Note: Descriptions are shown in the official language in which they were submitted.
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TITLE OF THE INVENTION
GOLD AND SILVER RECOVERY FROM POLYMETALLIC
SULFIDES BY TREATMENT WITH HALOGENS
FIELD OF THE INVENTION
[0001] The present invention relates to gold and silver recovery from
polymetallic sulfides by treatment with halogens.
BACKGROUND OF THE INVENTION
[0002] The use of chemical agents, particularly halides, for the
recovery of gold and silver is well known. It was noted very early that the
adjunction of sodium chloride to mercury improved the performances of the
amalgamation process. This discovery translated into the Patio or Cazo
processes, which were implemented on an empirical basis from the early
1600's in Central and South America more than 150 years before the discovery
of elemental chlorine by Scheele in 1774. The Patio method involved the
digestion of a finely divided gold ore with mercury and sodium chloride, in
the
presence of air and moisture over a three month period. The values were then
collected by further leaching with mercury, followed by amalgam distillation
(T.
Egleston, The Metallurgy of Silver, Gold and Mercury in the United States,
Vol.
1, p. 261, John Wiley, 1887).
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[0003] Later, in the late 1700s, chloridizing roasting followed by
barrel amalgamation was developed in Central Europe as an improved method
for gaining access to precious metals from sulfide ores. This process called
upon a high temperature treatment of the gold/silver ores in the presence of
sodium chloride, air and steam, in order to transform the precious metal
sulfides into their corresponding chlorides. The gold and silver was then
recovered either by amalgamation or cementation on pure copper (T. Varley et
al, U.S. Bureau of Mines, Bulletin N 211, 1923). However, it was discovered
that the high temperature chloridizing of gold or silver ores resulted in very
important losses of values by volatilization. In some cases these losses
reached 80 % or more of the precious metal content (S.B. Christy, Transaction
of the American Institute of Mining Engineering, Vol. 17, p. 3, 1888).
[0004] It appeared that the presence of pyrites or iron sulfides
contributed significantly to the volatilization of gold and silver during high
temperature chloridization with NaCI (S. Croasdale, The Engineering and Mining
Journal, August 29, 1903, p. 312). It was finally established that the
mechanism
explaining these losses involves the formation of a mixed chloride of gold and
iron
(AuCI3 = FeCI3), which is highly volatile at chioridization temperatures (J.
A. Eisele
et af. U.S. Bureau of Mines, Report N 7489).
[0005] Elemental chlorine dissolved in water, introduced by Plattner
around 1850, constituted an alternative to high temperature chloridization.
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However, this process was characterized by low efficiency.
[0006] The general characteristics of the various processes involving
chlorine, either as elemental chlorine or as chlorides, either at ambient
temperatures or at high temperatures, were not attractive. The yields obtained
with these processes were generally low (often below 50 %) and the values were
collected as amalgams or as cemented products on copper or iron. In addition,
complex procedures were involved in order to obtain the precious metals in a
pure form. The environmental impacts of such operations, where large amounts
of sulfur are disposed with the tailings, would have been completely
unacceptable
by current standards.
[0007] The advent of cyanide extraction in 1916, terminated the
extraction of gold by various forms of chloridation. The cyanide process calls
upon the action of a cyanide salt such as sodium cyanide on gold in the
presence
of oxygen, to give a soluble gold salt (Eq. I):
2 Au + 4 NaCN + 112 OZ + H2 --> 2 Na[Au(CN)2] + 2 NaOH (Eq. I)
[0008] The gold can then be recovered from the cyanide complex by
the action of excess zinc (Eq. II):
2 Na[Au(CN)Z] + Zn(ex,ess) -+ Na2[Zn(CN)4] + 2 Au (Eq. II)
[0009] Under the best circumstances, gold recovery can be as high as
98 %. This process calls for a contact time of one to three days at near
ambient
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temperature in the presence of air.
[0010] In some instances the cyanide process performs very poorly.
Ores refractory to cyanide extraction can be grouped under the general term of
polymetallic ores. In such ores, one finds small amounts of base metals such
as copper or zinc, typically 0.1% Cu or 0.3% Zn. Such small amounts qualify
the ore as of very low grade for the production of copper or zinc. If such a
polymetallic ore body contains some gold (for example, 4 gfT Au or Ag or a
mixture of both), the cyanide extraction process does not perform well. The
poor performance is due to the base metals, either copper or zinc, (as well as
silver), having a much stronger ability to form complexes with cyanide than
gold. In fact, this inherent property is used to recover gold from a pregnant
solution by zinc treatment following cyanide extraction (see Eq. {f). The base
metals will consume all the cyanide present and the gold extraction will only
begin after all the available base metals, as well as silver, have been
dissolved.
Because of the excessive consumption of relatively costly cyanide, this
process
for recovering gold is uneconomical.
[0011] Polymetallic ores constitute complex mixtures of sulfides. The
tailings discarded as a result of gold and silver extraction using the cyanide
process, as well as by other methods, still contain very substantial amounts
of
sulfur. This sulfur is prone to bio-oxidation (Thiobacillus ferrooxidans), and
the
resulting drainage is quite acidic and toxic due to its metallic content.
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[0012] The spent cyanide solutions, kept in large ponds following gold
recovery, represents a substantial environmental hazard and has recently
created
major disasters in Guyana and Central Europe, thus restricting the use of the
cyanide process in many areas.
5 [0013] In the last twenty years, chloridation has been reconsidered as
a process for extracting base metals such as copper, nickel or silver. The
Intec
Base Metal Process (J. Moyes and F. Houllis, Chloride Metallurgy 2002, Vol.
11, p.
577, Canadian Institute of Mining, Metallurgy and Petroleum) constitutes a
typical
example. This process calls for the digestion at 85 C, over a period ranging
from
12 to 14 hours, of the sulfides of copper or zinc in a concentrated brine
solution
(250 g/I NaCI) comprising a cupric mixed halide (BrCl2)Cu prepared
electrolytically. The mixture is aerated and the copper is collected as
cuprous
chloride. The cuprous chloride is decomposed at the cathode to elemental
copper
by electrolysis upon regeneration of the mixed halide of copper (Eq. !II):
2 CuFeS2 + 5 BrC12 -a 2 Cu+2 + 2 Fe+3 + 4 S + 5 Br + 10 Cl' (Eq. (I1)
[0014] The above described chloridation process was reported as also
extracting gold, if present. However, the requirement of recycling copper so
as to
have the cupric/cuprous system needed to oxidize iron and sulfur, makes this
approach very cumbersome when the main concern is gold recovery rather than
copper recovery. Further, the electrolytical oxidation of sulfur via the
cupric salt,
which is regenerated by electrolysis, is a very costly process rendering the
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treatment of a gold ore having a modest gold content uneconomical. Finally,
the
presence of elemental sulfur in the tailings is a potential source of acid
drainage.
[0015] Another chloridation process called Platsol, was reported as
being very efficient for the recovery of base and precious metals from sulfide
ores
(C.J. Ferron et al, Chloride Metallurgy 2002, Vol. 4, p.11, Canadian Institute
of
Mining, Metallurgy and Petroleum). This process involves a pressure oxidation
in
the presence of oxygen and sulfuric acid in an autoclave at a temperature
above
200 C. The implementation of such a technique is very capital-incentive,
calling
for titanium autoclaves and a source of pure oxygen. The operation of this
equipment is also prone to problems due to scaling of the reactor,
complicating
heat transfer. The sulfur resulting from the operation is in an innocuous
form, i.e.
a hydrated iron sulfate (jarosite). The high capital and operating costs
render this
approach unattractive for polymetallic suffides having a modest gold content.
[0016] Other techniques such as the Plint process (C. Frias et al,
Chloride Metallurgy 2002, Vol. I, p.29, Canadian lnstitute of Mining,
Metallurgy
and Petroleum) or, the Ito process (D.W. Kappes et al, Chloride Metallurgy
2002,
Vol. 1, p.69, Canadian Institute of Mining, Metallurgy and Petroleum), are
techniques used for the recovery of gold and silver from sulfides, by
oxidation
with ferric chloride in concentrated brine. The ferrous chloride is re-
oxidized to
ferric chloride by chlorine alone or by exposure to air and hydrochloric acid
(Eq.
IV):
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2 PbS = Ag2S = 3 SbZS3 + 24 FeC13 -> 24 FeC1Z + 2 PbC12 + 2 AgCI + 6 SbC13 +
12 S (Eq. IV)
[0017] In these processes, sulfur is again oxidized electrochemically
via the oxidation of ferrous chloride by chlorine or HCI. As explained
previously,
such an approach is costly for the recovery of gold or silver from sulfide
ores,
because of the electrochemistry involved. Elemental sulfur is again discarded
with
the tailings, generating a potential source of acid drainage.
[0018] There thus remains a need for an improved method for the
recovery of gold and silver from polymetallic ores.
[0019] The present invention seeks to meet these and other needs.
SUMMARY OF THE INVENTION
[0020] The present invention relates to a method for treating a
polymetailic sulfide ore containing gold andlor silver, and further containing
base metals selected from the group consisting of iron, aluminum, chromium,
titanium, copper, zinc, lead, nickel, cobalt, mercury, tin, and mixtures
thereof,
comprising the steps of:
(a) grinding the polymetallic ore to produce granules;
(b) oxidizing the granules at temperatures of at least about 300 C to
produce oxidized granules;
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(c) chloride leaching the oxidized granules to produce a pregnant
solution of solubilized metal chlorides and a barren solid;
(d) recovering the barren solid from the pregnant solution to produce
a purified pregnant solution; and
(e) selectively recovering gold and/or silver from the purified pregnant
solution yielding a solution essentially deprived of gold and/or
silver.
[0021] The present invention further relates to a method for the
recovery of gold and silver from polymetallic sulfide ores, characterized by
low
operational and cost investments.
[0022] The present invention also relates to a method for the
recovery of gold and silver from polymetallic sulfide ores, characterized by
being carried out at atmospheric pressure and at low oxidation temperatures
prior to leaching.
[0023] In addition, the present invention relates to a method for the
recovery of gold and silver from polymetallic sulfide ores, characterized by
producing tailings devoid of elemental sulfur, sulfides, or soluble sulfates
and
by fast reaction rates allowing for high rates of treatment.
[0024] Furthermore, the present invention relates to a method for the
recovery of precious metals such as gold and silver, as well as base metals
such
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as copper, nickel, cobalt, zinc, tin and lead from polymetallic sulfide ores,
in addition
to relating to the safe removal of sulfur, arsenic and mercury as well as to
the
disposal of iron, chromium, aluminum and titanium in an inert and insoluble
form.
[0024a] The present invention further relates to a method for treating a
polymetallic sulfide ore containing gold or silver, and further comprising a
base
metal selected from the group consisting of iron, aluminum, chromium,
titanium,
copper, zinc, lead, nickel, cobalt, mercury, tin, and mixtures thereof, the
method
comprising:
(a) providing a granulated polymetallic sulfide ore containing
gold or silver having a particle size of less than about 35
mesh;
(b) oxidizing the granulated polymetallic sulfide ore at
temperatures of at least about 300 C to produce oxidized
granules having a sulfur content of about 0.5% or less;
(c) chloride leaching of the oxidized granules, wherein the
chloride leaching involves contacting the oxidized granules
with a leaching solution comprising dissolved elemental
chlorine, a high concentration of chloride, and a catalytic
amount of bromide, to produce a pregnant solution of
solubilized metal chlorides and a barren solid;
(d) recovering the barren solid from the pregnant solution to
produce a purified pregnant solution; and
(e) selectively recovering gold or silver from the purified
pregnant solution,
wherein the method is carried out at atmospheric pressure.
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[0025] Further scope and applicability will become apparent from the
detailed description given hereinafter. It should be understood however, that
this
detailed description, while indicating preferred embodiments of the invention,
is given
by way of illustration only, since various changes and modifications within
the spirit
and scope will become apparent to those skilled in the art.
BRIEF DESCRIPTION OF THE DRAWINGS
[0026] In the appended drawings:
[0027] Figure 1 is a block diagram illustrating the various steps of the
method of the present invention;
[0028] Figure 2 is a block diagram illustrating the various steps of the
sulfur removal aspect of the method of the present invention;
[0029] Figure 3 is a block diagram illustrating the various steps of the
gold and silver recovery aspect of the method of the present invention; and
[0030] Figure 4 is a block diagram illustrating the various steps of the
base metal recovery aspect of the method of the present invention; and
[0031] Figure 5 is a schematic illustration of an electrolytic cell used
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in the method of the present invention.
DETAILED DESCRIPTION OF THE INVENTION
[0032] Unless defined otherwise, the scientific and technological
terms and nomenclature used herein have the same meaning as commonly
5 understood by a person of ordinary skill. As defined herein, the terminology
"recovering" is understood as being an operation resulting in the separation
of a
solid from a liquid. Non-limiting examples of such an operation include
filtration
techniques such as gravity filtration, pressure filtration, vacuum or suction
filtration and centrifugation.
10 [0033] In a broad sense, the present invention relates to a new
method for the recovery of precious metals such as gold and silver from
polymetallic sulfide ores. In an other aspect, the present invention also
relates to
the safe removal of sulfur, arsenic and mercury as well as to the disposal of
iron,
chromium, aluminum and titanium in an inert and insoluble form. This is
achieved
at considerably lower cost than with the current chloridation or cyanide
processes,
by avoiding sulfur oxidation by electrochemical means. The method of the
present
invention is very time efficient, of the order of a few hours, and is carried
out at
atmospheric pressure and at oxidation temperatures of at ieast about 300 C and
preferably ranging from about 400 to about 600 C. The method allows for the
separation of the precious metals as well as the base metals from the common
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metals, while recycling the reagents and releasing only inert waste materials
into
the environment.
[0034] In a preferred embodiment, gold and silver, and optionally base
metals such as copper, zinc, lead, tin, nickel, cobalt and mercury can be
recovered from polymetallic sulfide ores in yields generally well above 80 %
by
the method of the present invention comprising the following preferred steps:
[0035] oxidizing the polymetallic sulfide ore, preferably using lean air
having about 10% 02, at a temperature ranging from about 400 to about 600 C,
to reduce the sulfur content of the ore to about 0.5 % S (as sulfide) or less.
Temperatures above 600 C are also suitable but energy consumption is
increased and sintering of the ore results. The resulting SO2 is fixed by
calcium
carbonate as calcium sulfite, which auto-oxidizes to calcium sulfate dihydrate
(gypsum). This results in the elimination of sulfur in a manner compatible
with
environmental regulations;
[0036] leaching the sulfur-free ore with a near-saturated (275 to 300
g/1) aqueous solution of sodium chloride (sodium brine), or a near saturated
(190
to 225 g/1) aqueous solution of potassium chloride (potassium brine) and
adding a
solution of chlorine/HCI/hypochlorous acid such that the precious metals and
the
base metals are chlorinated and dissolved in the strongly complexing brine
milieu.
The chtoridation reaction is advantageously and significantly accelerated by
the
preferred presence of a catalytic amount, less than one percent of the halides
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present in the brine, of bromide ions. The chlorine/HCI/hypochlorous acid
solution, containing a catalytic amount of bromine, is generated by
circulating a
portion of the brine solution used to slurry the oxidized ore through the
anodic
compartment of an electrolytic cell, at a rate sufficient to dissolve the
chlorine in
the brine solution. Following the slurring operation, the ore is maintained in
suspension in the acidic halogenated brine at a temperature ranging from about
35-45 C by slow stirring, without aeration, for a period of 2-3 hours for most
ores,
and up to 5 hours for exceptionally refractory ores. After separating the
barren
solid followed by washing with brine, the combined filtrate and rinsings are
circulated over activated carbon for gold and silver recovery; and
[0037] treating the solution deprived of precious metals with a sodium
hydroxide solution (or a potassium hydroxide solution if potassium brine was
used) raising the pH to about 2.5-3.5. The sodium hydroxide (or potassium
hydroxide) required to achieve this partial neutralization is produced by
circulating
the initial brine solution through the cathodic compartment of the
electrolytic cell.
The caustic sodium hydroxide solution (or potassium hydroxide solution) is
generated concomitantly at the cathode, in a stoochiometric ratio, with the
chlorine/hydrochloric acid/hypochlorous acid solution produced at the anode of
the electrolytic cell. Raising the pH to about 2.5-3.5 induces the
precipitation of
iron, aluminum, chromium and titanium as insoluble oxides of these metals, in
various hydrated forms. These oxides are filtered and washed with brine.
Raising
the pH of the resulting filtrate to values above 3.5, induces the
precipitation of the
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base metals such as copper, zinc, lead, tin, nickel and cobalt as a base metal
concentrate.
[0038] Any arsenic, often present in significant amounts in polymetallic
sulfide ores, is eliminated along with the sterile solids following leaching
as ferric
arsenate, an insoluble and inert arsenic salt. Mercury, if present, is largely
recovered with the flue dusts after oxidation, and any remaining traces of
this
metal are lixiviated by the chlorinated brine, and recovered on carbon
together
with gold and silver.
[0039] The brine solution, following the removal of the metals, is
recirculated for further leaching. The sterile solids are rinsed with water
and the
rinsings concentrated by evaporation, using waste heat from the sulfide
oxidation
step. The concentrated rinsings, along with the brine solution, are then
recycled
so as to prevent salt losses or salt release into the environment.
Sulfur removal (Figure 2)
[0040] The gold and/or silver containing ore, additionally comprising
variable amounts of base metals such as Cu, Zn, Pb, Sn, Ni, and Co, is a
sulfide
or complex sulfide. The ore may further incorporate one or more other common
metals such as iron, aluminum, titanium, chromium, as well as elements such as
arsenic, antimony or bismuth. Mercury is occasionally also present in the ore.
[0041] The ore is reduced to a particle size of less than about 140
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mesh by standard methods known in the art, such as crushing. The sulfur
content
of the ore, which can be as high as 15 %, is reduced to about 0.5% or less (as
sulfides) by controlled oxidation in a reactor or kiln. The reactor or kiln
provides for
a control of the oxygen content in the reaction chamber. A relatively low
oxidation
temperature, typically ranging from about 400 to about 600 C, is very
advantageous since it prevents any sintering of the material and generates a
solid
product having a large surface area and having good reactivity. This treatment
is
much preferred to standard roasting where temperatures as high as 1200 C have
been observed. Such high reaction temperatures induce much sintering and
volatilization. Standard roasting involves the free burning of the sulfides in
the
presence of excess air.
[0042] The control of the low oxidation temperatures is achieved by
recycling part of the lean air back to the reactor. This allows for the oxygen
content in the reactor to be maintained at values not exceeding 10% 02. It is
important to prevent sodium chloride present in the ore from being oxidized.
It is
well known that sodium chloride contaminations as low as 0.01 percent, can
induce significant volatilization of gold and silver.
[0043] The gas stream from the oxidation reactor is cooled in a settling
chamber, allowing for the collection of volatile oxides such as arsenic oxide,
traces of zinc oxide, and metallic mercury if present in the starting ore, as
well as
other products generated during the oxidative treatment. Dusts carried
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mechanically from the fines in the reactor are also collected in the settling
chamber. The amount of solids collected is generally small; less than one
percent
of the weight of the ore treated. The solids thus collected can be recovered
and
used for recuperation of values such as As2 3 or mercury, or they can be
safely
5 disposed of in sealed containers. The gas at the exit of the settling
chamber,
essentially composed of SO2 and lean air, is partly redirected back to the
oxidation reactor for oxygen level control, and partly directed to a SO2
scrubbing
unit. The SOZ is adsorbed using a finely divided limestone slurry (200 mesh),
allowing for the transformation of essentially all of the SO2 (about 98%) into
10 calcium sulfite, which auto-oxidizes to calcium sulfate dihydrate or
gypsum.
Gypsum is a very stable and inert product representing a definitive solution
for the
safe disposal of sulfur. It can be used as a building material in the
production of
Portland cement or as land fill. The water following the dewatering of the
gypsum
is recirculated back to a water thank. Since gypsum is a dihydrate, there is a
net
15 consumption of water in the scrubbing process. The gases freed of SO2, are
vented through a flue duct.
[0044] In the first step of the method therefore, the ore was made
more reactive towards leaching, and essentially all of the sulfur initially
present
has been disposed of in a safe and environmentally compatible manner. The
present approach constitutes an economically attractive alternative to the
presently available methods. The current cost of electrochemically oxidizing
1%
of sulfur in one metric ton of sulfide ore is $US 4.71 per unit percent of S2-
per ton
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with a KWh at $US 0.09 per kilowatt and with an efficiency of 80 %. The cost
of
oxidizing the sulfide content of an ore containing 10% S2- to elemental
sulfur,
using an electrochemically-produced reagent such as chlorine, would be in the
best case scenario $US 47.10 per ton of ore for power only. The controlled
oxidation of the sulfur content using lean air can be done at 10 % or less of
that
cost, and transforms the sulfur into a safe and environmentally disposable
form.
The electrochemical oxidation process leaves elemental sulfur in the tailings
generating a potential source of acid drainage.
Gold/silver recovery (Figure 3)
[0045] The recovery of gold and silver from the oxidized ore is
achieved by leaching with a reagent comprising elemental halogens. The
halogens (Br2, C12) have significantly different behaviors towards gold.
Bromine
can readily dissolve gold at room temperature, even in the absence of water
(Kruss and Schmidt, Berichte der Deutschen Chemichen Gesellschaft, 20, 2634,
1887). Gold, on the other hand, is inert to dry chlorine at room temperature,
and
the attack of this gas on gold requires the presence of water and slight
heating
(Voigt and Biltz, Z. anorg. Chem., 133, 277, 1924). Even though bromine is the
more reactive reagent with gold, chlorine is more electronegative (W.M.
Latimer,
the Oxidation State of the Elements, pp. 56 and 62, Prentice Hall, 1952):
CI" -a C12 (-1.359 V);
Br ~ Br2 (-1.07 V).
[0046] It is possible to take advantage of this reactivity difference to
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accelerate gold leaching from the oxidized ore, if a catalytic amount of a
bromide
is introduced into the leaching solution. The leaching solution is a brine
solution
having a high concentration of chloride, i.e. from 275 to 300 gli of NaCI or
from
190 to 225 g/l of KCI. Lower salt concentrations yielded lower percentages of
silver recovery, when silver was associated with gold in the oxidized ore. A
portion of the concentrated brine solution, also containing a trace (1-3 g/I)
of NaBr
or KBr, is circuiated in the anodic compartment of an electrolytic cell, at an
appropriate rate, so as to dissolve the halogen liberated at the anode. As
mentioned above, the bromide ion will be reduced first, followed by some
chloride
ions so as to give a mixture of halogens dissolved in the brine solution. The
brine
solution containing dissolved C12 and Br2 is mixed with fresh brine from a
brine
tank to provide a volume of liquid necessary to form a 20% slurry with the
oxidized ore in a reactor kept at 35-45 C. The slurry is slowly stirred in
order to
prevent settling of the ore. The reacting mass was not aerated since aeration
was
neither improving the reaction rate nor the reaction yield, instead it
resulted in the
loss of dissolved halogens. Due to the trace amounts of bromine in the system,
the gold leaching process is believed to involve the initial formation of gold
tribromide (Eq. V):
2Au+ 3 Br2-+ 2AuBr3 (Eq. V)
[0047] The gold tribromide is then believed to be transformed,
because of the stronger oxidizing capacity of C12, into gold trichloride with
the
concomitant regeneration of elemental bromine (Eq. VI):
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2 AuBr3 + 3 CI2 -+ 2 AuCI3 + 3 Br2 (Eq. VI)
[0048] A similar type of reaction is obtained for silver, the high
concentration of chloride allowing the solubilization of the silver halides by
complexation.
[0049] In the course of the leaching reaction, the other ions are
similarfy solubilized, and exist at their maximum valency; copper as cupric
chloride, iron as ferric chloride, tin as stannic chloride, and arsenic as
arsenate
(As+5). Particularly with arsenic, the strong oxidizing environment leads to
the
precipitation of all the arsenic as an insoluble and inert ferric arsenate
(Eq. VII):
Fe3++As04'3 --> FeAsO4 (Eq. V@I)
[0050] The pH of the reaction mixture drops below 0.1 as the leaching
reaction proceeds. This strong acidification is an indication of the reaction
of
chlorine with water (Eq. VIII):
H20 + C12 -* HCI + HOCI (Eq. VNf)
[0051] The presence of hypochlorous acid could account for the
observed chloridation of gold by chlorine in the presence of water. A similar
equation can be written to describe the behavior of bromine, which is in,
equilibrium with hydrobromic acid and hypobromous acid. The solubilized
species
can therefore be seen as a mixture of chlorides and hypochlorides, which
eventually end up as chlorides when the hypochlorous ion decomposes with the
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concomitant evolution of nascent oxygen (Eq. IX):
HOCI -+ HCI + 11202 (Eq. IX)
[0052] The production of nascent oxygen accounts in part for the very
strong oxidizing capability of the system without aeration of any sort.
[0053] The duration of the leaching, preferably at 35-45 C in the
reactor, usually ranges from 2 to 3 hours. With exceedingly refractory ores it
is
necessary to extend the contact time to, for example, about 5 hours. Following
the leaching, the slurry is filtered or centrifuged, producing a pregnant
solution
and a waste or barren solid.
[0054] The barren solid was first rinsed with brine in order to recover
any held-up values in the cake, followed by washing with water to recover any
salt. The so-obtained tailings contain arsenic as an iron arsenate, and are
free of
sulfur and of soluble base metals. The pregnant solution is circulated over
carbon
to collect the gold and silver. Following the recovery of gold and silver from
the
carbon by known methods, these precious metals are obtained by electrowinning
or other standard techniques such as ion exchange and precipitation. The
gold/silver-free solution is then recovered to be further treated so as to
collect the
base metals.
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Recovery of base metals (Figure 4)
[0055] The base metals to be obtained from the leaching of gold-
bearing polymetallic sulfide ores are of two categories. The first category
contains
metals of relatively high commercial value, often obtained by
pyrometallurgical
5 operations. This category contains metals such as nickel, cobalt, copper,
zinc,
lead, tin and mercury. The second category contains metals of low economic
value, and comprises predominantly iron with considerably smaller amounts of
aluminum, titanium, chromium and traces of the p-bloc elements.
[0056] In order to isolate these two types of base metals, sodium
10 hydroxide is generated in the cathodic compartment of the electrolytic
cell. The
sodium hydroxide solution is accumulated in a caustic tank and is then used to
raise the pH of the previously produced barren solution, devoid of gold and
silver,
leaving the carbon columns, from below 1 to about 2.5 to about 3.5. At a pH
ranging from about 2.5 to about 3.5, any iron existing as Fe+3 is
instantaneously
15 precipitated by hydrolysis as a hydrated iron oxide. Titanium, aluminum and
chromium react similarly within this pH range. The hydrated oxides are removed
by filtration. The solids are rinsed with brine in order to recuperate any
base
metals of values held up in the solid cake, followed by washing with water to
remove any traces of salt. The salt-free mixture of oxides is then discarded
as an
20 insoluble and inert material of littie or no commercial vaiue.
[0057] The solution obtained from the filtration and the brine rinsings
CA 02448999 2003-11-12
21
contains the base metals of value. Mercury, if present, was recovered on
carbon
together with gold and silver. The pH of the mercury-free solution, pH between
about 2.5-3.5, is further raised using an additional portion of the sodium
hydroxide
solution to values above 3.5, causing all of the base metals (Ni, Co, Cu, Zn,
Pb,
Sn) to precipitate as oxides or hydrated oxides. The oxides are removed from
the
mixture by filtration and are rinsed with water to remove any traces of salt,
to
provide a concentrate of metals having significant commercial value. The
brine,
being free of metals, is recycled back to the fresh brine reservoir. The
rinsings are
concentrated by evaporation so as to give a brine solution of appropriate
concentration, and which is aiso recycled back to the fresh brine reservoir.
[0058] The implementation of the process of the present invention,
using a large variety of gold-bearing polymetallic sulfide ores, provides for
the
recovery of gold and silver in high yields, essentially always above 80 % and
frequently above 85 %. The process of the present invention also provides for
the
recovery in high yields of the base metals of commercial value, frequently
above
85%.
[0059] Of all the base metals of little commercial value, iron is
generally the predominant one. Following the oxidation of the sulfides at 400-
600 C, the resulting iron oxide is quite inert and no more than about 20-25 %
of
this iron is leached, thus significantly decreasing the power consumption of
the
process. In fact, for a KWh costing US$ 0.09, and with an efficiency at the
CA 02448999 2003-11-12
22
electrolytic cell of 80 %, each percent of iron in the ore would cost US$ 1.00
of
power to take care of, and each percent of base metals such as copper or zinc
in
the ore would cost US$ 2.36 of power to extract. Thus, for an ore having 1 %
copper and 8 % iron, the value of recovered copper (US$ 16.50 at US$ 0.75/lb
for
copper) covers all the electrolytical power costs (US$ 10.36) plus a fair
reserve
and no power imputations have to be made against the gold and silver values
recovered.
[0060] Using the process of the present invention, polymetallic sulfide
ores containing gold and/or silver which do not qualify for base metals
extraction
either because of a low base metal content, problems of enrichments by
flotation
or other restrictions, can be treated economically from the return generated
by the
base metals in order to collect the precious metals. Consequently, the process
of
the present invention provides for an attractive alternative to the currently
available technologies, allowing the treatment of ores or tailings previously
not
attractive, at a profit.
[0061] The recycling of the brine solution, and the disposal of sulfur,
arsenic and metal oxides as stable and inert solids, reduces the environmental
impacts of the operation to a minimum. Furthermore, the implementation of the
process of the present invention at low oxidation temperatures, at near
ambient
chloridation temperatures and at atmospheric pressure, reduces the investment
per unit weight of ore to very competitive values. Finally, the low
temperature
CA 02448999 2003-11-12
23
oxidation of sulfur being an exothermic process, the energy consumption at
that
level is minimal and much lower than the corresponding electrochemical
oxidation
of sulfide to elemental sulfur.
[0062] The process of the present invention was tested using a variety
of polymetallic sulfide ores and tailings containing gold and silver.
Example 1
[0063] A Canadian ore sample (90 g) from the Sudbury (Ontario) area
containing 4.5 g/T Au, 8 g/T Ag, 0.1 % As, 7.5 % S, 5.5 % Fe, 0.1 % Ni, 0.008
Co
and 0.5 % Cu was reduced to a particle size of about 140 mesh and heated at
585-600 C in an atmosphere composed of N2 (50%) and air (50 %), over a period
of two hours in a VycorTM tube heated extemally in a LindbergTM furnace. The
temperature was measured inside the mass being oxidized. The external heating
was reduced when the oxidation began at around 400 C.
[0064] A small amount of a white deposit, arsenic oxide, could be
observed at the discharge side of the VycorTM tube. The color of the oxidized
material changed from black to brown and the weight loss during the process
was
about 12 %.
[0065] A sample of the oxidized material (25.0 g) was placed in a
three-necked one liter flask, along with 500 g of water, 150g of sodium
chloride
and 1.2 g of sodium bromide. The suspension was stirred magnetically and the
CA 02448999 2003-11-12
24
flask was closed so as to exclude air from entering the system.
[0066] The slurry was extracted from the flask through one of the
necks using a peristaltic pump, and was subsequently circulated through the
anodic compartment of an electrolytic cell operating with a brine solution
having
the same concentration as the brine solution in the flask (anode of graphite,
operation at 2.5 V). The anodic fluid was returned to the flask after
dissolving
chlorine. The cell was operated on and off in such a manner as to maintain a
slight reddish coloration in the flask indicative of the presence of free
bromine.
[0067] The reaction flask was maintained at 40 C for a period of 2.5
hours after which it was filtered on a Buchner funnel. The solid was rinsed
three
times with a brine solution containing 300 g/l NaCI. The mixed filtrate and
rinsings
were very acid, having a pH below 1Ø The acidic filtrate and rinsings were
then
treated with 30 g of carbon (NoritT"" R03515) so as to collect gold and
silver. The
barren solid was then rinsed with water to completely remove any traces of
brine
(negative test to AgNO3), dried at 110 C (16.8 g) and submitted to elemental
analysis. The elemental analysis indicated that 96% of the gold and 94% of the
silver initially present in the oxidized material, were leached out and then
adsorbed on the carbon.
[0068] The solution following contacting with carbon was combined
with the aqueous rinsings and was submitted to elemental analysis. The
solution
was found to be essentially free of gold and silver, and contained 99% of the
CA 02448999 2003-11-12
extracted iron, 98 % of the nickel and copper and 91 % of the cobalt present
in
the starting oxidized ore sample. Adjusting the pH to about 3.5 with sodium
hydroxide resulted in the precipitation of the iron. Further raising the pH to
about
8.5 precipitated the nickel, cobalt and copper. The brine, being essentially
free of
5 metals, is available for further use. It was noted by elemental analysis
that the
bromine content in the brine did not change during the process, taking into
account the dilution induced by the rincings. Further, it was found that the
gold
and silver content following treatment (in the sterile residue), was below
0.05 gfT
and 0.16 g/T respectively, while 23% of the iron was extracted.
10 [0069] The process was repeated using several types of polymetallic
sulfide ores containing gold, silver or both, along with base metals of value.
AII the
operational parameters, except the duration of the digestion, were the same as
in
Example 1. Those results are reported in Table I.
Example 2
15 [0070] A sample of ground ore (100 - 200 mesh) from the Pueblo
Viejo deposit (100 g), Dominican Republic, and containing 3.0 g/T Au, 2.25 g/T
Ag, 0.28% Zn, 0.025% As, 5.8% Fe and 4.9% S (as sulfides) was oxidized at
about 600 C for a period of 2 hours in lean air (about 10% 02).
[0071] The oxidized material was then leached using KCI brine (50.0 g
20 of oxidized ore in 500 mL of KCI brine (200 g KCI/L) containing 2.0 g KBr).
The
suspension was stirred at 45 C for a period of two hours, while in the
presence of
CA 02448999 2003-11-12
26
chlorine (0.7 g), added to the slurry at the beginning of the contact.
[0072] The slurry was filtered, the cake rinsed with KCI brine (200 g
KCI/L) and then washed with water. The combined brine filtrate, rinsings and
washings were analyzed for gold, silver and zinc. The gold recovery was of the
order of 87%; the silver recovery was of the order of 61%; and the zinc
recovery
was of the order of 99%. Essentially all of the arsenic was found in the
barren
solid, and none was present in the brine or water rinsings.
[0073] Although the present invention has been described
hereinabove by way of preferred embodiments thereof, it can be modified,
without departing from the spirit and nature of the subject invention as
defined
in the appended claims.
CA 02448999 2003-11-12
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