO + C = Pb + CO
2PbO + C = Pb + CO2
Scrap Iron is also charged in blast furnace
PbSiO3 + Fe = FeSiO2 +Pb
PbO + Fe = FeO + Pb
In the blast furnace smelting
produces four distinct layers
First : Slag ( Sp. Gr ~ 8.6)
Second - Matte containing
copper ( Sp. Gr~5) and other
elements
Third – Spiss FeAs4 + other
impurities ( Sp. Gr ~ 6)
Fourth – Lead ( Sp Gr ~11)
called base bullion
Base bullion contains many
impurities which must be
eliminated systematically.
Modern Developments in Lead Smelting
At present, about 90 per cent of the world’s primary lead is provided by the conventional process,
i.e. sintering followed by blast furnace reduction, and about 10 per cent by the Imperial Smelting
process. Recent environment protection laws, which seek to control the emission of lead fumes
into the atmosphere , coupled with a shortage of energy needed for lead production have
necessitated the development of processes that are both cleaner and less energy consuming. In
this regard, the one-step(continuous) smelting process appears attractive. Here PbS is oxidized to
yield lead according to the reaction.
PbS +O2 --- Pb + SO2
Figure shows the equilibrium
phases in the Pb-S-O system at
12000C. At this temperature ,
metallic lead, with 1-3 per cent
sulphur, can be produced in
presence of pure SO2.
Extraction of Zinc
Uses : Maximum use in protective coating ( galvanizing) because
it forms and impervious ZnCO3 layer . Zinc protects irons by
being more electropositive. Used widely as Cu-Zn brasses Zn is
attractive because of low m.pt, high structural strength, good
dimensional stability.Ideal for die casting
Rolled zinc plates are used in dry cell batteries.
Also used in paints and pigments
Several methods of Production
Sources
Sphalerite(ZnS), Zincite(ZnO), also others
Low grade ores need concentration. ZnS melts at 15000C. However
concentrate ( 55 % Zn) can be roasted at 8000 C, then ground, agglomerated,
sintered to provide feed for retorts. ZnO is reduced ~12000 C, Zinc distils off
for collection in condensers.
Important Points
During roasting one must avoid formation of ferrite ZnO. Fe2O3 which is not
leachable ( but easily reducible)
Since Zn is high is electrochemical series the electrolytic solution must be free
of Cd, Pb, Cu, Fe, Ag and Ge or else Zn deposited on cathode will be
contaminated. Hence purification is necessary prior to electrolysis. For Cu
purification is done afterwards.
Imperial Smelting Process (ISP)
Advantages
• Simultaneously smelt low-grade complex mixed
charges of Zn and Pb ores and concentrates to
recover both Zn and Pb
• Overall thermal efficiency is higher, Zn recovery more
economical
• No electricity is required
• Variety of furnace sizes available
• Furnace operation is continuous and fully automated
• Mechanism is robust – can stand shut downs and
restarts
This process involves the following basic steps :
(1) The treatment of ores to obtain a concentrate rich in zinc.
(2) The roasting of the concentrate to convert the zinc into a soluble form
(3) The treatment of the roasted concentrate to form a zinc sulphate solution
(4) The purification of the zinc sulphate solution by precipitation of impurities
(5) The removal of zinc from purified solution by electrolysis
(6) The melting of zinc sheets to form ingots
Extraction of Nickel
Both oxidic and sulphidic ores are available. Ni is important as an alloying
element – alloys in chemical processing, space research, nuclear reactor
engineering. More than 3000 commercial alloys for mechanical properties
and corrosion resistance. Ni-Cu alloys are known as Monel metals, Ni-Cu-Zn
alloys make German silver.
Sulphide ores – Ni2FeS4, ( NiFe)9S3 etc.
Nickel from Oxide Ores
More abundant than sulphudic ores. Ores are Nickeliferrous laterites typically
~1% Ni and 40-50% Fe in ‘overburden’ , more Ni ( ~1.5%) in layers below in
limonite or goethite, ( FeNi) O, (OH).nH2O. Ni in solid solution with FeO.
In India this is there in chromite mines of Sukinda ( Orissa)
Usual Procedure : Dry, reduce around 800-10000C, reduce Ni – Leach ( acid
or ammonia) - treat by solvent extraction -- ppt. NiCO3--Ni or else
leach – electrolyze
At no stage should ore be heated beyond 10000C.
By keeping controlled
temperature and pco one
produces Ni and FeO ( easy to
leach)
In reduction smelting calcined
ores are smelted in electric
furnace ( 1550-16500C) with
carbon to produce Ferro-nickel
( 25-40% Ni) Gangue silicates
form a slag.
One has to strike a balance
between % Ni and total Nickel
recovery.
There is often Co( ~ 1/10 – 1/20
of Ni) associated with Ni