U.S. patent application number 10/706325 was filed with the patent office on 2004-08-12 for gold and silver recovery from polymetallic sulfides by treatment with halogens.
This patent application is currently assigned to Nichromet Extraction Inc.. Invention is credited to Lalancette, Jean-Marc.
Application Number | 20040156765 10/706325 |
Document ID | / |
Family ID | 34572695 |
Filed Date | 2004-08-12 |
United States Patent
Application |
20040156765 |
Kind Code |
A1 |
Lalancette, Jean-Marc |
August 12, 2004 |
Gold and silver recovery from polymetallic sulfides by treatment
with halogens
Abstract
A method for treating a polymetallic sulfide ore containing gold
and/or 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, is
disclosed. The method comprises the steps of grinding the
polymetallic sulfide ore to produce granules, oxidizing the
granules to produce oxidized granules, and chloride leaching the
granules using a brine solution including dissolved halogens, as
well as chloride and bromide salts.
Inventors: |
Lalancette, Jean-Marc;
(Sherbrooke, CA) |
Correspondence
Address: |
FULBRIGHT & JAWORSKI L.L.P.
600 CONGRESS AVE.
SUITE 2400
AUSTIN
TX
78701
US
|
Assignee: |
Nichromet Extraction Inc.
|
Family ID: |
34572695 |
Appl. No.: |
10/706325 |
Filed: |
November 12, 2003 |
Related U.S. Patent Documents
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Application
Number |
Filing Date |
Patent Number |
|
|
60446517 |
Feb 12, 2003 |
|
|
|
Current U.S.
Class: |
423/40 |
Current CPC
Class: |
Y02P 10/20 20151101;
C22B 11/04 20130101; C22B 3/10 20130101; Y02P 10/234 20151101 |
Class at
Publication: |
423/040 |
International
Class: |
C22B 011/06 |
Claims
1. 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) grinding said polymetallic ore to produce
granules; (b) oxidizing said granules at temperatures of at least
about 300.degree. C. to produce oxidized granules; (c) chloride
leaching said oxidized granules to produce a pregnant solution of
solubilized metal chlorides and a barren solid; (d) recovering said
barren solid from said pregnant solution to produce a purified
pregnant solution; and (e) selectively recovering gold or silver
from said purified pregnant solution yielding a solution
essentially deprived of gold or silver.
2. The method of claim 1, further comprising subsequent treatment
of the solution deprived of gold or silver so as to precipitate and
remove solubilized base metal chlorides.
3. The method of claim 1, wherein in step (b) said oxidizing is
performed using lean air.
4. The method of claim 1, wherein in step (c) said chloride
leaching involves contacting said oxidized granules with a leaching
solution comprising a brine solution including a dissolved
halogen.
5. The method of claim 1, wherein in step (d) said recovering
eliminates the barren solid from the pregnant solution of
solubilized metal chlorides as a filtrate, and wherein the barren
solid is washed with a brine solution to produce washings and a
sterile solid, the washings being combined with the filtrate to
produce said purified pregnant solution.
6. The method of claim 2, wherein said solution deprived of gold or
silver is treated with a caustic solution to produce a first
reaction mixture having a pH ranging from about 2.5 to about 3.5,
further producing a precipitate comprising a first set of base
metals comprising a hydrated metal oxide selected from the group
consisting of iron, aluminum, chromium and titanium, and recovering
said precipitate yielding a first solution essentially devoid of
iron, aluminum, chromium and titanium.
7. The method of claim 6, further comprising the subsequent step of
treating said first solution with a caustic solution to produce a
second reaction mixture having a pH ranging from about 3.5 to about
14, further producing a precipitate including a second set of base
metals comprising a hydrated metal oxide selected from the group
consisting of nickel, copper, cobalt, zinc, lead and tin, and
recovering said precipitate yielding a second solution essentially
devoid of nickel, copper, cobalt, zinc, lead and tin.
8. The method of claim 3, wherein following said oxidizing, said
lean air is cooled in a settling chamber allowing for a volatile
species to be collected; wherein a first portion of said lean air
and sulfur dioxide is recycled from said settling chamber to said
oxidizing step; and wherein a second portion of said lean air and
sulfur dioxide is directed to a sulfur dioxide scrubbing unit.
9. The method of claim 8, wherein in said scrubbing unit said
sulfur dioxide is converted to calcium sulfate dihydrate following
treatment with an aqueous limestone slurry.
10. The method of claim 9, wherein said calcium sulfate dihydrate
is subsequently dried.
11. The method of claim 8, wherein said volatile species comprise
mercury, arsenic oxide and zinc oxide.
12. The method of claim 8, wherein said lean air includes an oxygen
content of about 10%.
13. The method of claim 3, wherein said oxidizing is performed at
temperatures ranging from about 400 to about 600.degree. C.
14. The method of claim 4, wherein a first portion of a brine
solution is circulated through an electrolytic cell to separately
and concomitantly produce a caustic solution and said brine
solution including dissolved halogens, and wherein said brine
solution including dissolved halogens is combined with a second
portion of said brine solution to produce said leaching
solution.
15. The method of claim 14, wherein said brine solution includes a
concentration of sodium chloride ranging from about 275 g/L to
about 300 g/L.
16. The method of claim 14, wherein said brine solution includes a
concentration of potassium chloride ranging from about 190 g/L to
about 225 g/L.
17. The method of claim 1, wherein in step (e) said purified
pregnant solution is treated with carbon to produce a reaction
mixture including a carbon cake rich in gold or silver, and wherein
the carbon cake is removed from the reaction mixture to produce
said solution essentially deprived of gold and silver.
18. The method of claim 17, wherein said gold or silver is stripped
from said carbon cake and wherein said gold or silver is
selectively recovered by a process selected from leaching followed
by electrowinning, and precipitation.
19. The method of claim 5, wherein said sterile solid is washed
with water to produce a salt containing solution, said salt
containing solution being concentrated and recycled to said
leaching step (c).
20. The method of claim 6, wherein said precipitate is washed with
a brine solution to produce washings and a washed residue, said
washings being combined with said first solution essentially devoid
of iron, aluminum, chromium and titanium.
21. The method of claim 7, wherein said second solution essentially
devoid of nickel, copper, cobalt, zinc, lead and tin is recycled to
said leaching step (c).
22. The method of claim 19, wherein said salt containing solution
includes salts selected from the group consisting of sodium
chloride and sodium bromide.
23. The method of claim 19, wherein said salt containing solution
includes salts selected from the group consisting of potassium
chloride and potassium bromide.
24. The method of claim 14, wherein said halogens are selected from
the group consisting of chlorine and bromine.
25. The method of claim 5, wherein said brine solution comprises a
concentration of sodium chloride ranging from about 275 g/L to
about 300 g/L.
26. The method of claim 5, wherein said brine solution comprises a
concentration of potassium chloride ranging from about 190 g/L to
about 225 g/L.
27. The method of claim 14, wherein said brine solution further
includes a bromide salt selected from the group consisting of
sodium bromide and potassium bromide.
28. The method of claim 27, wherein said bromide salt is present in
a catalytic amount.
29. The method of claim 28, wherein said catalytic amount is
ranging from about 1.0 g/L to about 3.0 g/L.
30. The method of claim 4, wherein said chloride leaching is
carried out at ambient temperatures over a period ranging from
about 2 to about 5 hours.
31. The method of claim 30, wherein said ambient temperatures range
from about 35 to about 45.degree. C.
32. The method of claim 6, wherein said caustic solution is a
sodium hydroxide solution.
33. The method of claim 6, wherein said caustic solution is a
potassium hydroxide solution.
34. The method of claim 17, wherein said carbon is activated
carbon.
35. The method of claim 3, wherein said granules have a particle
size ranging from about 35 mesh to about 200 mesh.
36. The method of claim 35, wherein about 80% of said granules have
a particle size of less than 35 mesh and wherein about 20% of said
granules have a particle size of less than 200 mesh.
37. The method of claim 35, wherein about 20% of said granules have
a particle size of less than 35 mesh and wherein about 80% of said
granules have a particle size of less than 200 mesh.
38. The method of claim 13, wherein said oxidized granules have a
sulfur content inferior to about 0.5%.
39. The method of claim 19, wherein said sterile solid may include
ferric arsenate.
40. The method of claim 1, wherein said gold or silver are
recovered in yields in excess of about 80%.
41. The method of claim 1, wherein said polymetallic sulfide ore
comprises gold and silver.
Description
[0001] This application claims the benefit of U.S. Provisional
Application No. 60/446,517, filed Feb. 12, 2003. The entire text of
the above provisional application is specifically incorporated by
reference.
FIELD OF THE INVENTION
[0002] The present invention relates to gold and silver recovery
from polymetallic sulfides by treatment with halogens.
BACKGROUND OF THE INVENTION
[0003] 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 (Egleston, 1887).
[0004] 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 (Varley et al,
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 (Christy,
1888).
[0005] It appeared that the presence of pyrites or iron sulfides
contributed significantly to the volatilization of gold and silver
during high temperature chloridization with NaCl (Croasdale 1903).
It was finally established that the mechanism explaining these
losses involves the formation of a mixed chloride of gold and iron
(AuCl.sub.3.FeCl.sub.3), which is highly volatile at chloridization
temperatures (Eisele et al.).
[0006] Elemental chlorine dissolved in water, introduced by
Plattner around 1850, constituted an alternative to high
temperature chloridization. However, this process was characterized
by low efficiency.
[0007] 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.
[0008] 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):
2Au+4NaCN+1/2O.sub.2+H.sub.2O.fwdarw.2Na[Au(CN).sub.2]+2NaOH (Eq.
I)
[0009] The gold can then be recovered from the cyanide complex by
the action of excess zinc (Eq. II):
2Na[Au(CN).sub.2]+Zn.sub.(excess).fwdarw.Na.sub.2[Zn(CN).sub.4]+2Au
(Eq. II)
[0010] 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 temperature in the presence of air.
[0011] 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 g/T 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. II). 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.
[0012] 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.
[0013] 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.
[0014] 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 (Moyes and Houllis, 2002)
constitutes a typical example. This process calls for the digestion
at 85.degree. C., over a period ranging from 12 to 14 hours, of the
sulfides of copper or zinc in a concentrated brine solution (250
g/l NaCl) comprising a cupric mixed halide (BrCl.sub.2)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. III):
2CuFeS.sub.2+5BrCl.sub.2.sup.-.fwdarw.2Cu.sup.+2+2Fe.sup.+34S.sup.o+5Br.su-
p.-+10Cl.sup.- (Eq. III)
[0015] 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 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.
[0016] Another chloridation process called Platsol, was reported as
being very efficient for the recovery of base and precious metals
from sulfide ores (Ferron et al, 2002). This process involves a
pressure oxidation in the presence of oxygen and sulfuric acid in
an autoclave at a temperature above 200.degree. 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 sulfides
having a modest gold content.
[0017] Other techniques such as the Plint process (Frias et al,
2002) or, the Ito process (Kappes et al, 2002), 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):
2PbS.Ag.sub.2S.3Sb.sub.2S.sub.3+24FeCl.sub.3.fwdarw.24FeCl.sub.2+2PbCl.sub-
.2+2AgCl+6SbCl.sub.3+12.sup.o (Eq. IV)
[0018] In these processes, sulfur is again oxidized
electrochemically via the oxidation of ferrous chloride by chlorine
or HCl. 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.
[0019] There thus remains a need for an improved method for the
recovery of gold and silver from polymetallic ores.
[0020] The present invention seeks to meet these and other
needs.
SUMMARY OF THE INVENTION
[0021] The present invention relates to a method for treating a
polymetallic sulfide ore containing gold and/or 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:
[0022] (a) grinding the polymetallic ore to produce granules;
[0023] (b) oxidizing the granules at temperatures of at least about
300.degree. C. to produce oxidized granules;
[0024] (c) chloride leaching the oxidized granules to produce a
pregnant solution of solubilized metal chlorides and a barren
solid;
[0025] (d) recovering the barren solid from the pregnant solution
to produce a purified pregnant solution; and
[0026] (e) selectively recovering gold and/or silver from the
purified pregnant solution yielding a solution essentially deprived
of gold and/or silver.
[0027] 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.
[0028] 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.
[0029] 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.
[0030] 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 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.
[0031] 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
[0032] In the appended drawings:
[0033] FIG. 1 is a block diagram illustrating the various steps of
the method of the present invention;
[0034] FIG. 2 is a block diagram illustrating the various steps of
the sulfur removal aspect of the method of the present
invention;
[0035] FIG. 3 is a block diagram illustrating the various steps of
the gold and silver recovery aspect of the method of the present
invention; and
[0036] FIG. 4 is a block diagram illustrating the various steps of
the base metal recovery aspect of the method of the present
invention; and
[0037] FIG. 5 is a schematic illustration of an electrolytic cell
used in the method of the present invention.
DETAILED DESCRIPTION OF THE INVENTION
[0038] Unless defined otherwise, the scientific and technological
terms and nomenclature used herein have the same meaning as
commonly 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.
[0039] 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 least about
300.degree. C. and preferably ranging from about 400 to about
600.degree. C. The method allows for the separation of the precious
metals as well as the base metals from the common metals, while
recycling the reagents and releasing only inert waste materials
into the environment.
[0040] 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:
[0041] oxidizing the polymetallic sulfide ore, preferably using
lean air having about 10% O.sub.2, at a temperature ranging from
about 400 to about 600.degree. C., to reduce the sulfur content of
the ore to about 0.5% S (as sulfide) or less. Temperatures above
600.degree. C. are also suitable but energy consumption is
increased and sintering of the ore results. The resulting SO.sub.2
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;
[0042] leaching the sulfur-free ore with a near-saturated (275 to
300 g/l) aqueous solution of sodium chloride (sodium brine), or a
near saturated (190 to 225 g/l) aqueous solution of potassium
chloride (potassium brine) and adding a solution of
chlorine/HCl/hypochlorous acid such that the precious metals and
the base metals are chlorinated and dissolved in the strongly
complexing brine milieu. The chloridation reaction is
advantageously and significantly accelerated by the preferred
presence of a catalytic amount, less than one percent of the
halides present in the brine, of bromide ions. The
chlorine/HCl/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.degree.
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
[0043] 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
stochiometric 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 base
metals such as copper, zinc, lead, tin, nickel and cobalt as a base
metal concentrate.
[0044] 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.
[0045] 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.
[0046] Sulfur Removal (FIG. 2)
[0047] 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.
[0048] The ore is reduced to a particle size of less than about 140
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.degree.
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.degree. 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.
[0049] 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% O.sub.2. 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.
[0050] 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 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
As.sub.2O.sub.3 or mercury, or they can be safely disposed of in
sealed containers. The gas at the exit of the settling chamber,
essentially composed of SO.sub.2 and lean air, is partly redirected
back to the oxidation reactor for oxygen level control, and partly
directed to a SO.sub.2 scrubbing unit. The SO.sub.2 is adsorbed
using a finely divided limestone slurry (200 mesh), allowing for
the transformation of essentially all of the SO.sub.2 (about 98%)
into 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 consumption of water in the scrubbing
process. The gases freed of SO.sub.2, are vented through a flue
duct.
[0051] 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 S.sup.2- per ton 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% S.sup.2- 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.
[0052] Gold/Silver Recovery (FIG. 3)
[0053] The recovery of gold and silver from the oxidized ore is
achieved by leaching with a reagent comprising elemental halogens.
The halogens (Br.sub.2, Cl.sub.2) have significantly different
behaviors towards gold. Bromine can readily dissolve gold at room
temperature, even in the absence of water (Kruss and Schmidt,
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, 1924). Even
though bromine is the more reactive reagent with gold, chlorine is
more electronegative (Latimer, 1952):
Cl.sup.-.fwdarw.Cl.sub.2(-1.359V);
Br.sup.-.fwdarw.Br.sub.2(-1.07V).
[0054] It is possible to take advantage of this reactivity
difference to 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 g/l of NaCl or from
190 to 225 g/l of KCl. 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/l) of NaBr or KBr, is
circulated 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 Cl.sub.2 and Br.sub.2 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.degree. 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+3Br.sub.2.fwdarw.2AuBr.sub.3 (Eq. V)
[0055] The gold tribromide is then believed to be transformed,
because of the stronger oxidizing capacity of Cl.sub.2, into gold
trichloride with the concomitant regeneration of elemental bromine
(Eq. VI):
2AuBr.sub.3+3Cl.sub.2.fwdarw.2AuCl.sub.3+3Br.sub.2 (Eq. VI)
[0056] A similar type of reaction is obtained for silver, the high
concentration of chloride allowing the solubilization of the silver
halides by complexation.
[0057] In the course of the leaching reaction, the other ions are
similarly solubilized, and exist at their maximum valency; copper
as cupric chloride, iron as ferric chloride, tin as stannic
chloride, and arsenic as arsenate (As.sup.+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):
Fe.sup.3++AsO.sub.4.sup.-3.fwdarw.FeAsO.sub.4 (Eq. VII)
[0058] 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. VII):
H.sub.2O+Cl.sub.2.fwdarw.HCl+HOCl (Eq. VIII)
[0059] 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
concomitant evolution of nascent oxygen (Eq. IX):
HOCl.fwdarw.HCl+1/2O.sub.2 (Eq. IX)
[0060] The production of nascent oxygen accounts in part for the
very strong oxidizing capability of the system without aeration of
any sort.
[0061] The duration of the leaching, preferably at 35-45.degree. 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.
[0062] 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.
[0063] Recovery of Base Metals (FIG. 4)
[0064] 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 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.
[0065] In order to isolate these two types of base metals, sodium
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.sup.+3
is instantaneously 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 insoluble and inert material of little or no
commercial value.
[0066] The solution obtained from the filtration and the brine
rinsings 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 also recycled
back to the fresh brine reservoir.
[0067] 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%.
[0068] Of all the base metals of little commercial value, iron is
generally the predominant one. Following the oxidation of the
sulfides at 400-600.degree. 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
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.
[0069] 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.
[0070] 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 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.
[0071] The process of the present invention was tested using a
variety of polymetallic sulfide ores and tailings containing gold
and silver.
EXAMPLES
Example 1
[0072] 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.degree. C. in an atmosphere composed of
N.sub.2 (50%) and air (50%), over a period of two hours in a
Vycor.TM. tube heated externally in a Lindberg.TM. furnace. The
temperature was measured inside the mass being oxidized. The
external heating was reduced when the oxidation began at around
400.degree. C.
[0073] A small amount of a white deposit, arsenic oxide, could be
observed at the discharge side of the Vycor.TM. tube. The color of
the oxidized material changed from black to brown and the weight
loss during the process was about 12%.
[0074] A sample of the oxidized material (25.0 g) was placed in a
three-necked one liter flask, along with 500 g of water, 150 g of
sodium chloride and 1.2 g of sodium bromide. The suspension was
stirred magnetically and the flask was closed so as to exclude air
from entering the system.
[0075] 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.
[0076] The reaction flask was maintained at 40.degree. 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 NaCl. The mixed filtrate and rinsings were very
acid, having a pH below 1.0. The acidic filtrate and rinsings were
then treated with 30 g of carbon (Norit.TM. RO3515) 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
AgNO.sub.3), dried at 110.degree. 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.
[0077] 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 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 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 g/T and 0.16 g/T respectively, while 23%
of the iron was extracted.
[0078] The process was repeated using several types of polymetallic
sulfide ores containing gold, silver or both, along with base
metals of value. All 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
[0079] 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.degree. C. for a period of 2 hours in
lean air (about 10% O.sub.2).
[0080] The oxidized material was then leached using KCl brine (50.0
g of oxidized ore in 500 mL of KCl brine (200 g KCl/L) containing
2.0 g KBr). The suspension was stirred at 45.degree. C. for a
period of two hours, while in the presence of chlorine (0.7 g),
added to the slurry at the beginning of the contact.
[0081] The slurry was filtered, the cake rinsed with KCl brine (200
g KCl/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.
[0082] 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.
1TABLE 1 Treatment of polymetallic ores Precious metals content
Base metals Deposit (g/T) content (%) Sulfur Duration Recovery %
Ex. site Country Au Ag Cu Zn Others content % (hours) Au Ag Cu Zn
Others 2 Zacateca*** Mexico 3.5 8.0 0.3 0.1 Pb: 0.8 7.5 2.0 94 92
98 96 Pb: 91 3 Cassandra* Greece 28 12 -- 1.0 Pb: 1.5 11.0 3.0 96
95 -- 98 Pb: 94 4 Potosi*** Bolivia 3.0 5.8 0.5 -- Sn: 1.9 8.8 3.0
96 92 99 -- Sn: 89 5 Red Lake* Canada 17.0 16.5 0.2 0.8 -- 7.3 2.5
95 96 98 -- -- 6 Rosario*** Dom. 3.37 34.7 0.01 1.1 -- 4.5 3.5 85
91 95 95 -- Republic 7 Moore* Dom. 5.5 8.0 0.01 1.1 -- 6.0 5.0 85
88 98 99 -- Republic 8 Italian Italy 52 5100 1.13 8.06 Pb: 5.18
11.8 2.5 97 87 96 97 Pb: 99 Smelter** Hg: 1130 ppm Hg: 99.9 9 Rio
Spain 231 248 25.2 0.39 Pb: 0.14 18.5 3.5 98 96 99 95 Pb: 92
Narcea** *fresh ore; **concentrate; ***tailings
References
[0083] The following references, to the extent that they provide
exemplary procedural or other details supplementary to those set
forth herein, are specifically incorporated herein by
reference.
[0084] Christy, Transaction of the American Institute of Mining
Engineering, 17:3, 1888.
[0085] Croasdale, J. Engineering and Mining, 312, 1903.
[0086] Egleston, In: The Metallurgy of Silver, Gold and Mercury in
the United States, 1:261, John Wiley, 1887.
[0087] Eisele et al. U.S. Bureau of Mines, Report N.sup.o 7489.
[0088] Ferron et al, In: Chloride Metallurgy, Vol. I:11, Canadian
Institute of Mining, Metallurgy and Petroleum, 2002.
[0089] Frias et al, In: Chloride Metallurgy, Vol. I:29, Canadian
Institute of Mining, Metallurgy and Petroleum, 2002.
[0090] Kappes et al, In: Chloride Metallurgy, Vol. I:69, Canadian
Institute of Mining, Metallurgy and Petroleum, 2002.
[0091] Kruss and Schmidt, Berichte der Deutschen Chemichen
Gesellschaft, 20:2634, 1887.
[0092] Latimer, In: Oxidation State of the Elements, 56-62,
Prentice Hall, 1952.
[0093] Moyes and Houllis, In: Chloride Metallurgy, Vol. II:577,
Canadian Institute of Mining, Metallurgy and Petroleum, 2002.
[0094] Varley et al, U.S. Bureau of Mines, Bulletin N.sup.o 211,
1923.
[0095] Voigt and Biltz, Z. anorg. Chem., 133:277,1924.
* * * * *