A Technical Paper presented at the Canadian Mineral Processors Annual
Meeting, a division of CIMM
(Week of the 18 January 2004)
Crystallex’s Las Cristinas Gold Project
Prepared by,
J.R. Goode, P.Eng., Metallurgical Consultant1
K.G. Thomas, P.Eng., PhD, Chief Operating Officer2
1
2
J.R. Goode and Associates
Crystallex International Corporation
ABSTRACT
The Las Cristinas deposit contains about 10.2 million ounces of gold reserves at a grade of 1.3
g/t. The deposit comprises a layer of fully oxidized saprolite (SAPO), a layer of sulphideenriched saprolite (SAPS), carbonate leached bedrock (CLB) and carbonate stable or un-leached
bedrock (CSB). Gold occurs at about the same level in all lithological groups. The bedrock
units contain minor amounts of copper (about 0.1%) as chalcopyrite, while SAPO contains
virtually no copper it having been leached and deposited in the SAPS zone.
A previous owner started investigation of the Las Cristinas deposit in Venezuela in 1991. Over
the next seven years, 1174 holes with a total length of 159 km were drilled, extensive
metallurgical testwork, including pilot plant operations, were conducted and feasibility studies
were completed. Construction was started in 1997 and again in 1999 but was suspended.
Crystallex has been producing gold in Venezuela since the early 1990’s. In 1997, Crystallex
acquired Inversora Mael which had held two of the claims to the Las Cristinas concessions since
1986. In September 2002, Crystallex entered into a definitive agreement with the Corporación
Venezolana de Guayana, (CVG), to develop the Las Cristinas deposit.
An extensive program of studies and testwork was started in early 2003 to determine an optimum
development plan for Las Cristinas. Mine Development Associates, SNC–Lavalin Engineers
and Constructors, SGS Lakefield Research Limited, J.R. Goode and Associates, and Professor A.
Laplante at McGill University acted as the main contractors.
Because of the potential value of by-product copper, earlier flowsheets used carbon-in-leach for
SAPO, but processed all other ore types by flotation to produce a gold-copper concentrate with
cyanide leaching of cleaner tailings to maximize gold recovery. Cyanide recovery was
necessitated by the high copper content of the leach feed. Initial studies of available data by
Crystallex revealed that direct leaching of all ore types would provide about 11% more gold
recovery than the previously selected flotation route with reduced capital and operating costs.
Bench and pilot scale studies have confirmed that a SAG-ball mill-gravity-CIL route is very
effective for all ore types and will give about 89% gold recovery. This paper describes the
recent testwork, discusses the proposed plant design, and presents economic data.
INTRODUCTION
Venezuela is best known as a major oil producer supplying about 12% of US imports in 2002 or
roughly the same as Saudi Arabia or Mexico. Venezuela is less well known as a gold producer
having produced only 9 t in 2002. Most gold production has come from Bolivar State located
in the south-east of the country and in particular from the El Callao district. Gold is produced
by local mining companies such as Compañia General de Mineria de Venezuela (CVG
Minerven) but, recognizing the potential of the country, several foreign mining companies are
also active in Venezuela including Crystallex International, Hecla Mining, Bolivar Gold, Gold
Reserve, Gold Fields, and China’s Shandong Gold Group.
The Bolivar State deposits are in Archean to
Guayana Shield. The Shield hosts several
including Cambior’s Omai mine in Guyana
deposits are geologically similar to those in
layer overlying primary ore.
early Proterozoic granite-greenstone terrain of the
world class deposits besides those in Venezuela
and its Gross Rosebel mine in Surinam. These
Bolivar State and comprise an oxidized saprolite
Crystallex has been producing gold in Venezuela since the early 1990’s and the company is well
versed in operating mining and milling facilities in the region. In 1997, Crystallex acquired
Inversora Mael which had held the rights to two of the Las Cristinas concessions since 1986.
Figure 1 indicates the location of Crystallex’s current operations and Las Cristinas.
On September 17, 2002,
Crystallex entered into a
definitive Mining Operation
Contract
(MOC)
with
Corporación Venezolana de
Guayana, (CVG), to develop the
Las Cristinas deposit.
The
MOC provides Crystallex with
the exclusive right to explore,
design and construct facilities,
exploit, process and sell gold
from Las Cristinas. An official
translated version of the MOC is
available on the Company’s
website (www.crystallex.com).
Following signing of the MOC,
Crystallex initiated studies of the
significant amount of geological
and metallurgical data that was
available on the deposit. These
studies confirmed the viability of
the project and also suggested an
alternative processing route to
that which had previously been
considered.
In early 2003,
Crystallex shipped several tonnes
Figure 1. Location of Crystallex properties in Venezuela
of samples to Canada for
metallurgical testing to confirm
the suitability of the selected process route.
This paper describes the metallurgical work that has been recently completed and the resulting
development plan for Las Cristinas.
LOCATION AND EXISTING FACILITIES
The Las Cristinas concessions are located in Bolivar state, Venezuela, about 6 km from Troncal
10, the main paved highway that runs from Puerto Ordaz to the Brazilian border. The mining
concessions are in relatively flat terrain ranging from 130 to 160 m above sea level. The
climate is tropical.
The property is provided with an airstrip and a 3000 man construction camp. A major 400 kV
power line supplying power to Brazil parallels the highway and a 150 MVA switching station
was constructed in 2001 for the Las Cristinas project at Las Claritas some 6 km from the site.
GEOLOGY, MINERALOGY, RESERVES, AND MINING
There are two main deposits at Las Cristinas: Conductora/Cuatro Muertos and Mesones/Sofia
(see Figure 2). At Conductora/Cuatro Muertos, gold and copper mineralization are associated
with pyrite-chalcopyrite disseminations, veinlets (2-5% sulphides) and blebs generally oriented
parallel to the foliation. The occurrence of sulphide mineralization is not associated with any
particular rock type, but rather with alteration assemblages that include secondary biotite and a
younger carbonate-epidote assemblage. On a microscopic scale, gold can be found as free
grains in quartz and as blebs and fracture fillings in pyrite and/or chalcopyrite. Silicatecarbonate-sulphide veins tend to parallel foliation. At Mesones/Sofia, gold-copper mineralization
occurs within tourmaline breccia zones, which have obliterated primary tuffaceous textures.
Sulphide concentrations are coarser grained and more chalcopyrite rich than those at
Conductora/Cuatro Muertos.
Extensive weathering has led to
the development of saprolite to
depths of over 90 m locally. The
upper part of the saprolite is
oxidized. Within the oxidized
saprolite, copper has been
predominantly leached, but the
gold remains generally in its
original distribution. The sulphide
saprolite, which has been enriched
in copper leached from the
overlying oxide saprolite, also
retains
the
original
gold
distribution.
The
secondary
copper minerals in the sulphide
saprolite are soluble in cyanide
solutions.
Copper and gold
grade distributions in the bedrock
have not been affected by
weathering.
Figure 2. Plan of Las Cristinas deposit
Earlier exploration work generated a database including information on 1,174 drill holes
covering 160,600 m of drilling, 108 trenches, 162,806 gold assays, 145,221 silver assays,
145,547 total copper assays, and 40,655 cyanide-soluble copper assays. Crystallex determined
that the prior exploration and sampling procedures conformed to or exceeded industry standards.
Nevertheless, Crystallex drilled an additional 2,188 m in twelve diamond drill holes, for a total
of 1,087 core samples, to verify the presence and tenor of mineralization. In addition, 275 quality
assurance/quality control (QA/QC) samples were analyzed. The Crystallex drill results and check
samples corroborate the general tenor of gold mineralization reported by the previous operator.
For additional confirmation, Crystallex re-assayed 262 pre-existing pulps, 200 pre-existing
coarse rejects and 342 pre-existing quarter core samples. Mean grades are similar for both
datasets.
Based on all available data, MDA generated a total resource estimate of 499,000,000 t grading
1.17 g Au/t for a total of 18,807,000 ounces of gold (Measured, Indicated and Inferred) which
agrees well with earlier estimates.
MDA developed a mine plan based on a conventional truck and shovel operation. The total
reserves, using the Canadian Institute of Mining, Metallurgy and Petroleum reserve definitions,
are 246 Mt at a grade of 1.29 g/t for a total of 10.2 million ounces of gold.
From a processing viewpoint, the mass and analysis of the different ore types is important and
especially the mass of the sulphide saprolite. This material contains cyanide soluble copper
minerals which affect processing costs. The distribution of the different lithologies is
summarized in Table 1.
Table 1. Distribution of ore types and grades
Lithology
Saprolite Oxide
Saprolite Sulphide
Carbonate Leach Bedrock
Carbonate Stable Bedrock
Tonnage
Mt
38,868
23,532
54,666
128,593
Percentage of
total reserve
%
16
10
22
53
Gold grade
g/t
1.12
1.38
1.20
1.36
Cyanide
soluble copper
%
~0.005
0.093
~0.01
~0.005
METALLURGICAL TESTWORK
Background
Earlier investigators recognized that cyanide leaching could be economically applied to all ore
types. However, it was decided to additionally recover copper from the deposit and a gravityflotation circuit was developed that would produce a copper-gold flotation concentrate for
custom processing by an off-shore smelter. To give adequate overall gold recovery, it was
necessary to cyanide leach certain flotation products.
The flotation flowsheet was demonstrated in several pilot plant runs operated at solids flowrates
of up to 150 kg/h. As the metallurgical development work continued, an acidificationvolatilization-recovery (AVR) plant was added to the flowsheet because of excessive cyanide
consumption when leaching a copper-bearing flotation intermediate product for gold recovery.
In early 2003, Crystallex and its consultants reviewed available metallurgical test data and
performed various trade-off studies. These analyses indicated that the production and off-shore,
smelting of a copper-gold flotation concentrate, as proposed earlier, was not the preferred
alternative. Direct leaching of most or all of the ore and on-site production of bullion would
give higher gold recovery, simplify the process, improve plant operability, and give lower capital
and operating costs, and a higher Internal Rate of Return (IRR).
Crystallex arranged for new samples to be prepared in Venezuela and SGS Lakefield Research
Limited (Lakefield) was engaged to test the direct leach process. The program ran from the
time that samples arrived at Lakefield in early April until September 2003.
Samples
Nine composite samples of the four different ore types from the Conductora deposit were
prepared from drill core stored at the mine site in Venezuela. Each sample was composited
from individual drill core intervals with a mass of between 0.5 and about 7 kg and averaged
about 2 kg across all samples. A listing of the major samples and their assays is presented in
Table 2.
Table 2. Summary of main Conductora samples tested
Sample
SAPO 1
SAPS1
SAPS2
SAPS(2)
SAPS3
SAPS4
CSB1
CLB/CSB2
Mine
estimate
Au – g/t
1.59
1.55
2.29
1.29
1.64
1.53
1.38
-
Major Assays – Lakefield
Au – g/t
Ag – g/t
Cu – %
CNSCu – %
1.63
1.32
2.16
1.33
2.15
1.84
1.28
1.38
1.2
2.2
1.9
1.4
6.1
1.7
0.9
1.8
0.038
0.14
0.15
0.11
0.21
0.43
0.15
0.14
0.004
0.018
0.033
0.037
0.12
0.31
0.006
.016
Crystallex also shipped samples of Conductora waste and ore from the Mesones deposit.
Graphitic carbon assays were obtained as the difference between CTotal and CCO2 on all samples
and found to be in the range of 0.01 to 0.08%. Preg robbing tests were done on the earlier
samples and CSB, SAPS2, and SAP(2) were found to be mildly preg robbing with 11, 9, and
16% of a 10 ppm spike adsorbed after 24 h. SAPO and the other SAPS samples returned values
of 4% or less. Mercury assays in the various samples were either 0.3 g/t or <0.3 g/t except for
SAPS1 which was reported as 0.4 g/t.
The as-received screen analyses of SAPO, SAPS2 and SAPS 3 were 63, 182, and 69
micrometres m respectively. All other samples were competent rock provided as fragments of
drill core.
Various composites were produced for metallurgical testwork as presented in Table 3. The
composites were made to represent possible operating combinations and to investigate the impact
of varying cyanide soluble copper levels on operating parameters.
Table 3. Main composites used in testwork
Composite name
SAPO-CSB
Comp S1
Comp S2
Comp S3
Comp S4
Comp S5
SAPS350
CLB-CSB Comp. 1
CSB2
CLB-CSB Comp. 2
Mine blend
Composition
20% SAPO1, 80% CSB
10% SAPO1, 10% SAPS2, 80% CSB1 (target 85 ppm CNSCu)
10% SAPO1, 20% SAPS2, 70% CSB1 (target 112 ppm CNSCu)
10% SAPO1, 10% SAPS3, 80% CSB1 (target 172 ppm CNSCu)
10% SAPO1, 20% SAPS3, 70% CSB1 (target 286 ppm CNSCu)
10% SAPO1, 10% SAPS4, 80% CSB1 (target 362 ppm CNSCu)
40% SAPS1, 50% SAPS2, 10% SAPS3 (target 350 ppm CNSCu)
68% CLB, 32% CSB (selected from CLB-CSB sample)
100% CSB (balance of CLB-CSB sample)
27% CLB, 73% CSB
15% SAPO, 5% SAPS (350ppm CNSCu), 10% CLB, 70% CSB
Grinding tests
A great deal of grinding testwork had been performed on Las Cristinas ore by earlier
investigators, including A.R. MacPherson Consultants Ltd. Bond rod and ball mill work index
determinations and abrasion index measurements were made in the recent Lakefield program
which confirmed the earlier data.
The Bond ball mill work index data (metric) for CSB and CLB are 15.3 and 10.5 respectively.
The abrasion indices are about 0.1 g for CLB, and 0.2 g for CSB.
SAPO and SAPS have not been subjected to Bond work index tests because the as-received
material is too fine to test. However, an apparent work index for saprolite can be calculated
from the work index measurements for blends containing this material. This calculation
method yields values between 6 and 11.
Gravity recovery of gold
The feed for bottle roll leach tests were prepared by grinding 2 kg batches of ore to the desired
grind then passing the sample through a 3” Knelson concentrator, then upgrading the concentrate
on a Mozley table. The leach feed was then made by mixing the Knelson and Mozley tailings.
Average data for the different ore types from twenty small-scale gravity recovery tests are
provided in Table 4.
Table 4. Average data from gravity tests ahead of bottle roll leach tests
Sample
SAPO
SAPS
COMP S
SAPO/CSB1 20/80
CSB
CSB depth
CLB/CSB2
Grind
K80, µm
35
50
63
77
67
94
99
Gravity Conc
Wt %
Au, g/t % Rec'y
0.031
252
5.3
0.060
900
18.4
0.097
356
22.9
0.082
278
15.7
0.091
328
22.5
0.086
254
17.2
0.026
1198
22.2
Tail
Au, g/t
1.35
1.39
1.14
1.00
0.96
1.03
1.07
Head, g/t Au
calc.
direct
1.47
1.63
1.71
1.48
1.43
1.19
1.38
1.24
1.24
1.25
1.29
1.38
1.46
The four samples of Mesones CSB and CLB-CSB mixtures were also processed by gravity
concentration and responded well. From an average feed grade of 1.1 g/t, 37% of the gold was
recovered to a 711 g/t concentrate.
As is described later, a pilot plant was operated to process about 1 tonne of Las Cristinas material
over a 20-day period. The first part of the pilot plant run used a feed comprising 20% SAPO
and 80% CSB. The second part used a feed comprising 15% SAPO, 5% SAPS, 10% CLB, and
70% CSB – the Mine Blend. The feed for the pilot plant was prepared in 30 kg batches which
were processed by the same Knelson-Mozley flowsheet as described above. Gravity recovery
data from the pilot plant are provided below in Table 5.
Table 5. Gravity concentration data from pilot plant feed preparation work
Feed
SAPO-CSB
SAPO-CSB
Mine Blend
Phase
PP1-1
PP1-2
PP2
Gold assays – g/t
Head
Tail
Conc
1.50
0.95
775
1.39
0.95
1760
1.38
0.90
1920
Recovery to conc. – %
Mass
Gold
0.071
36
0.025
32
0.025
34
In the first part of the pilot plant (Phase PP1-1), the mass of Mozley concentrate was set at 15 to
25 g per 30 kg batch grind or about 0.07% mass pull. In the later operation (PP1-2 and PP2) the
mass pull was reduced to 5 to 10 g of concentrate or about 0.025% mass. The tabulated
concentrate assays are based on the assay head and the gravity tail assay estimated from the
cyanidation data.
Gravity recovery in the pilot plant was far higher than in the small-scale tests as indicated in
Figure 3. This is as expected and reinforces the importance of processing large samples to
determine the potential for the gravity recovery of gold.
About 40 kg of SAPO and 100 kg of CSB were studied by Professor André Laplante at McGill
University using the Laplante Gravity Recoverable Gold (GRG) test procedure. This work
established that SAPO contained 39% GRG while CSB contained 46%. It was noted that about
10% of the total gold in each sample was –20 m in size and would be difficult to recover.
Based on an analysis of the data, Laplante concluded that about 25% gold recovery would be
obtained by gravity processing.
Gravity gold recovery - %
40
30
20
10
0
-
500
1,000
1,500
2,000
Conc. grade - g/t
80/20 Blend
CSB
SAPO
Comp S1 - S5
CSB depth
SAPS1 - 2
PP1-high mass
PP1-low mass
PP2
Figure 3. Gravity concentration data illustrating sample size effect
Intensive cyanidation of gravity concentrate
Samples of the concentrates produced during the three different segments of the gravity recovery
portion of the pilot plant were subjected to intensive cyanidation using 2% NaCN solution, H2O2
as an oxidant, and a leach time of 48 h. Results are summarized in Table 6.
Table 6. Intensive cyanidation of gravity concentrate
Feed
PP
NaCN
kg/t
Add Cons
1-1
233
74
1-2
240
78
2
260
100
SAPO-CSB
Mine blend
Metal extraction - %
Au
Ag
Au
Ag
Au
Ag
2h
90
95
91
102
101
104
6h
84
90
93
95
96
95
12 h
90
96
101
103
108
101
24 h
95
96
99
100
99
98
Tail
48 h
98.6
95.7
99.5
98.3
99.3
98.3
g/t
6.6
2.3
6.3
2.3
8.5
2.3
Calc
head
g/t
484
54
1,378
138
1,246
136
The intensive cyanidation testwork gave very encouraging results and this method will be used to
process the gravity concentrates at Las Cristinas.
In summary, the data show that gravity recovery should be very effective at Las Cristinas and
give well over 20% recovery of the gold in the feed. The concentrates are very amenable to
intensive cyanidation.
Bottle roll leach tests
All bottle roll tests were preceded by the removal of coarse gold by gravity concentration as
detailed earlier. The results discussed in this section are overall gold recovery, i.e., gravity
recovery plus leach extraction.
An initial series of CIL tests investigated the effects of grind (P80 of 110, 75, and 50
micrometres) and time (12, 24, 48 h) on leaching of the SAPO-CSB blend with a cyanide
strength of 0.5 g/L. This work showed that a grind of 75 micrometres and CIL time of 36 h was
optimum and most additional leach tests were then conducted under said conditions.
A second series of tests on SAPO-CSB looked at cyanide addition strategy and showed that an
initial 0.5 g/L held for 4 h gave low tailings (0.15 g/t) and lower cyanide addition (0.9 kg/t).
Tests on SAPO showed that 99% extraction (tailings of 0.02 g/t) was possible after 36 h of CIL
with 0.9 kg/t NaCN addition. Extraction from SAPO at 24 h was 98% (0.03 g/t tailings). A
36 h leach of CSB gave 85% recovery (0.17 g/t tails) following 0.8 kg/t NaCN addition.
Copper leaching from the SAPO, CSB, and blends was generally less than 5% from heads of
about 0.05% for SAPO and 0.15% Cu(Total) for CSB.
Numerous leach tests were completed on samples of SAPS selected and composited to give a
range of cyanide soluble copper levels between 180 and 1200 ppm. With sufficient cyanide,
the leach extraction of gold varied from 85% to 95%. The extraction of cyanide soluble copper
from the SAPS-bearing material was between 2 and 45% of the analytical cyanide soluble
content of the ore.
The chemical consumption of cyanide and lime in the SAPS tests varied according to the
following equations:
Cyanide addition (kg/t) = 7.93 x Cyanide soluble copper (%) + 0.62 (R2 = 0.83)
Lime addition (kg/t) = 16 x Cyanide soluble copper (%) + 0.36 (R2 = 0.68)
The regression equations were consolidated with data obtained in earlier testwork on Las
Cristinas ore and operating cost factors and the following equation developed for SAPS:
SAPS processing cost ($/t) = 24.2 x Cyanide soluble copper (%) + 2.487
This equation was used to determine the cut-off grade for the SAPS during mine planning and
calculation of the ore reserves.
Pilot plant
Configuration
The CIL pilot plant operation included the batch ball milling of 30 kg charges of feed material
followed by the removal of a gravity concentrate using a 3” Knelson concentrator, upgrading of
the composite on a Mozley table, and combination of the table and Knelson tails. Initially the
table was operated to give a 0.07% mass pull but this was changed early in the operation to a
0.025% mass pull.
Gravity tailings were transferred to a holding tank ahead of the CIL pilot plant where the density
was adjusted and trash removed on a 28# screen. Feed slurry was then pumped at a rate
corresponding to 1.9 kg/h of solids to the first of 6 CIL tanks providing a total of 36 h residence
time. Lime was added to adjust the pH and, for the SAPO-CSB blend, a total of 0.7 kg/t of
NaCN was added as a solution – 67% to the first tank with the balance to the second tank.
During processing of the Mine Blend, which contains CNSCu-containing SAPS, the NaCN
addition was increased to 0.8 kg/t.
Each CIL tank contained 4 g/L of activated carbon during the initial operation. This was
changed to 8 g/L part way through the first pilot plant run (PP1) because it was suspected that the
concentration was too low. Carbon was retained in each tank with a 20 # screen located on the
tank outlet and was manually advanced every 12 h. Based on isotherm and kinetic tests and
modeling studies performed by Lakefield, a carbon loading of 1500 g/t was selected in the design
of the pilot plant operation. The carbon used in each CIL tank prior to start-up was pre-loaded
with gold to ensure rapid attainment of equilibrium. A 28# safety screen was fitted to the last
CIL tank.
A feed sample was taken from each batch of feed to the Knelson concentrator and every 8 h from
the feed to the CIL plant. Tailings were sampled every hour, filtered, and combined to form 4 h
composites. A full profile through the CIL circuit (solids, solution, carbon) was taken every
day and screen analyses were periodically checked.
Results
As noted earlier in Table 6, after adjustment of the Knelson procedure to give a high
concentration ratio, the gravity concentration section of the pilot plant achieved better than 30%
gold recovery to a concentrate assaying more than 1700 g/t.
Lakefield data show that the average gold content of the pilot plant tailings when the plant was at
equilibrium was 0.15 g/t when processing both the SAPO-CSB blend and the Mine Blend.
Corresponding overall gold extraction levels are 89%.
The cyanide addition during the pilot plant operation was set at 0.7 kg/t for the SAPO-CSB blend
and 0.8 kg/t for the Mine Blend. The cyanide consumption was 0.3 kg/t during the last four
days of PP1B and 0.34 kg/t during the last four days of PP2. Residual cyanide concentration
must be added to the chemical consumption to arrive at expected total cyanide addition.
Results are summarized in Table 7 and in the graph presented as Figure 4.
Table 7. Summary of pilot plant data
PP1A
(day 7-9)
20/80 Blend
4
0.71
0.27
0.87
0.95
0.17
36.3
82.2
88.6
Parameter
Feed
C Concentration, g/L
NaCN Addition, kg/t
NaCN Consumption, kg/t
CaO Addition, kg/t
Average CIL Feed Assay, g/t Au
Average CIL Tail Assay, g/t Au
% Gravity recovery
% Extraction in CIL
% Overall recovery
PP1B
(day 10-14)
20/80 Blend
8
0.70
0.30
0.78
0.95
0.15
31.4
84.8
89.6
PP1 O’all
(day 7-14)
20/80 Blend
0.70
0.28
0.84
0.95
0.16
33.0
83.6
89.0
PP2
(day 16-20)
Mine Blend
8
0.77
0.34
0.85
0.90
0.15
34.5
83.6
89.3
100
1.5
95
1.0
90
0.5
Initial feed
20%SAPO 80%CSB
Started Mine Blend
Carbon conc. doubled
85
Hours
Knelson feed
CIL tails
O'all recovery
Figure 4. Pilot plant operation – time variation in major parameters
432
408
384
360
336
312
288
264
240
216
192
168
144
120
96
72
48
80
24
0.0
O'all recovery - %
2.0
0
Gold assay - g/t
Las Cristinas Pilot Plant
Lines are 36 h running averages
Recovery calculated from running averages using 36 h offset
The pilot plant operated quite smoothly although some issues arose and had to be solved.
During the initial operation, it was observed that tailings assays were very low but also that the
tailings were very fine with P80 of 38 micrometres. This was due to the SAPO component of
the ore traversing the leach train at a faster rate than the coarser bedrock particles. Agitator
speeds were changed and equilibrium was attained. Another problem was evidenced by
occasional spikes in the tailings assay. This was found to be due to contamination of tailings
with partially loaded carbon. Tailings samples were rescreened at 28 mesh and operating
procedures changed.
Selected tailings samples were re-leached to determine if longer leach times would be warranted.
An additional 48 h of leaching (more than double the 36 h leach time of the CIL pilot plant) gave
a reduction in tailings assay of between 0.01 and 0.04 g/t suggesting that longer leach times
would not be justified.
Carbon elution
Two samples composited from loaded carbon from the pilot plant were eluted using the high
pressure Zadra approach. Data are summarized in Table 8.
Table 8. Carbon stripping results
Parameter
Loaded carbon assay
Acid washed assay
Eluted carbon assay
Recovery
Unit
g/t
g/t
g/t
%
PP1 Loaded Carbon
Au
Ag
Cu
1552
185
334
1598
306
366
32
1.2
<20
98.0
99.6
94.6
PP2 Loaded carbon
Au
Ag
Cu
1534
287
555
1615
319
364
38
40
20
97.6
87.8
96.4
The data indicate no problems with eluting gold from carbon.
Viscosity measurements
Lakefield measured the viscosity of various Las Cristinas slurries using a Haake rheometer.
The data are discussed here in terms of the critical solids density (CSD) which is defined as the
percentage solids where the yield stress exceeds 8 Pa – a level where slurry handling problems
can be expected.
One series of tests measured the rheology of limed and flocculated mixtures of CSB and SAPO
across a range of percentage solids. The data show that 100% SAPO has a CSD of about 40%
solids, while 70% SAPO-30% CSB has a CSD of about 47%. At 50% SAPO, the CSD is about
52% solids which increases to about 60% solids at 20% SAPO. Figure 5 presents selected
viscosity data.
Data show that the CSD for pure SAPS is about 56% solids which is far higher than the 40%
indicated for SAPO. The CLB and CSB have negligible yield stress (<1 Pa) up to 60% solids.
24
Yield Stress, Pa
20
SAPS 350
CSD
SAPO
CSD
16
SAPO
Operating
domain
12
SAPS 350
Operating
domain
8
4
0
30
33
36
39
42
45
48
51
54
57
60
Solids density, % wt.
SAPO 100% M 368 pH 10.6
SAPS 350 100% M 368 pH 10.6
SAPO 100% M 919 pH 10.6
SAPS 350 100% M 919 pH 10.6
Figure 5. Viscosity data for pure SAPO and SAPS
The rheology data indicate that the percentage solids should be held below 36% solids if pure
SAPO were to be handled but that higher percentage solids can be accommodated as the
percentage SAPO is decreased by blending with CSB and CLB.
Thickening tests
Initial small-scale flocculant scoping tests and static settling tests allowed selection of the
anionic flocculant Magnafloc 919 as suitable for the promotion of thickening of Las Cristinas
slurries. After the preliminary tests, Outokumpu was requested to operate its pilot thickener on
different ore blends.
Outokumpu operated its continuous 0.1 m2 pilot-scale thickener at Lakefield in three campaigns.
In all, Outokumpu conducted 58 thickening tests on nine ore blends ranging from pure SAPO
through various SAPO-SAPS-CLB-CSB blends, to a simple mixture of CSB and CLB.
The results of the Outokumpu tests are summarized in Figure 6. The data show that, with the
correct flocculant, thickener underflow solids concentrations of 50% or greater can be obtained
at a loading rate of 0.47 t/m2/h or lower (scale up to be applied) with all of the ore types/mixtures
that were tested provided that the saprolite content of the thickener feed does not exceed 50%.
0.7
Solids loading rate - t/m2/h
0.6
0.5
0.4
0.3
0.2
0.1
0
20
30
40
50
60
70
Underflow % solids - %
27% CLB-73% CSB
PP1 feed, SAPO-CSB
PP2 feed, M ine Blend
20% SAPO-80% CSB
25% CSB - 75% SAPO
35% SAPO-65% CSB
50% SAPO-50% CSB
50% SAPO-25 SAPS-25 CSB
100% SAPO
Figure 6. Results of Outokumpu pilot thickening tests
The average flocculant dose in all of the tests that were performed was 27 ppm and a dosage of
30 to 40 ppm will probably be needed in the plant. Overflow clarity was generally good and
well under 500 ppm of suspended solids.
Environmental tests
Lakefield completed modified EPA Acid Base Accounting (ABA) tests on SAPO, CSB,
20%SAPO:80%CSB blend (PP1 pilot plant feed), PP1 and PP2 pilot plant tailings, SAPS2
(about 330 ppm CNSCu, 0.7% S), 50% SAPS3:50% SAPS4 (about 2100 ppm CNSCu, 1.2%S),
samples of Mesones ore, and waste rock from Conductora.
Based on the samples that were examined, it was concluded that Conductora SAPO ore and
waste would be classified as non-acid generating and that the SAPS blend with very high
cyanide soluble copper and a SAPS waste sample may be acid generating. The acid generating
potential of the other samples was deemed uncertain.
It should be mentioned that previous investigators performed over 2100 ABA tests on different
rock types from Las Cristinas and concluded that acid generation would not be a problem.
Standard settling test, without rakes, were performed on flocculated but degraded tailings from
pilot plant operations with SAPO-CSB blend and Mine Blend. After seven days, settled solids
reached 60 to 61% solids. Consolidation tests up to 5 bar were performed in a consolidation
(Rowe) cell.
Natural degradation
Natural degradation tests were performed on tailings from the pilot plant operation with SAPOCSB blend (PP1) and Mine Blend (PP2) in two 57 L aquarium located outside at Lakefield.
Results are summarized in Figure 7 presented below. The natural degradation of PP1 was
terminated after 55 d when the CNWAD level had dropped to less than 15 ppm whereas it took the
PP2 tailings, which contained 40% more copper in solution, 98 days to reach the same level.
Assay - ppm
1000
100
10
0
2
4
6
8
10
12
14
Time - weeks
PP1 CNwad
PP1 Cu
PP2 CNwad
PP2 Cu
Figure 7. Removal of CN and Cu from PP tailings by natural degradation
Figure 7 shows that cyanide and copper are quite rapidly removed by natural degradation in the
Lakefield environment. Faster rates are expected at the mine site based on experience gained at
the Crystallex Revemin operation in Venezuela.
Cyanide destruction tests
Nine continuous cyanide destruction tests were performed on naturally degraded tailings solution
from PP1 and PP2 using the INCO Air/SO2 process. CN(wad) levels below 1 ppm were readily
obtained and copper was also effectively eliminated. Retention times of about 30 minutes were
adequate and an SO2 addition of about 5 g/g CN(wad) was effective.
ENGINEERING DESIGN
As the testwork at Lakefield progressed, Crystallex and its consultants reviewed earlier testwork
and data generated by Lakefield and developed the process flowsheet and project design criteria.
The flowsheet will comprise a mineral sizer for crushing saprolite and a gyratory crusher for
bedrock. Crushed material will be fed to a SAG mill – ball mill circuit with gravity recovery
facilities including intensive cyanidation and electrowinning of gravity gold. The overflow
from the grinding circuit cyclones will be thickened, leached in a CIL circuit, gold recovered
using a high pressure Zadra strip circuit, and the CIL tailings sent to the tailings area. Reclaim
water will be returned to the mill and treated for cyanide destruction before re-use in the mill or
discharge to the environment.
Key process design criteria are tabulated below in Table 9.
Table 9. Principal design criteria for process plant
Parameter
Ore throughput
Head grade
Initial ore blend
Gravity recovery
Overall gold recovery
Annual gold recovery
Grind P80 size
Blended SAG Wi
Blended Ball mill Wi
Thickener area
Thickener underflow
CIL residence time
Carbon concentration
Net carbon loading
Carbon advance rate
Lime addition
Cyanide addition
Cyanide destruction
Units
t/a
t/d
g/t Au
%SAPO
Average %
%
oz/a Year 1-5
oz/a Mine life
m
metric
metric
m2/t/h
% solids
h
g/L
g/t Au
t/d
kg/t
kg/t
Data
7,300,000
20,000
1.29
50
21
89
311,000
266,000
70
12.4
11.5
0.4
50
36
6
1500
13
1
0.33
Inco on reclaim water
The numerous mine and infrastructure facilities will not be described here.
ECONOMICS
Capital costs
Project capital costs as developed in the Feasibility Study are presented in Table 10.
Table 10. Summary of Las Cristinas capital costs
Item
Mine
Process Plant
Tailings Management Facility
Infrastructure
Sub-Total Direct Costs
Owner’s Cost
Indirect Costs (including contingency)
Total Costs
VAT1
Total Initial Capital Requirement
Cost Estimate (US$,000)
27,258
80,196
24,490
27,728
160,672
10,000
72,095
242,767
38,843
$281,610
1
VAT of 16.5% has been applied to the total capital costs. This is fully recovered over
the first two and one half years from gold sales revenues.
Operating Costs (at US$325 gold)
Total cash costs for the first five years of production are estimated at US$130 per ounce before
royalties and US$144 per ounce including royalties. Over the life of mine, average total cash
costs are estimated at US$182 per ounce of gold ($6.70/tonne of ore) before royalties and
US$196 per ounce including royalties. Unit operating costs by area are presented below in
Table 11.
Table 11. Las Cristinas operating costs
Item
Mining
Processing
General and administration
Grand total1
1
Operating costs
$/t ore
$/oz gold
$2.94
$80
$3.38
$92
$0.38
$10
$6.70
$182
Excludes royalties; add $14/oz at a gold price of $325/oz.
Financial Analysis
SNC-Lavalin Capital prepared a financial model for the Las Cristinas Project. The model was
run in US dollars with no allowance for inflation. The model includes all capital costs,
operating costs, royalties and a 34% income tax. Depreciation was conservatively assumed for a
20 year life on a straight line basis. An existing investment tax credit of 10% of the
development capital cost was utilized to offset income taxes during the first two years of
production. For simplicity, the model assumed that the Project is financed entirely with equity;
however, the application of project debt will improve the already robust internal rate of
return. The Base Case model used a gold price of $325 per ounce. Key results are presented in
Table 12.
Table 12. Las Cristinas Economics – Before Tax and Unleveraged (US$)
Gold price
Cumulative Free Cashflow
Internal Rate of Return
Payback
$325/oz
$742 million
13.8%
4.7 a
$375/oz
$1,2 billion
19.4%
6.9 a
A sensitivity analysis was performed which considered the impact on the financial results of
changes to the gold price, capital costs and operating costs. The analysis indicated that the Las
Cristinas financial results are most sensitive to changes in the gold price. On a pre-tax basis, a
10% increase in the gold price resulted in a 29% increase in the IRR to 17.8%, while similar
decreases to the capital or operating costs yielded IRR increases of only 15% in both cases.
The Feasibility Study was provided to the CVG in accordance with the Mining Operation
Contract signed in September 2002.
NEXT STEPS – PROJECT IMPLEMENTATION
The next phase of advancing Las Cristinas, which is estimated to extend through the first quarter
of 2004, will focus on awarding an Engineering Procurement and Construction Management
mandate, initiating Detailed Engineering work, completing the Preliminary and Final EIS
reports, and securing the Land Use Permit and the Permit to Impact the Environment.
The Company will also continue to advance the social programs committed to under the terms of
Crystallex’s Mining Operation Contract. These include providing new water treatment
facilities, sewerage systems, houses and road improvements for the local communities.
In addition, the Company will continue to work with its financial advisors to determine the
optimum financing structure and sources of debt financing for Las Cristinas.
ACKNOWLEDGEMENTS
Many people have contributed to the development of the Las Cristinas project to this date
including staff at Crystallex, Lakefield, SLEC, MDA, Knelson, Gekko, Outokumpu, McGill, and
various other service and supply groups. Their input is gratefully acknowledged.