Note: Descriptions are shown in the official language in which they were submitted.
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Recovery of Precious Metals
The present invention relates to thiosulfate
leaching of material containing precious metals.
The present invention relates particularly to
thiosulfate leaching of gold from gold-bearing material,
such as ores and concentrates of ores.
It is known to extract gold from ores using
thiosulfate-based lixivient systems. US patents 4,369,061
and 4,269,622 to Kerley describe processes which include
lixiviating with an ammonium thiosulfate leach solution
containing copper to recover gold from ores, particularly
difficult-to-treat ores containing copper, arsenic,
antimony, selenium, tellurium and/or manganese. US
4,654,078 to Perez et al discloses a modification of the
process disclosed in US patent 4,269,622 and is based on
lixiviating ores with copper-ammonium thiosulfate in a
solution that is maintained at a minimum pH of 9.5. Other
known processes that are based on the use of thiosulfate
lixiviants include US patent 5,785,736 to Thomas et al
(assigned to Barrick Gold Corporation) and US patent
5,354,359 to Wan et al (assigned to Newmont Gold Co).
An object of the present invention is to provide
an alternative process for leaching precious metals, such
as gold, using thiosulfate-based lixiviants.
According to the present invention there is
provided a process for leaching precious metals from
material containing precious metals, which process includes
the steps of
(i) treating the material by oxidising
precious metal in the material into a
form that is leachable in a subsequent
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leaching step; and thereafter as a
separate step
(ii) leaching the precious metal with a leach
solution containing a thiosulfate-based
lixiviant.
The present invention is based on the realisation
that high levels of precious metal recovery can be achieved
on a cost-effective basis by carrying out precious metal
oxidation and precious metal leaching as separate steps.
The material may be any material that contains
precious metals.
The present invention relates particularly to
materials in the form of ores and concentrates of the ores.
Preferably, the ores and concentrates are gold
bearing ores and concentrates. The gold may be contained
in oxidic or sulfidic ores.
In one embodiment treatment step (i) includes
forming agglomerates of the precious metal-bearing material
and an oxidant.
Preferably the agglomerates are formed by
contacting the material and a solution containing the
oxidant.
More preferably this embodiment includes forming
agglomerates of the material, a binder, and the oxidant.
More preferably the agglomerates are formed by
mixing the material (such as an ore or concentrate of the
ore) and the binder and thereafter contacting the mixture
with a solution containing the oxidant.
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Preferably, this embodiment includes curing the
agglomerates.
Preferably the curing step is carried out in air
for a period of at least 24 hours.
The treatment step (i) may include forming
agglomerates of the precious metal-bearing material and the
oxidant and a thiosulfate-based lixiviant.
In another embodiment the treatment step (i)
includes forming agglomerates of the precious metal-bearing
material (with or without a binder) and thereafter
contacting the agglomerates with a solution containing the
oxidant.
The treatment step (i) may include contacting the
agglomerates with a solution containing a thiosulfate-based
lixiviant.
In a further embodiment the treatment step (i)
includes contacting the material (without agglomerating the
material first) with a solution containing the oxidant.
The treatment step (i) may include contacting the
material with a solution containing thiosulfate-based
lixiviant.
In each of the above embodiments, preferably the
amount of the solution of the oxidant is relatively small,
typically between 10 and 20%, more preferably, between 12
and 15%, by weight of the weight of the precious metal-
bearing material.
In each of the above embodiments, the treatment
step (i) may include treating the material with ammonia or
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an ammonium salt, such as ammonium carbonate, to stabilise
the oxidant.
The oxidant may be any soluble source of copper
ions.
Preferably, the oxidant is selected from the
group consisting of copper sulfate, copper salt, and
ammonium complex of divalent copper.
The thiosulfate lixiviant may be any suitable
soluble thiosulfate compound.
Preferably the thiosulfate lixiviant is selected
from the group consisting of sodium thiosulfate and
ammonium thiosulfate.
The binder may be any suitable binder, such as a
cement or an organic binder.
The process of the present invention may be
carried out under any suitable pH conditions. In this
connection, the applicant has found in experimental work
that the subject process can be operated over a wider pH
range than prior art processes. Moreover, the applicant
has found that the subject process is more flexible With
operating pH than a number of prior art processes and
consequently pH adjustment may not be necessary - as is the
case in these prior art processes.
The present invention may be carried out on a
heap of precious metal-bearing material, such as gold-
bearing ores and concentrates of the ore, by:
(i) passing the solutiq~ of the oxidant through
the heap;
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(ii) allowing the oxidant solution to drain from
the heap;
(iii)passing the leach solution containing the
thiosulfate-based lixiviant through the
heap; and
(iv) allowing the leach solution containing
leached precious metal to drain from the
heap.
The above sequence of process steps may be
repeated as required to maximise recovery of precious metal
from the heap.
The process may include a further step of
processing the oxidant solution that drains from the heap
to recover the oxidant.
Preferably this step further includes recycling
the oxidant to the process.
The process may also include a further step of
treating the precious metal-bearing leach solution that
drains from the heap to recover precious metal, such as
gold, from the solution.
Preferably, this step also includes recycling
thiosulfate-based lixiviant to the process.
The present invention is not confined to process
precious metal-bearing material in a heap and, by way of
example, extends to other processing options such as
continuously stirred tank reactors.
The process of the present invention can be
applied to both oxidic and sulfidic ores.
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In the c«se of sulfidic ores, the conventional
wisdom in the induf~txy is that such ores are refractory and
that the sulfidic content of the ores must be at least
partially oxidised. However, it has been surprisingly
found by the applicant that the process of the present
invention can be used to selectively oxidise the precious
metal in the ore while minimising or substantially avoiding
oxidation of the sulphide ore to sulfate.
The applicant has carried out experiment work on
gold-bearing oxidic and sulphidic ores. This experimental
work is discussed below.
The experimental work included the following
basic process steps:
Step 1 Copper pretreatment
A solution containing cupric ion (either as
copper, copper diammine or copper tetrammine) in a
predetermined concentration was prepared by dissolving a
predetermined weight of anhydrous copper sulfate in a known
amount of water. To this solution was added either ammonia
(so as to form copper tetrammine) or ammonium carbonate
(AC) or bicarbonate (ABC) (so as to form copper diammine).
This cupric solution thus prepared was contacted with the
ore for a fixed period before separation by filtration
(small scale) or natural draining (columns).
Step 2 Intermediate wash (Optional)
If an intermediate wash was used, a predetermined
volume of a wash solution (either water or ammonia -0.87M)
3S was contacted with the filtered/drained ore for a fixed
period before further filtration/draining.
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Step 3 Thiosulfate wash
The copper pretreated and (when performed) washed
ore was then contacted with a predetermined volume and
concentration of either ammonium or sodium thiosulfate
solution for a fixed period before filtration or draining.
Thiosulfate washing was repeated until little or no Au was
detected in the collected filtrate. In some instances the
ore was left in for extended periods between washes.
EXAMPLE 1.
This example relates to small-scale leaching of
high-grade oxide ore (-. 250ppm Au)
The objective of this experimental work was to
investigate at ambient temperature the influence of:
(i) using CuS04 as a source of Cu2+ as opposed to
different ammine systems (Cu-NH3 to yield
Cu (NH3 ) 42+ or Cu-AC to yield Cu (NH3 ) Za+
(ii) using sodium thiosulfate rather than
ammonium thiosulfate; and
(iii)exposure to air between sequential
thiosulfate washes.
Table 1.1 summarises the series of experiments
performed.
Table 1.1
Copper Pretreat Intermediate Thiosulfate wash
wash
Pretreat Copper
species
Compare copper species
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Cu-NH3 Cu (NH3 ) Water ammonium
4 +
thiosulfate
Cu-AC Cu ( NH3 ) Water ammonium
z'+
thiosulfate
Cu-AC Cu (NH3 ) Ammonia ammonium
z'+
thiosulfate
CuSOd Cuj+ Water ammonium
thiosulfate
Compare thiosulfate
type
CuSOd Cu2+ Water sodium
thiosulfate
The following is a summary of the experimental. conditions.
(i) Wt of ore used (g, dry basis): 64
(ii) Copper pretreatment
~ wt. of copper sulfate (g):1.0(0.025M)
~ Total pretreat volume (ml): 250
~ Contact time with ore before filtration (min):15
~ No. of washes: 1
(iii) Intermediate Wash(when used)
Water:
~ Total Volume (ml): 300
~ Volume per wash (m1):100
~ No of washes: 3
Ammonia solution:
~ Total ammonia pretreat volume (ml): 250
~ Concentration (M): 0.87
~ No. of washes: 1
(iv) Thiosulfate wash
~ Volume per wash (m1):100
~ wt of ammonium thiosulfate(s)(g/100 ml wash)(when
used): 3.7 (0.1M)
~ wt of sodium thiosulfate pentahydrate(s)(g/100 ml
wash)(when used): 6.2 (0.1M)
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_ g _
~ Contact time of wash soln. with ore before filtration
(min):5
~ No of washes . determined by Au content iri filtrate
(usually ~ 8 to 10)
Results are presented in Figures 1.1 and 1.2.
These Figures are plots of cumulative %Au or Cu recovered
in solution versus the number of washes respectively. Where
modifications to the usual sequence in sequential leaching
occurred these are highlighted in Figures 1.1 and 1.2.
Conclusion
~ In all cases with Cu pretreatment (of any form), the
overall Au extraction level is either approaching or
exceeding 90°0. This suggests that high extraction
levels may be achieved with the process of the present
invention regardless of the form of the cupric ion.
~ The rate and extent to which copper desorbs mimics the
trends apparent in gold extraction.
EXAMPLE 2.
This example relates to leaching of as received
and agglomerated low-grade oxide ore (~ 6ppm Au) using
columns.
The most likely field application of the process
of the present invention for low to moderate-grade ores
would be as a heap or vat leach.
In order to investigate this process application,
a series of columns were fabricated using PVC tubing
(D=50mm, L = 350-400mm) and packed with 1 kg of ore (dry
weight basis) as illustrated in Figure 2.1. Column leaching
(which is a form of heap leaching) was then performed using
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the process of the present invention and, to assess its
applicability in the field, several trials of varying
chemical configurat:iox~. were performed.
In general, columns were filled (to completely
cover the bed) by pumping '(from the bottom) or spraying
(from the top) a predetermined volume of liquid (either
pretreatment or leach). After soaking (usually between -. 8
and 24h), the liquid was allowed to drain and the ore
rested (usually between 1-3 days) before the next soak and
rest cycle was begun. Washings were collected and analysed
for Au and Cu by AAS.
The column leach trials involved the use of two
ore forms, generally referred to as:
(i) the as-received ore; and
(ii) agglomerated ore, where the ore was
agglomerated with cement only (usually using
5-6 kg of cement/t of ore.)
To determine the efficiency of column leaching
using the process of the present invention (without the
intermediate washing step) of a low grade oxide (~ 6ppm Au)
ore by varying:
(i) the form of the ore .
agglomerated vs as-received (non-
agglomerated);
(ii) the form of copper in pretreatment:
copper tetrammine vs copper sulfate;
and
(iii)the amount of copper in the copper
pretreatment step.
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The following table (Table 2.1) summarises the
experimental matrix performed.
Table 2.1
Column 'Ore type Weight Copper Thiosulfate
No. (kg) Pretreatment Leach
(dry Form of Cup+/ Concentration
basis) concentration (M)
equivalent
of
CuS04 in
g/1)
Compare
the
form
of copper
in pretreatment
C1 Agglomerated 1 Tetrammine 0.1
(4g/1)
C2 as received 1 Tetrammine 0.1
(4g/1)
C3 Agglomerated 1 CuSOa 0.1
(4g/1)
C4 as received 1 CuSOa 0.1
(4g/1)
Compare
the
amount
of copper
in the
copper
pretreatment
step
C5 Agglomerated 1 CuSOa 0.1
(2g/1)
C6 as received 1 CuSOa 0.1
(2g/1)
Results are presented Figures 2.1a and 2.2a.
These Figures are plots of %Au recovered solution versus
the cumulative weight of recovered solution for the two
comparisons.
Conclusion
Comparison of the form of copper in pretreatment (Cuz+ vs
Cu ( NH3 ) a2+ )
~ The best performed columns for Au extraction are those
where the ore was:
(i) pretreated with copper tetrammine (both
agglomerated or as received ore; or
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(ii) agglomerated and pretreated with CuS04 .
Comparison of the amount of copper in the copper
pretreatment step:
~ Halving the copper concentration of the copper-sulfate
pretreatment appeared to make little difference to Au
extraction rate in the as received ore but reduced
extraction rate in the agglomerated ore by about half
EXAMPLE 3.
This example relates to leaching of co-
agglomerated low-grade oxide ore (-6ppm Au) using columns.
In this example the ore was first pretreated with
copper before subsequent thiosulfate treatment was
performed. To reduce the number of treatment steps and
simplify operation in the field, it may be possible to
apply the required copper component by co-agglomerating it
(in addition to the cement) in the ore and thus avoid the
pretreatment step. Field operation would then require only
thiosulfate washing during extraction. To this end a series
of co-agglomerated ores were prepared where copper (as
copper tetrammine) was added during agglomeration with
cement.
Co-agglomeration was performed in the following
manner:
Columns 7 & 8 Co-Agglomeration with copper.
To 3 kilograms of ore 18g of cement was added.
While this was mixed 400m1s of a solution of 0.00156 moles
/litre of copper as copper tetrammine was added.
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Columns 9 & 14 Co-Agglomeration with copper and
ammonium thiosulfate.
To 3 kilograms of ore 18g of cement Was added.
While this was mixed 200m1s of a solution of 0.00312 moles
/litre of copper as copper tetrammine was added. In
addition to this 200m1s of 0.26M ammonium thiosulfate
solution was added.
Comparing the extraction efficiency of ores co-
agglomerated (besides cement) with either:
(i) small amounts of copper tetrammine (with and
without an added copper pretreatment step);
or
(ii) a combination of copper tetrammine and
thiosulfate.
Leaches were performed in the manner previously
described. The following Table (Table 3.1) presents the
experimental matrix performed..
Table 3.1
..-
Column Ore type Weight Ore Copper Thiosulfate
No. (kg) Bed Pretreatment Leach
(dry L/D Form of Cus'/ Concentration
basis) ratio concentration (M)
equivalent of
CuSOd (g/1)
C7 Co- 1 6.4 None 0.1
agglomerated
With copper
tetrammine
C8 Co- 1 6.6 CuSOa (1 0.1
agglomerated g/1)
with copper
tetrammine
C9 Co- 1 4:9 None 0.1
agglomerated
with copper
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tetrammine
+
thiosulfate
C14 Co- 1 0.26 None 0.1
Agglomerated -
with copper
tetrammine
+
thiosulfate
For comparison
C1 Agglomerate 1 6.6 Copper 0.1
tetrammine
(4g/1)
C3 Agglomerate 1 7 CuSOa 0.1
(4g/1)
Cli Agglomerate 1 0.26 CuSOa 0.1
(4g/1)
Results are presented in Figure 3.1. This Figure
is a plot of %Au recovered versus the cumulative weight of
recovered solution.
Conclusion
~ The best-performed column (wide column) was that where
the ore was co-agglomerated with copper tetrammine and
thiosulfate.
~ Extraction behaviour decayed towards what appeared to
be a limit of about 50%. To determine if the adsorbed
copper level was a limiting factor, the column was
dosed with a treatment of copper ammine before further
thiosulfate washing was undertaken.. Although some
subsequent increase in Au extraction occurred, it
appeared insubstantial and short-lived. This suggested
that, at this crush size, the ore might be limited to
an extraction level of about 50-60°0.
~ The treatments, where the ore was co-agglomerated with
copper tetrammine alone (narrow columns C7, C8) showed
no particular advantage and Were abandoned after about
10 wash cycles. Co-agglomeration in wider columns
appeared to have the "initial kick" observed in small-
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scale experiments.
EXAMPLE 4.
This example relates to leaching of co-
agglomerated low-grade oxide ore (~ 6ppm Au) using columns
without using free ammonia.
The inclusion of ammonia or ammonium into the
leaching system has a beneficial effect during the early
stages of the process of the present invention. However,
in some environments the use of ammonium thiosulfate may
not be feasible because of its unavailability and the use
of free ammonia may also be restricted and sodium
thiosulfate would be used as a source of thiosulfate.
However, if ammonium sulfate (as opposed to thiosulfate) is
freely available it represents a source of
ammonia/ammonium. On this basis, co-agglomerates were
prepared where copper sulfate and ammonium sulfate were co-
agglomerated to mimic the behaviour of copper tetrammine.
Co-agglomeration was performed in the following
manner:
Column 12
To 2.2 kg ore was added llgm cement (5gm/kg).
While mixing, 400m1 of a solution containing 4gm copper
sulfate and l6gm of ammonium sulfate was added. (HIGH
level)
Column 13
To 2.4 kg ore was added l2gm cement (5gm/kg).
While mixing, 400m1 of a solution containing lgm copper
sulfate and 8gm of ammonium sulfate was added. (LOW level)
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Table 4.1 presents the experimental matrix
10
performed.
Table 4.1
Column Ore type weight Ore Copper Thiosulfate
No. (kg) Bed Pretreatment Leach
(dry L/D Form of CuZ+/ Concentration
basis) ratio concentration (M)
equivalent of
CuSOd (g/1)
C12 Co- 1 1.1 None 0.1
agglomerated
With CuSOa
and
( NHa ) xSOa
HIGH level
C13 Co- 1 1.1 None 0.1
agglomerated
with CuSOQ
and
( NHa ) sSOQ
LOW
level
For
comparison
C14 Agglomerate 1 0.26 Tetrammine 0.1
(4g/1)
C11 Agglomerate 1 0.26 CuSO, 0.1
I (4g/1)
C1 Agglomerate 1 6.6 Tetrammine 0.1
(4g/1)
C3 Agglomerate 1 7 CuSOd 0.1
(4g/1)
Results are presented in Figure 4.1. This Figure
is a plot of %Au recovered versus the cumulative weight of
recovered solution.
Conclusion
~ With a co-agglomerated ore using high levels of Cu and
ammonium sulfate, Au extraction behaviour was similar
to that of an ore co-agglomerated with copper
tetrammine+thiosulfate
EXAMPLE 5
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This example relates to leaching of co-
agglomerated low-grade oxide ore (- 6ppm Au) in columns
using a copper tetrammine made from copper sulfate,
ammonium sulfate and sodium hydroxide and thiosulfate as
sodium thiosulfate.
Co-Agglomerated ores were made up as follows:
Ore Total Cement CuS04 Ammonium Adjusted Na2S203.5Hz0
Code ore (kg/t) (anhydrous) sulfate with NaOH (kg/t)
wt (kg/t) (kg/t) to make
(kg) tetrammin
a
404 3 5 2 8 Yes 6.6
405 3 5 2 8 Yes 3.3
Figure 5.1 presents °%Au extracted (based on 6ppm
of Au in ore) versus weight or volume of recovered
lixiviant per wash. Results for Au from the 404 and 405 are
compared with previous best performing columns that had
co-agglomerated ore with Cu-tetrammine+thiosulfate
co-agglomerated ore with CuS04 + Ammonium sulfate (high)
Conclusion
~ The presence of copper tetrammine (made from either
method) and thiosulfate in the co-agglomerated ore
improves the initial rate of extraction. Slight
differences observed between C14 and X-404/X-405 may be
accounted for by differences in the thiosulfate
concentration used in the co-agglomeration step.
~ Based on the recovered solution analysis, the maximum
extraction level was in the order of 50-60%.
~ At the end of the trials, residues from the best
performing columns were fire assayed for Au and the
extraction level calculated. This calculation indicated
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an extraction ~~f 64-67%~, a similar figure to that
determined on v~he as received ore from a cyanide-roll
bottle test (-56%). This suggests that the 'ore crush
size may indeed be a limiting factor.
~ To clarify this, a sample of as received ore was ring-
milled and then leached (in a high concentration
thiosulfate, ammonia containing lixiviant system as per
experiment 8). In this case, extraction level rose to -
77% confirming a limit on extraction due to crush size.
Many modifications may be made to the process of
the present invention described above without departing
from the spirit and scope of the present invention.
'
EXAMPLE 6
This example relates to leaching sulfide ores.
The copper pretreatment conditions were as
follows:
~ copper tetrammine concentration (M): 0.025M
~ ammonia concentration (M): 0.235 - 0.435M
~ Total volume (ml): 250
The thiosulfate was conditions Were as follows:
~ ammonium thiosulfate concentration (M): 0.1
~ volume per wash (ml): 100
Two ore/concentrates were examined: Kanowna Belle
(X-136) and KCGM (X-133). The following effects were
examined:
(i) premilling (by dry ring-milling for 5
minutes (RM))
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(ii) varying the form of Cu2+ in the
pretreatment step ( Cu2+ cf Cu ( NH3 ) 4a+ )
Sequential leaches of pyrite concentrates were
performed as described above with the incorporation of
various treatments. These treatments included:
(i) leaving exposed to air or soaking in
thiosulfate for extend periods;
(ii) increasing the concentration of
thiosulfate in the wash solution ; and
(iii) re-dosing ore with copper tetrammine.
Results based on solution analyses are presented in Figure
6.
Conclusion
~ The highest Au extraction level was ~50-60°o using
unmilled Kanowna Belle (X-136).
~ Premilling appears to inhibit Au extraction although a
greater proportion of copper is adsorbed on the ore
(60-70°o cf 30-40%) .
~ In all cases Cu adsorbed on the concentrate is readily
desorbed.
