CA2409378C - Method for thiosulfate leaching of precious metal-containing materials - Google Patents

Method for thiosulfate leaching of precious metal-containing materials Download PDF

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CA2409378C
CA2409378C CA002409378A CA2409378A CA2409378C CA 2409378 C CA2409378 C CA 2409378C CA 002409378 A CA002409378 A CA 002409378A CA 2409378 A CA2409378 A CA 2409378A CA 2409378 C CA2409378 C CA 2409378C
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Prior art keywords
thiosulfate
leach solution
precious metal
pregnant
thiosulfate leach
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CA2409378A1 (en
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Christopher Andrew Fleming
Ralph Peter Hackl
Jinxing Ji
Paul George West-Sells
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Priority to CA002617415A priority patent/CA2617415C/en
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    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B11/00Obtaining noble metals
    • C22B11/04Obtaining noble metals by wet processes
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/06Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
    • C22B3/08Sulfuric acid, other sulfurated acids or salts thereof
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/20Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/22Treatment or purification of solutions, e.g. obtained by leaching by physical processes, e.g. by filtration, by magnetic means, or by thermal decomposition
    • C22B3/24Treatment or purification of solutions, e.g. obtained by leaching by physical processes, e.g. by filtration, by magnetic means, or by thermal decomposition by adsorption on solid substances, e.g. by extraction with solid resins
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/20Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/42Treatment or purification of solutions, e.g. obtained by leaching by ion-exchange extraction
    • YGENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02PCLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00Technologies related to metal processing
    • Y02P10/20Recycling

Abstract

Processes are provided for recovering precious metals from refractory materials using thiosulfate lixiviants and an oxygen-containing gas at or above atmospheric pressure. The processes can employ lixiviants that include at mo st only small amounts of copper and/or ammonia to control the degradation of the thiosulfate lixiviant. The process can also operate at a relatively low pH, in relativel y inert atmospheres to control the production of polythionates which can complicate the recovery of the precious metals. The processes of the present invention provide relatively high rates of precious metal recovery.

Description

METHOD FOR THIOSULFATE LEACHING OF
PRECIOUS METAL-CONTAINING MATERIALS
FIELD OF THE INVENTION
The present invention is directed generally to the recovery of precious metals from precious metal-containing material and specifically to the recovery of precious metals from precious metal-containing material using thiosulfate lixiviants.

BACKGROUND OF THE INVENTION
A traditional technique for recovering precious metal(s) from precious metal-containing ore is by leaching the material with a cyanide lixiviant. As used herein, a "precious metal" refers to gold, silver, and the platinuin group metals (e.g., platinum, palladium, ruthenium, rhodium, osmium, and iridium). Many countries are placing severe limitations on the use of cyanide due to the deleterious effects of cyanide on the environment. Incidents of fish and other wildlife having been killed by the leakage of cyanide into waterways have been reported. The limitations being placed on cyanide use have increased substantially the cost of extracting precious metal(s) from ore, thereby decreasing precious metal reserves in many countries. Cyanide is also unable to recover precious metals such as gold from refractory ores without a pretreatment step.
"Refractory ores" refer to those ores that do not respond well to conventional cyanide leaching. Examples of refractory ores include sulfidic ores (where at least some of the precious metals are locked up in the sulfide matrix), carbonaceous ores (where the precious metal complex dissolved in the lixiviant adsorbs onto carbonaceous matter in the ores), and sulfidic and carbonaceous ores.
Thiosulfate has been actively considered as a replacement for cyanide.
Thiosulfate is relatively inexpensive and is far less harmful to the environment than cyanide. Thiosulfate has also been shown to be effective in recovering precious metals from pretreated refractory preg-robbing carbonaceous ores and sulfidic ores.
As used herein, "preg-robbing" is any material that interacts with (e.g., adsorbs or binds) precious metals after dissolution by a lixiviant, thereby interfering with precious metal extraction, and "carbonaceous material" is any material that includes one or more carbon-containing compounds, such as humic acid, graphite, bitumins and asphaltic compounds.
Where gold is the precious metal, thiosulfate leaching techniques have typically relied on the use of copper ions to catalyze and accelerate the oxidation of gold, ammonia to facilitate the formation and stabilization of cupric ammine ions and/or a pH
at pH 9 or above to maintain a region of stability where both the cupric ammine and gold thiosulfate complexes are stable.
It is well known in the art that the catalytic effect of copper and ammonia in conventional thiosulfate leaching of gold is described by the following sequence of reactions.
Formation of the cupric ammine complex:

CuZ' +4NH --> Cu(NH,);' (1) Oxidation of gold by cupric ammine, gold complexation as the gold-thiosulfate anion, and reduction of the cupric aminine to cuprous thiosulfate:

Au+ Cu(NH3)4+ + 5S2O3 -> Au(SaO3)2 + Cu(S2O3)3 + 4NH3 (2) Oxidation of the cuprous thiosulfate back to cupric ammine with oxygen:

Cu(S203)3 +4NH3 + 1402 + 12H2O-> eu(NH3)4+ +3S2O3 +OH (3) Summing equations (2) and (3) yields the overall thiosulfate leach reaction for gold:
Au+ 2S203 +/02 + Hz0~ Au(S2O3)2 + OH- (4) It can be seen from the above equations that copper and ammonia act as catalysts in that they are neither produced nor consumed in the overall leach reaction.
Cupper and ammonia can be a source of problems. Added copper tends to precipitate as cupric sulfide, which is speculated to form a passive layer on gold, thereby inhibiting gold leaching as well as increasing copper and thiosulfate consumption:

CZl2+S203 +20H -> CuS+SO4 +HZO (5) Rapid oxidation of thiosulfate by cupric ammine also occurs, leading to excessive degradation and loss of thiosulfate:

2Cu(NH3)4+ + gS203 -> 2CulS2O3)3 + 5406 + 8NH3 (6) Loss of ammonia by volatilization occurs readily, particularly in unsealed gas-sparged reactors operating at pH greater than 9.2, leading to excessive annnonia consumption:

NH4 +OH---~ NH3(aq)+H20-~ NH3(g) +HZO (7) Like cyanide, copper and ammonia are highly toxic to many aquatic lifeforms and are environmentally controlled substances.
Other problems encountered with thiosulfate leaching include difficulty in recovering gold out of solution as a result of the formation of polythionates, such as tetrathionate and trithionate, which adsorb competitively with gold onto adsorbents, such as resins. The formation of polythionates further increases thiosulfate consumption per unit mass of processed ore.

SUMMARY OF THE INVENTION
An object of the present invention is to provide a method for thiosulfate leaching of precious metal-containing materials. In accordance with an aspect of the present invention, there is provided a process for recovering a precious metal from a precious metal-containing material, comprising: contacting the precious metal-containing material with a thiosulfate lixiviant at superatmospheric pressure in the absence of at least one of added copper and added ammonia to solubilize the precious metal and form a pregnant thiosulfate leach solution containing the solubilized precious metal; and thereafter recovering the solubilized precious metal from the pregnant thiosulfate leach solution.
In accordance with another aspect of the invention, there is provided a process for recovering a precious metal from a carbonaceous precious metal-containing material, comprising: contacting the carbonaceous precious metal-containing material with a thiosulfate lixiviant in the substantial or complete absence of added copper and added ammonia to solubilize the precious metal and form a pregnant thiosulfate leach solution containing the solubilized precious metal;
and thereafter recovering the solubilized precious metal from the pregnant thiosulfate leach solution.
In accordance with another aspect of the invention, there is provided a process for recovering a precious metal from a precious metal-containing material, comprising: (a) contacting the precious metal-containing material with a thiosulfate lixiviant to solubilize the precious metal and form a pregnant thiosulfate leach 3a solution containing the solubilized precious metal and a polythionate; (b) contacting the pregnant leach solution with a sulfide-containing reagent to precipitate at least most of the solubilized precious metal and convert at least most of the polythionate to thiosulfate; and (c) thereafter recovering the precious metal precipitate from the thiosulfate leach solution.
In accordance with another aspect of the invention, there is provided a process for recovering a precious metal from a precious metal-containing material, comprising: (a) solubilizing a first portion of the precious metal in the precious metal-containing material to form a first pregnant leach solution, wherein the solubilizing step (a) is conducted at a first oxygen partial pressure; (b) solubilizing a second portion of the precious metal in the precious metal-containing material to form a second pregnant leach solution, wherein the solubilizing step (b) is conducted at a second oxygen partial pressure and wherein the first oxygen partial pressure is less than the second oxygen partial pressure; (c) separating at least the second pregnant leach solution from the precious metal-containing material; and (d) recovering the solubilized precious metal from the first and second pregnant leach solution.
In accordance with another aspect of the invention, there is provided a process for recovering a precious metal from a precious metal-containing material, comprising: contacting the precious metal-containing material with a thiosulfate lixiviant at superatmospheric pressure and at a pH less than pH 9 to solubilize the precious metal and form a pregnant thiosulfate leach solution containing the solubilized precious metal; and recovering the solubilized precious metal from the pregnant thiosulfate leach solution.
In accordance with another aspect of the invention, there is provided a process for recovering a precious metal from a precious metal-containing material, comprising: contacting the precious metal-containing material with a thiosulfate lixiviant at superatmospheric pressure and at a temperature ranging from about 40 to about 100 C to solubilize at least most of the precious metal in the material and form a pregnant thiosulfate leach solution containing the solubilized precious metal; and recovering the solubilized precious metal from the pregnant thiosulfate leach solution.
In accordance with another aspect of the invention, there is provided a process for recovering a precious metal from a precious metal-containing material, comprising: (a) contacting the precious metal-containing material with a thiosulfate lixiviant to solubilize the precious metal and form a pregnant thiosulfate leach 3b solution containing the solubilized precious metal; (b) contacting the pregnant thiosulfate leach solution with an extraction agent at a temperature of more than about 60 C to recover the precious metal from the pregnant thiosulfate leach solution and convert trithionates in the pregnant thiosulfate leach solution into thiosulfate; and (c) recovering the precious metal from the extraction agent.
In accordance with another aspect of the invention, there is provided a process for recovering a precious metal from a precious metal-containing material, comprising: contacting the precious metal-containing material with a thiosulfate lixiviant to solubilize the precious metal and form a pregnant thiosulfate leach solution comprising solubilized precious metal, thiosulfate, and at least one of trithionate and tetrathionate; after the contacting step, converting at least most of the at least one of trithionate and tetrathionate in the pregnant thiosulfate leach solution into thiosulfate; and thereafter recovering the solubilized precious metal from the pregnant thiosulfate leach solution.
In accordance with another aspect of the invention, there is provided a process for recovering a precious metal from a precious metal-containing material, comprising: contacting the precious metal-containing material with a lixiviant to solubilize the precious metal and form a pregnant leach solution containing the solubilized precious metal; and thereafter electrowinning the precious metal in the presence of sulfite.
In accordance with another aspect of the invention, there is provided a process for recovering a precious metal from a precious metal-containing material, comprising: contacting the precious metal-containing material with a thiosulfate leach solution to solubilize the precious metal and form a pregnant thiosulfate leach solution containing solubilized precious metal; maintaining a dissolved molecular oxygen content in at least one of the thiosulfate leach solution and the pregnant thiosulfate leach solution at or below about I ppm to inhibit the formation of trithionate and tetrathionate; and recovering the solubilized precious metal from the pregnant thiosulfate leach solution.
In accordance with another aspect of the invention, there is provided a hydrometallurgical process for the recovery of precious metal values from a refractory precious metal ore material containing precious metal values and pregrobbing carbonaceous compounds, comprising: (a) providing a body of particles and/or particulates of the refractory precious metal ore material; (b) contacting the body of particles and/or particulates with a thiosulfate lixiviant solution at 3c superatmospheric pressure and at a pH of less than pH 9 to form stable precious metal thiosulfate complexes; (c) recovering the thiosulfate lixiviant solution from the body of particles and/or particulates after a period of time which is sufficient for the thiosulfate lixiviant solution to become pregnant with precious metal values extracted from the ore material; and (d) recovering the precious metal values from the lixiviant solution.
In accordance with another aspect of the invention, there is provided a process for recovering a precious metal from a precious metal-containing material, comprising: (a) contacting a precious metal-containing material with a thiosulfate lixiviant to dissolve the precious metal and form a pregnant thiosulfate leach solution containing the dissolved precious metal; (b) contacting the pregnant thiosulfate leach solution with an adsorbent to load the precious metal onto the adsorbent; (c) contacting the loaded adsorbent with an eluant other than sulfite in the presence of sulfite to desorb the precious metal adsorbed on the loaded adsorbent and form a loaded eluate containing the dissolved precious metal, and (d) recovering the precious metal from the loaded eluate.
In accordance with another aspect of the invention, there is provided a process for recovering a precious metal for precious metal-containing material comprising: (a) contacting a precious metal-containing material with a thiosulfate lixiviant to dissolve the precious metal and form a pregnant thiosulfate leach solution containing the dissolved precious metal; (b) contacting the pregnant thiosulfate leach solution and/or a barren thiosulfate leach solution with a sulfide and/or bisulfide and/or a polysulfide to convert polythionates in the pregnant thiosulfate leach solution and/or barren thiosulfate leach solution into thiosulfate; and (c) thereafter contacting the pregnant thiosulfate leach solution and/or barren thiosulfate leach solution with an oxidant to solubilize precipitated precious metal precipitates.
In accordance with another aspect of the invention, there is provided a process for recovering a precious metal from a precious metal-containing material comprising: contacting the precious metal-containing material with a thiosulfate lixiviant at superatmospheric pressure to dissolve the precious metal and form a pregnant thiosulfate leach solution containing the dissolved precious metal, wherein the concentration of added sulfite during the contacting step is no more than about 0.01 M; and recovering the solubilized precious metal from the pregnant thiosulfate leach solution.
In accordance with another aspect of the invention, there is provided a 3d process for recovering a precious metal from a precious metal-containing material, comprising: (a) contacting the precious metal-containing material with a thiosulfate lixiviant to solubilize the precious metal and form a pregnant thiosulfate leach solution containing the solubilized precious metal, wherein the pregnant thiosulfate leach solution includes polythionates; (b) contacting the pregnant thiosulfate leach solution with a reluctant to convert the polythionate into thiosulfate; and (c) thereafter recovering the solubilize precious metal from the pregnant thiosulfate leach solution.
In accordance with another aspect of the invention, there is provided a process for recovering a precious metal from a precious metal-containing material, comprising: leaching the precious metal from the material with a thiosulfate lixiviant to form a pregnant leach solution including at least most of the precious metal in the material and a metal impurity; recovering the precious metal from the pregnant leach solution to form a barren leach solution; and contacting at least one of the pregnant leach solution and the barren leach solution with a reductant to reduce a concentration of the metal impurity, thereby inhibiting a reaction between the thiosulfate and the metal impurity.

These and other needs have been addressed by the methodologies and systems of the present invention. The methodologies can recover precious metals from a variety of materials, including refractory carbonaceous or sulfidic ores, double refractory ores (e.g., ores containing both sulfide-locked gold and carbonaceous preg-robbing matter), oxide ores, nonrefractory sulfidic ores, and ores also containing copper minerals and other materials derived from such ores (e.g., concentrates, tailings, etc.).
In one embodiment, a thiosulfate leaching process is provided that includes one or more of the following operating parameters:
(a) an oxygen partial pressure that is preferably superatmospheric and more preferably ranges from about 4 to about 500 psia;
(b) a leach slurry pH that is preferably less than pH 9;
(c) a leach slurry that is preferably at least substantially free of (added) ammonia and more preferably contains less than 0.05M (added) ammonia such that the leach slurry has a maximum total concentration of ammonia of preferably less than 0.05M and more preferably no more than about 0.025M;
(d) a leach slurry that is preferably at least substantially free of (added) copper ion and more preferably contains no more than about 15 ppm (added) copper ions;
(e) an (added) sulfite concentration that is preferably no more than about 0.01M such that the slurry has a maximum total concentration of sulfite of preferably no more than about 0.02M and more preferably no more than about 0.O1M; and/or (f) a leach slurry temperature preferably ranging from about 20 to about 100 C and more preferably from about 20 to about 80 C.
The foregoing parameters can yield a high level of precious metal extraction from the precious metal-containing material, which can be at least about 70% and sometimes at least about 80%.
The thiosulfate lixiviant can be derived from any suitable form(s) of thiosulfate, such as sodium thiosulfate, calcium thiosulfate, potassium thiosulfate and/or ammonium thiosulfate. Sodium and/or calcium thiosulfate are preferred.
The leaching process can be conducted by any suitable technique. For example, the leaching can be conducted in situ, in a heap or in an open or sealed vessel. It is particularly preferred that the leaching be conducted in an agitated, multi-compartment reactor such as an autoclave.
The precious metal can be recovered from the pregnant leach solution by any suitable technique. By way of exainple, the precious metal can be recovered by resin adsorbtion methods such as resin-in-pulp, resin-in-solution, and resin-in-leach or by solvent extraction, cementation, electrolysis, precipitation, and/or combinations of two or more of these techniques.
Reducing or eliminating the need to have copper ions and/or ammonia present in the leach as practiced in the present invention can provide significant multiple benefits. First, the cost of having to add copper and ammonia reagents to the process can be reduced significantly or eliminated. Second, environmental concerns relating to the presence of potentially harmful amounts of copper and ammonia in the tailings or other waste streams generated by the process can be mitigated. Third, the near-absence or complete absence of copper and ammonia in the leach can provide for a much more reliable and robust leaching process, yielding more stable leachates, able to operate over a wider pH and oxidation-reduction potential (ORP) range than is possible with conventional thiosulfate leaching. The latter process must operate in the relatively narrow window of pH and ORP where both the cupric ammine complex and the gold thiosulfate complex co-exist. With the process of the present invention, the pH of the thiosulfate lixiviant solution in the leaching step can be less than pH 9 and the ORP less than 200 mV (referenced to the standard hydrogen electrode). Fourth, minimizing the amount of copper in the system can lead to increased loading of gold onto resins due to 5 reduced competitive adsorption of copper ions. Resin elutions are also simplified as little, if any copper, is on the resin. Finally, the near-absence or complete absence of copper and ammonia in the leach can reduce or eliminate entirely a host of deleterious side reactions that consume thiosulfate and are otherwise difficult or impossible to prevent.
The elimination or near elimination of sulfite from the thiosulfate leach also can have advantages. Sulfite can depress the rate of dissolution of precious metal from the precious metal-containing material by reducing significantly the oxidation reduction potential (ORP) of the leach solution or lixiviant. As will be appreciated, the rate of oxidation of the gold (and therefore the rate of dissolution of the gold) is directly dependent on the ORP.
In another embodiment, an extraction agent is preferably contacted with a pregnant (precious metal-containing) thiosulfate leach solution at a temperature of less than about 70 C and more preferably less than about 60 C in the substantial absence of dissolved molecular oxygen to isolate the precious metal and convert polythionates in the pregnant leach solution into thiosulfate. In one configuration, the extraction agent is an adsorbent, such as a resin, which loads the precious metal onto the adsorbent. As used herein, an "adsorbent" is a substance which has the ability to hold molecules or atoms of other substances on its surface. Examples of suitable resin adsorbents include weak and strong base resins such as "DOWEX 21K", manufactured by Dow Chemical.
In another configuration, the extraction agent is a solvent extraction reagent that extracts the precious metals into an organic phase, from which the precious metals can be later recovered. As will be appreciated, the detrimental polythionates decompose into thiosulfate in the substantial absence of dissolved molecular oxygen.
In yet another embodiment, the pregnant leach solution from a thiosulfate leaching step is contacted, after the leaching step, with a reagent to convert at least about 50% and typically at least most of polythionates (particularly trithionate and tetrathionate) into thiosulfate. The reagent or reductant can be any suitable reactant to convert polythionates into thiosulfate, with any sulfide, and/or polysulfide (i.e., a compound containing one or a mixture of polymeric ion(s) Sx 2", where x = 2-6, such as disulfide, trisulfide, tetrasulfide, pentasulfide and hexasulfide) being particularly preferred. A sulfite reagent can also be used but is generally effective only in converting polythionates of the form SX062", where x = 4 to 6, to thiosulfate. The sulfite, sulfide, and/or polysulfide can be compounded with any cation, with Groups IA and IIA
elements of the Periodic Table, ammonium, and hydrogen being preferred.
In yet another embodiment, a precious metal solubilized in a solution, such as a pregnant leach solution or eluate, is electrowon in the presence of sulfite.
In the presence of sulfite, the precious metal is reduced to the elemental state at the cathode while the sulfite is oxidized to sulfate at the anode. Sulfite is also believed to improve the precious metal loading capacity of the resin by converting loaded tetrathionate to trithionate and thiosulfate.
In yet another embodiment, the formation of polythionates is controlled by maintaining a (pregnant or barren) thiosulfate leach solution in a nonoxidizing (or at least substantially nonoxidizing) atmosphere and/or sparging a nonoxidizing (or at least substantially nonoxidizing) gas tlirough the leach solution. As will be appreciated, the atinosphere or gas may contain one or more reductants, such as hydrogen sulfide and/or sulfur dioxide. The molecular oxygen concentration in the atmosphere and/or sparge gas is preferably insufficient to cause a dissolved molecular oxygen concentration in the leach solution of more than about 1 ppm and preferably of more than about 0.2 ppm.
Preferably, the inert atmosphere (or sparge gas) is at least substantially free of molecular oxygen and includes at least about 85 vol. % of any inert gas such as molecular nitrogen and/or argon. By controlling the amount of oxidant(s) (other than thiosulfate and polythionates) in the atmosphere and/or (pregnant or barren) leach solution the rate or degree of oxidation of thiosulfates to form polythionates can be controlled.

BRIEF DESCRIPTION OF THE DRAWINGS
Fig. 1 is a flow schematic of a first embodiment of the present invention;
Fig. 2 is a flow schematic of second embodiment of the present invention;
Fig. 3 is a flow schematic of a third embodiment of the present invention;
Fig. 4 is a flow schematic of a fourth embodiment of the present invention;
Fig. 5 is a plot of gold extraction in percent (vertical axis) versus leach time in hours (horizontal axis);
Fig. 6 is another plot of gold extraction in percent (vertical axis) versus leach time in hours (horizontal axis);
Fig. 7 is another plot of gold extraction in percent (vertical axis) versus leach tiine in hours (horizontal axis);
Fig. 8 is another plot of gold extraction in percent (vertical axis) versus leach time in hours (horizontal axis); and Fig. 9 is a plot of gold extraction in percent (left vertical axis) and thiosulfate remaining in percent (right vertical axis) versus leach time in hours (horizontal axis).
DETAILED DESCRIPTION
The present invention provides an improved thiosulfate leaching process for the recovery ofprecious metals from precious metal-bearing material. The precious metal(s) can be associated with nonprecious metals, such as base metals, e.g., copper, nickel, and cobalt. The precious metal-bearing material includes ore, concentrates, tailings, recycled industrial matter, spoil, or waste and mixtures thereof. The invention is particularly effective for recovering precious metals, particularly gold, from refractory carbonaceous material.
Figure 1 is a flow chart according to a first embodiment of the present invention.
The process of the flow chart is particularly effective in recovering gold from carbonaceous material and oxide material and mixtures thereof.
Referring to Figure 1, a precious metal-bearing material 100 is subjected to the steps of wet and/or dry crushing 104 and wet and/or dry grinding 108 to reduce the particle size of the material sufficiently to enable the solids to be suspended in an agitated vessel and to allow for the efficient leaching of the precious metals. Preferably, wet grinding is employed with the recycled thiosulfate leach solution and water being used as the liquid component in the slurry. In that event, the slurry 112 containing the comminuted material typically contains from about 0.05 to about 0.1 M
thiosulfates and from about 0.0005 to about 0.025 M polythionates. The fully comminuted material particle size is preferably at least smaller than 80% passing about 48 mesh (300 microns), more preferably 80% passing about 100 mesh (150 microns), and most preferably 80%
passing about 200 mesh (75 microns). The typical solids content of the slurry 112 ranges from about 20 to about 30 wt.%. As will be appreciated, other techniques can be used to comminute the material to the desired particle size(s). By way of illustration, blasting can be used alone with or without crushing and grinding and crushing and grinding can be used alone with or without another comminution technique.
The ground slurry 112 is then thickened 116 to adjust the pulp density to a value suitable for leaching. The ideal leach pulp density will vary according to the type of material being leached. Typically, the pulp density ranges from about 20 to about 50%
solids by weight, but could be as low as about 1% or as high as about 60%.
Thickening 116 will generally not be required if the desired pulp density (after wet comminution or forination of the comminuted material into a slurry) is less than about 20%.
The thickener overflow solution 120 is recycled back to grinding 108 in the event that wet grinding is employed. Otherwise, the overflow solution 120 is returned to the optional slurry formation step (not shown).
Fresh makeup thiosulfate is added, as necessary, at any suitable location(s), such as to the slurried material during comminution 108 and/or in the thickener 116, to the underflow or overflow solution 124, 120, to leaching 132 and/or to the regenerated thiosulfate solution 128 (discussed below). In any event, the optimum solution thiosulfate concentration to maintain during leaching 132 will depend on the nature of the material being leached, but will preferably range from about 0.005 to about 2 molar (M), more preferably about 0.02 to about 0.5 M, and even more preferably from about 0.05 to about 0.2 M. The source of makeup thiosulfate can be any available thiosulfate-containing compound, such as sodium thiosulfate, potassium thiosulfate, calcium thiosulfate, or any other thiosulfate-containing material or thiosulfate precursor.
Ammonium thiosulfate can also be used but its use is less preferred for environmental reasons. Alternatively, thiosulfate can be generated in situ or in a separate step by reaction of elemental sulfur with a source of hydroxyl ions, in accordance with the following reaction:
2(x + 1)S + 60H- -~ S2O3 + 2Sx + 3H20 (8) where x= 3-6, or by reaction of bisulfide with bisulfite:

2HS- + 4HS03 -> 3S2O3 + 3H20 (9) or by reaction of elemental sulfur with sulfite:

S + S03 -> S2 03 (10) If the desirable temperature is above ainbient, it may be desirable to recover waste heat for recycle to leaching. In that event, the underflow slurry 124 is directed through an indirect heat exchanger 136, preferably a shell and tube heat exchanger system in which the hot slurry from resin-in-pulp pretreatment 140 (discussed below) is passed through the inner tubes and the cold feed (or underflow) slurry 140 is passed through the annular space between the tubes (or vice versa). In this way waste heat is transferred from the leached slurry 144 to the feed (or underflow) slurry 124, reducing the amount of new heat that must be added in leaching 132 to maintain the desired leach temperature. Typically, the approach temperature of the incoming feed slurry 148 is from about 2 to about 5 C below the leach temperature (discussed below) and heat is added to the leach vessel by suitable techniques to makeup the difference.
The heated slurry 148 is subjected to leaching 132 in the presence of oxygen and thiosulfate. Leaching is conducted in the presence of an oxygen-enriched atmosphere at atmospheric pressure, or at a pressure above atmospheric pressure using an oxygen-containing gas to reduce or eliminate the need for the presence of copper and/or ammonia in the leach. Using gold as an example, the thiosulfate leaching of precious metal-bearing material in the absence or substantial absence of copper and ainmonia under elevated oxygen partial pressure can be illustrated by the following reaction:

Au+ 2S2O 3 + 4 02 +'2H20 __> Au(S2O3)2 + OH (11) The increased oxygen partial pressure in the leach increases the rate of the above reaction in the absence or near absence of copper and ammonia. To accomplish this goal, the oxygen-containing gas may include atmospheric air, or it may include relatively pure (95%+) oxygen such as that produced from any commercially available oxygen plant, or it may include any other available source of oxygen. The desired oxygen partial pressure (POz) maintained during leaching will depend on the material being leached, but it will be at least higher than that provided under normal ambient conditions by air at the 5 elevation the process is applied. Thus, if the process is practiced at sea level for example the oxygen partial pressure will be in excess of about 3 pounds per square inch absolute pressure (psia) to as high as about 500 psia, preferably from about 10 to about 115 psia, and most preferably from about 15 to about 65 psia. The total operating pressure is the sum of the molecular oxygen partial pressure and the water vapor pressure at the 10 temperature employed in the leaching step 132, or preferably ranges from about 15 to about 600 psia and more preferably from about 15 to about 130 psia.
The leaching temperature will be dictated by the type of material being leached.
The temperature will vary typically from about 5 C to about 150 C, preferably from about 20 to about 100 C, and most preferably from about 40 to about 80 C.
Higher temperatures accelerate the leaching of precious metals but also accelerate the degradation of thiosulfate. If required, a source of makeup heat such as steam is added to the leach reactors to maintain the desired temperature.
The leaching retention time is dependent on the material being leached and the temperature, and will range from about 1 hour to 96 hours, preferably from about 2 to about 16 hours, and most preferably from about 4 to about 8 hours.
The absence or substantial absence of copper and/or ammonia in the leach greatly simplifies the process. Elimination or near-elimination of ammonia and copper from the leach provides the advantage of allowing for a consistently high and reproducible precious metal extraction over a broader pH range than was previously possible with the other thiosulfate leaching processes. Preferably, the (added and/or total solution) copper concentration is no more than about 20 ppm, more preferably no more than about ppm, and even more preferably no more than about 10 ppm while the (added and/or total solution) ammonia concentration is no more than about 0.05 M, more preferably no more than about 0.03 M, and even more preferably no more than about 0.01 M. In the present invention leaching can be operated at about pH 7-12, preferably about pH 8-11, more preferably about pH 8-10, and even more preferably at a pH less than pH 9. The oxidation-reduction potential (ORP) preferably ranges from about 100 to about 350 mV
and more preferably from about 150 to about 300 mV (vs. the standard hydrogen electrode (SHE)).
Oxidative degradation ofthiosulfate ultimately to sulfate can also occur, possibly by the following sequence of reactions that involve the formation of intermediate polythionates (polythionates can be represented by SõO6 , where n = 2-6):

Tetrathionate formation: 2S203 +/ O2 + H20 -4 S406 - + 2OH- (12) Trithionate formation: 3S4O6 + 2 02 + H20 -4 4S3O6 + 2H} (13) Sulfite formation: S306 +/ 02 + 2H20 -4 3S03 + 4H+ (14) Sulfate formation: 2SO3 + 02 -4 2SO4 (15) Overall: S2 03 + 202 + HZ O--~ 2SO4 + 2H+ (16) Oxidative degradation of thiosulfate to polythionates and sulfates is accelerated markedly in the presence of copper ions and/or ammonia. The oxidative degradation reactions are slowed considerably at elevated oxygen partial pressure in the absence or near-absence of copper and ammonia.
The leaching step 132 may be conducted in a batch or continuous basis but continuous operation is preferred. Continuous leaching is carried out in a multiple series of one or more reactors that are agitated sufficiently to maintain the solids in suspension.
Agitation may be accomplished by mechanical, pneumatic or other means. In a preferred configuration, gassing impellers are employed, such as those disclosed inU.S.
PatentNo.
6,183,706.
Such impellers can significantly enhance the amount of dissolved molecular oxygen in the leach solution.
Leaching may also be carried out in a multi-compartment autoclave containing one or more compartments, (with 4 to 6 compartments being preferred) similar in design to the autoclaves used to pressure oxidize sulfide-bearing ores or concentrates.
However, owing to the non-acidic, moderate temperature, relatively mild conditions employed in the present invention, the autoclave materials of construction are much less expensive than those found to be necessary when oxidizing sulfide minerals. The latter autoclaves are normally constructed of a steel shell fitted with a lead liner and refractory brick liner and containing metallic components constructed of titanium or other expensive corrosion-resistant alloys. The leach reactors and contained metallic components employed by the present invention can be simply constructed of stainless steel and do not require lead or brick liners.

The extraction of precious metals in the leaching step 132 is relatively high, particularly for carbonaceous ores. Typically, at least about 50%, more typically at least about 70%, and even more typically at least about 80% of the precious metal in the precious metal-containing niaterial is extracted or solubilized into the pregnant leach solution 144. The concentration of the dissolved precious metal in the pregnant leach solution typically ranges from about 0.05 to about 100 ppm and more typically from about 1 to about 50 ppm.

The pregnant leach slurry 144 containing the precious metal-bearing leach solution and gold-depleted solid residue may optionally be directed to RIP
pretreatment 140 to reduce the concentration ofpolythionates in solution. As will be appreciated, the molecular oxygen sparged through the leach slurry in the leaching step 132 will oxidize a minor portion of the thiosulfate into polythionates. Polythionates have the undesired effect of reducing the loading of precious metals on to resin by competitive adsorption.
Lowering the polythionate concentration will have the beneficial effect of increasing the loading of precious metals on to resin, thereby improving the efficiency of resin recovery of precious metals.

The RIP pretreatment step 140 can be performed using any one or more of a number of techniques for converting polythionates to other compounds that do not compete with the precious metal for collection by the extraction agent.
In one embodiment, a polythionate reductant is added to the slurry 144 to reduce polythionates to thiosulfates. Any of a number of reductants are suitable for performing the conversion.
By way of example, a sulfide-containing reagent can reduce the polythionates back to thiosulfate, as shown by the following reactions:

2S4O6 + S2 + 32 HZO ~ 2S2 O3 + 3H+ (17) S3O6 + SZ ~ 2S2O3 (18) Any reagent that releases sulfide ions on dissolution will suffice, such as sodium bisulfide, NaHS, sodium sulfide, Na2S, hydrogen sulfide gas, H2S, or a polysulfide. The use of a sulfide reagent has the advantages of rapidly and efficiently reducing polythionates to thiosulfate at ambient or moderately elevated temperature.
The treatment can be carried out in an agitated reactor in batch mode or in a series of 1-4 reactors operating in continuous mode, or in a multi-compartment autoclave.
Alternatively the treatment can be carried out in a pipe reactor or simply by injecting sulfide ions in the piping system directing the leach slurry to gold recovery, or the first stage of RIP. The treatment is carried out at a controlled pH of about pH 5.5 to about pH
10.5, preferably about pH 7 to about pH 10, most preferably less than about pH
9. The temperature employed can range from about 20 C to about 150 C, preferably from about 40 to about 100 C, more preferably from about 40 to about 80 C, and even more preferably from about 60 to about 80 C. The retention time can range from as low as 5 minutes, preferably greater than 30 minutes, most preferably from about 1 to about 3 hours.
Alternatively, a sulfite-containing reagent can also reduce polythionates to thiosulfates as shown by the following reaction:

S4O6 + SO3 -~ S2O3 + S3O6 (19) Sulfite treatment is effective in reducing tetrathionate quickly, but a disadvantage is it is ineffective in reducing tritliionate. The sulfite can be added in any fornn and/or can be formed in situ. For example, sulfite can be added in the form of sodium metabisulfite or sulfur dioxide. , When using sulfite, the temperature of the leach slurry in the RIP
pretreatment 140 is preferably less than 60 C to inhibit, at least substantially, the precipitation of precious metal(s) from the leach slurry 144. More preferably, the RIP
pretreatment 140 with sulfite is performed at a temperature in the range of from about 10 to about 50 C
and even more preferably at ambient temperature.

When using sulfite, the residence time of the leach slurry 144 in the regeneration step 140 is preferably at least about 1 second, more preferably greater than about 5 minutes, and even more preferably greater than about 10 minutes and no more than about 1 hour, with about 15-30 minutes being most preferable.

The pH of the leach slurry during sulfite treatment typically ranges from about pH 5.5 to about pH 10.5 and more typically from about pH 7 to about pH 10.

Other suitable polythionate reductants include hydrogen, fine, reactive elemental sulfur, carbon monoxide, and mixtures thereof.

In another embodiment, the pretreatment step 140 is performed by maintaining the tenlperature of the leach slurry at a sufficiently high value in the absence of oxygen to effect the following hydrolytic disproportionation reactions:

4S40g + 5H20 -4 7S2O3 + 2SO4 + 10H+ (20) 5306 + H2O -> S2O3 + SO4 + 2H+ (21) Hydrolytic treatment can be carried out in an agitated reactor in batch mode or in a series of 1-4 reactors operating in continuous mode, or a multi-compartment autoclave. The temperature is preferably maintained in the range of from about 60 to about 150 C, preferably of from about 70 to about 100 C, and most preferably of from about 80 to about 90 C by adding a source of heat, such as steam. The retention time typically ranges from about 15 minutes to 8 hours, preferably from about 1 to about 6 hours, and most preferably from about 2 to about 4 hours. Hydrolytic treatment is generally less preferable than sulfide treatment because the former method results in irreversible loss of some of the polythionate to sulfate.

Alternatively, any or all of the above-techniques for converting polythionate(s) into thiosulfate can be combined in the same process configuration.

In a preferred embodiment, the reductant used to convert polythionates into thiosulfates is the sulfide reagent. Sulfide addition is preferred because one sulfide reacts with one tri- or two tetrathionates to form multiple thiosulfates without any sulfur-containing byproducts. Sulfite addition only reduces tetrathionate and is not capable of reducing trithionate at common operating temperatures and pH's. Heating of the leach solution is energy intensive and produces byproducts. Trithionate and tetrathionate are 5 each converted into thiosulfate, sulfate, and hydrogen ions, thus the thiosulfate yield is not as high as with sulfide addition.

RIP pretreatment 140 can be performed in any suitable vessel(s), preferably agitated. Preferably, RIP pretreatment is performed in a series of tanks or in a multistaged vessel.
10 The addition of a sulfide such as NaHS is preferred. Preferably, the amount of the reductant generally and, sulfide reagent specifically added to the slurry 144 is sufficient to convert at least most of the polythionates into thiosulfate. The amount of sulfide contacted with the slurry 144 preferably is at least about 100 to about 150% of the stoichiometric amount required to convert at least substantially all of the 15 polythionates in the slurry into thiosulfates. Typically, at least about 50%, more typically at least most, and even more typically from about 80 to about 95% of the polythionates are converted into thiosulfates in RIP pretreatment 140.

The temperature of the slurry 144 preferably is at least about 60 C and the ORP
of the exiting slurry 152 is at least below about 100 mV (SHE) and more preferably ranges from about -100 to about 100 mV (SHE) to substantially minimize precious metal precipitation.

The exiting RIP pretreated slurry 152 is passed through heat exchanger 136 and conditioned in a conditioner 156 to resolubilize any precious metal precipitated during RIP pretreatment 140 and/or heat exchange 136. Conditioning 156 is performed in an agitated single- or multi-compartment vessel which has an oxidizing atmosphere, such as air, to cause solubilization of the precious metal precipitates. Although polythionates will form in the presence of an oxidant, such as molecular oxygen, the rate of conversion of thiosulfate to polythionates is much slower than the rate of precious metal solubilization. Preferably, the residence time (at ambient temperature and pressure) is selected such that at least about 95 % of the precious metal precipitates are solubilized while no more than about 5 % of the thiosulfate is converted into polythionates.
Preferably, the slurry residence time in conditioning 156 is no more than about 12 hrs and more preferably ranges from about 1 to about 6 hrs.
The conditioned slurry 160 is next subjected to resin-in-pulp treatment 164 to extract the precious metal from the conditioned slurry 160. The resin-in-pulp step 164 can be performed by any suitable technique with any suitable ion exchange resin.
Examples of suitable techniques include that discussed in U. S. Patents 5,536,297 and 5,785,736. Preferred resins include anion exchange resins, preferably a strong base resin including a quatemary amine attached to a polymer backbone. A
strong base resin is preferred over a weak base resin. The precious metal loading capacity of a strong base resin is typically greater than that of a weak base resin, such that a lower volume of resin is required. Gel resins and macroporous resins are suitable.
Suitable resins include all commercial strong-base resins of either Type I
(triethylamine functional groups) or Type II (triethyl ethanolamine functional groups).
Specific strong-base ion exchange resins include "A500" manufactured by Purolite"`, "A600"
manufactured by Purolite"', "21K" manufactured by Dow Chemical, "AmberliteT``
IRA
410" manufactured by Rohm and Haas, "Amberlite"`' IRA 900" manufactured by Rohm and Haas, and "Vitrokele'T" 911" supplied by Signet. Because the RIP
pretreatment and resin-in-pulp steps 140 and 164 are preferably performed in the same vessel (though they may be performed in different vessels), the temperature, leach slurry pH, and residence time typically depend on which of the above techniques are used to reduce the polythionate concentration.
Resin-in-pulp treatment can be performed in any suitable vessel. A preferred vessel is a Pachuca tank, which is an air-agitated, conical bottomed vessel, with air being injected at the bottom of the cone. An advantage of the Pachuca system is reduced resin bead breakage and improved dispersion of the resin beads in the slurry as compared to mechanically agitated systems. The RIP recovery is preferably carried out in four or more tanks connected in series, more preferably between four and eight such Pachuca tanks.
During resin-in-pulp 164, the resin will become "loaded" with the dissolved precious metals. Typically, at least about 99% and more typically at least about 99.8%
of the precious metal(s) in the leach slurry will be "loaded" or adsorbed onto the resin.
To inhibit the formation of polythionates and the consequent precious metal recovery problems and increased reagent consumption, the leach slurry can be maintained in an inert (or an at least substantially nonoxidizing) atmosphere and/or an inert (or an at least substantially nonoxidizing) gas can be sparged through the leach slurry. The atmosphere is preferably maintained (and/or gas sparging used) during RIP
pretreatment 140 and resin-in-pulp 164. As used herein, "inert" refers to any gas which is at least substantially free of oxidants, such as molecular oxygen, that can cause thiosulfate to be converted into a polythionate. For example, an "inert" gas would include a reducing gas. Typically, the inert atmosphere will include at least about 85 vol % of an inert gas, preferably nitrogen gas, and no more than about 5 vol %
oxidants, such as oxygen gas, that can cause thiosulfate conversion into a polythionate. The molecular nitrogen can be a byproduct of the oxygen plant that is employed in the leaching step to provide superatmospheric partial pressures of oxygen gas. As will be appreciated, the leach slurry 144 during transportation between the leaching and RIP
pretreatment steps 132 and 140 and if applicable from the RIP pretreatment and resin-in-pulp steps 140 and 164 (except during conditioning 156) is typically in a conduit that is not open to the surrounding atinosphere. If the leach slurry is open to an atmosphere during transportation in either or both of these stages, the leach slurry should be maintained in the presence of the inert atmosphere during any such transportation.

While not wishing to be bound, it is believed that sparging is more effective than an inert atmosphere without sparging in controlling polythionate production.
Sparging appears to inhibit molecular oxygen ingress into the solution, even where the reactor is open to the ambient atmosphere, because of the outflow of inert gas from the surface of the solution.

The barren leach slurry 168 (which will typically contain no more than about 0.01 ppm precious metals or 1% of the precious metal(s) in the leach solution 144) is subjected to one or more stages of counter current decantation ("CCD") 172. In CCD
172, the solids are separated in the underflow 176 from the barren leach (or overflow) solution 180 and sent to the tailings pond. The barren leach solution 180 is separated in the overflow from the solids and forwarded to regeneration step 184 to convert polythionates to thiosulfate. As will be appreciated, CCD performs liquidlsolid separation, provides water balancing in the circuit, and prevents build up of impurities in the leach circuit by removing a portion of the leach solution with the solids.
Regeneration 184 can be performed in one or more vessel(s) and/or by in line sulfide (and/or sulfite) addition to a conduit carrying the stripped lixiviant solution. If a number of the techniques are employed, they can be performed simultaneously (in the same reactors) or sequentially (in different reactors), as desired.
The regenerated lixiviant solution 128 is recycled to the grinding step 108 along with the thickener overflow 120 and ultimately to the leaching step 132.
The loaded resin 188 is screened 190 and washed with water to remove any leach slurry (liquid and/or leached material) from the resin beads.
The recovered beads 192 are contacted with an eluant to strip or elute 194 adsorbed precious metal into the eluate and form a pregnant solution 196 containing typically at least most (and more typically at least about 95%) of the precious metal on the resin and a stripped resin 197.
The eluant can be any suitable eluant that can displace the adsorbed precious metal from the loaded resin beads. The eluant could include salts, such as one or more types of polythionate ions, and a nitrate, a thiocyanate, a sulfite, a thiourea, a perchlorate and mixtures thereof.
Typically, the concentration of the eluant in the pregnant solution 196 ranges from about 0.25 to about 3 M; the temperature of elution 194 from about 5 to about 70 C, and the pH of elution 194 from about pH 5 to about pH 12. Under the conditions, at least about 90% and more typically from about 95 to about 99% of the precious metal adsorbed on the resin is displaced by the eluant into the pregnant solution 196.
The stripped resin 197 is recycled to the resin-in-pulp step 164. Optionally, the stripped resin 197 can be regenerated (not shown) by known techniques prior to reuse of the resin. As will be appreciated, the resin can be regenerated by acid washing the resin with an acid such as nitric acid or hydrochloric acid. The acid wash removes adsorbed eluant and/or impurities from the resin and frees up the functional sites on the resin surface (previously occupied by the eluant) to adsorb additional precious metal.
In the case of a polythionate eluant, the resin can be regenerated by contacting the resin with sulfide and/or sulfite to reduce the polythionate ions to thiosulfate ions and sulfate ions. After regeneration, the resin and regeneration product solution are separated by screening and washing.
The pregnant solution 196, which includes the eluant and typically no more than about 100 ppm and more typically from about 10 to about 500 ppm solubilized precious metals, is subjected to electrowinning 198 to recover the solubilized precious metals and form a barren solution 199.

When the eluant is a polythionate the polythionate and thiosulfate tend to be co-reduced with the precious metal at the cathode to produce elemental sulfur, which interferes with the efficient continued operation of the electrowinning circuit while the polythionate and thiosulfate are also wastefully oxidized to sulfate ions at the anode.
These problems are overcome by the present invention through the use of sulfite in the pregnant solution. Sulfite is added to the eluant and/or to the pregnant solution 196 prior to, during, or after electrowinning. Preferably, sulfite is added to the eluant prior to the elution step 194. In the presence of sulfite, the precious metal is reduced at the cathode while the sulfite is oxidized to sulfate at the anode. This has the benefit of lowering the cell voltage required. Preferably, the concentration of sulfite in the pregnant solution 196 (in the elution and electrowinning steps 194, 198) is at least about 0.01M and more preferably ranges from about 0.1 to about 2 M. The sulfite is preferably in the pregnant solution with another eluant, such as any of the eluants noted above.
The stripped or barren solution 199 is removed from the electrowinning cell(s) and returned to the elution step 194. A bleedstream (not shown) of the barren solution 199 can be used to control buildup of impurities such as sulfate.
The recovered precious metal 195, which contains the precious metal recovered in electrowinning and impurities, is subjected to retorting 193 by known techniques to remove the impurities and form precious metal sludge. The sludge, which contains at least most of the precious metal in the recovered precious metal 195, is refined to produce a precious metal product of high purity.

Fig. 2 depicts another embodiment of a process for thiosulfate leaching of a 5 refractory precious metal-containing material. Fig. 2 shows an alternative to resin-in-pulp for precious metal recovery. Following leaching 132, the precious metal bearing solution 144 is separated 200 from the solids by any suitable means, such as by counter-current decantation washing and/or filtration. Preferably, at least about 95%
and more preferably at least about 99% of the precious metal is separated from the solids with the 10 latter going to tailings impoundment.

The separated precious metal bearing solution 204 is directed to the precious metal precipitation - thiosulfate regeneration step 208. This process can be carried out in any suitably agitated reactor or plurality of agitated reactors. The pH of the precious metal bearing solution 204 is adjusted if necessary to about pH 5.5-12, more preferably 15 about pH 7-11, even more preferably about pH 9-11 using a suitable basic reagent such as sodium hydroxide and the solution is contacted with a reductant, preferably a sulfide and/or bisulfide and/or polysulfide reagent to precipitate at least about 99%
of the precious metal and convert at least about 90% of the polythionates to thiosulfate, effectively regenerating the thiosulfate lixiviant. The effectiveness of the conversion 20 causes significantly less thiosulfate reagent to be consumed during the process than for conventional thiosulfate leaching processes. The use of a sulfide and/or bisulfide and/or polysulfide has the added benefit of reducing impurities such as copper or mercury or manganese from solution thereby reducing the rate of thiosulfate degradation.
While not wishing to be bound by any theory, it is believed that the most likely composition of the precipitate is the metallic precious metal and/or a precious metal sulfide, such as Au2S.
Maximum precipitation of gold and regeneration of thiosulfate is accomplished by adding at least a stoichiometric amount of reductant (relative to the dissolved precious metal and polythionate concentrations) to reduce the solution ORP to at least about -150 mV (SHE). The temperature is preferably maintained in the range of about 5 to 40 C, and more preferably at ambient temperature, about 20 C. The retention time is about 5 minutes to about 2 hours, more preferably about 15 minutes to about 1 hour.
The process is conducted under oxygen-depleted conditions, with the solution preferably containing no more than about 1 ppm dissolved molecular oxygen and more preferably less than about 0.2 ppm dissolved molecular oxygen concentration, by bubbling an oxygen-deficient gas such as nitrogen into the slurry and/or maintaining a blanket of nitrogen in the atmosphere over the slurry as noted above.

The precious metal bearing precipitate is separated from the regenerated solution 212 by any suitable method such as filtration, CCD, and the like and the separated precious metal 216 is recovered by refining in furnaces.

The regenerated solution 220 is directed to the conditioning step 224, which can be conducted in any suitably agitated reactor or plurality of reactors. The solution pH is adjusted to a value suitable for recycling the solution back to grinding 108 and/or for precious metal scavenging 228. Preferably, the pH ranges from about pH 7 to about pH
12, more preferably about pH 8 to pH 10. The solution 220 is agitated in the presence of an oxygen-containing atmosphere, such as air, to oxidize any remaining reductant (such as sulfide or bisulfide or polysulfide) carried over from the precious metal precipitation - thiosulfate regeneration step 208. The duration of the conditioning step 224 is preferably not sufficient to cause more than about 5% of the thiosulfate to form polythionates, or to yield a polythionate concentration of more than about 0.003M. The majority (typically at least about 80 vol%) of the conditioned solution 232 is then recycled in recycle solution 236. A minor portion (e.g., from about 2 to about vol%) of the conditioned solution or bleed stream 240 may have to be bled to tailings to control the buildup of impurities, such as soluble sulfate and metallic impurities. Prior to discharge to tailings the bleed portion 240 of the conditioned solution 232 is directed to the precious metal scavenging step 228 to recover any precious metals remaining in solution that were not recovered in the precious metal precipitation -thiosulfate regeneration step 208. Precious metal scavenging can be accomplished, by any suitable gold recovery technique such as by passing the bleed solution 240 tlirough a column containing a strong base resin to adsorb the precious metal. While not wishing to be bound by any theory, precipitated precious metal can be redissolved due to trace amount of molecular oxygen in the solution and incomplete reduction of polythionates in the solution. Because the amount of polythionates in the bleed is negligible, a resin-in-column recovery technique will have an excellent ability to load any remaining dissolved precious metal.

In an alternative configuration (not shown), the precious metal precipitates are redissolved in a suitable solvent, such as nitric/hydrochloric acid, cyanide, thiosulfate, thiourea chloride/chlorine and bromide/bromine to provide a precious metal-containing solution. The precious metal can then be recovered by electrolysis as noted above in connection with step 198 of Fig. 1.

This process is preferred in certain applications over the process of Fig. 1.
For certain precious metal-containing materials, it is difficult to obtain high rates of precious metal adsorption onto resins while maintaining the precious metal in solution.
The use of an RIP pretreatment step, though beneficial, can be difficult to use without experiencing some precious metal precipitation. Conditioning 156 may not be completely effective in redissolving gold precipitates, which would be discarded with the barren solids to tailings. The process of Fig. 2 can also be more robust, simpler, and therefore easier to design and operate than the process of Fig. 1.

Fig. 3 shows an alternative to Fig. 2 in which thiosulfate leaching is conducted in two stages to achieve more effective recovery of the precious metal content. Leaching is first conducted at atmospheric pressure and ambient temperature in the presence of an oxygen-containing gas such as air or industrially available oxygen (step 300) to dissolve from about 30 to 95% of the leachable precious metal content. The leachable precious metal content is defined as that portion of the precious metal content that is physically accessible to the thiosulfate lixiviant and is not encapsulated within constituents contained in the host material. The precious metal bearing solution 304 is separated from the solids 308 (step 200), the solids 308 are repulped with a portion 310 of the recycle solution 236, and the resulting slurry 308 is then directed to pressure leaching (step 312) to dissolve the majority, ie. about 5-70%, of the remaining leachable precious metal content that was not recovered in atmospheric leaching 300. In pressure leaching the solids are leached under superatmospheric conditions such as the conditions described previously (step 132 of Fig. 1). The molecular oxygen partial pressure in leach 300 preferably ranges from the molecular oxygen partial pressure at ambient conditions (e.g., more than about 3 psia at sea level) to about 15 psia and the molecular oxygen partial pressure in leach 312 preferably ranges from more than 15 psia to about 500 psia. The slurry 316 exiting pressure leaching 312 is then processed in essentially the same manner as the slurry exiting leaching 132 in Fig. 2. That is, the slurry 316 is subjected to solid/liquid separation 320 in the presence of wash water to separate the barren solid material 324 from 'the (second) pregnant leach solution 328. The first and second pregnant leach solutions 304, 328 are subjected to precious metal precipitation -thiosulfate regeneration 208, further solid/liquid separation 212, conditioning 224 and precious metal scavenging 228 as noted above in connection with Fig. 2.
The process of Fig. 3 typically performs the bulk of the leaching, or precious metal dissolution, under ambient conditions, which is much cheaper than leaching under superatmospheric conditions. The more-difficult-to-dissolve precious metals and weakly preg-robbed precious metals are then dissolved in a higher pressure leach.
Because less precious metal remains to be dissolved, the high pressure leach can have a shorter residence time and therefore lower capacity than would be possible in the absence of the ambient pressure leach.
Fig. 4 depicts another embodiment of the present invention. The process is similar to those discussed above except that thiosulfate leaching is performed by heap leaching 400 techniques. The comminuted precious metal-containing material 404 can be directly formed into a heap (in which case the material would have a preferred P80 size of from about 2 inches to about 1/4 inch, possibly agglomerated and formed into a heap.
The thiosulfate lixiviant (which commonly includes a recycled thiosulfate lixiviant 236 mixed with a makeup (fresh) thiosulfate solution(not shown)) is applied to the top of the heap using conventional techniques, and the pregnant leach solution 408 is collected from the base of the heap. Refming can be performed using any of the techniques noted above.

To facilitate extraction of gold from sulfidic and/or carbonaceous materials, the thiosulfate leach step in any of the above processes can be preceded by one or more pretreatment steps to destroy or neutralize the carbon-containing and/or sulfidic minerals.
By way of example, the intermediate steps can include one or more of biooxidation or chemical oxidation to oxidize sulfides, ultrafine grinding to liberate occluded precious metals, conventional roasting to destroy carbon- and/or sulfide-containing minerals, and/or microwave roasting.

A gold ore from Nevada, designated Sample A, was subjected to thiosulfate leaching under oxygen pressure at varying temperatures. The ore assayed 24.1 g/t gold, 2.59% iron, 0.31% total sulfur, 0.28% sulfide sulfur, 3.40% total carbon, 1.33% organic carbon and 0.02% graphitic carbon. From a diagnostic leaching analysis of the ore it was determined that a maximum of 83% of the contained gold was capable of being solubilized while the remaining gold was inaccessible to a lixiviant because it was encapsulated within pyrite and/or other minerals contained in the ore.

The ore was ground to 80% passing 200 mesh (75 m). Samples of the ore were slurried with water to a pulp density of 33% solids in a mechanically agitated laboratory autoclave. The natural pH of the slurry at ambient temperature was 8.3. The pH
of the slurry was adjusted to 9 with sodium hydroxide and a quantity of sodium thiosulfate reagent was added to adjust the initial leach solution thiosulfate concentration to 0.1 molar (M). The autoclave was sealed and pressurized to 100 psig oxygen with pure (95%
plus) oxygen gas and the slurry was heated to the desired temperature (if required).
Leaching was maintained for 6 hours, during which pulp samples were taken at 2 and 4 hours in order to monitor gold extraction with time. Upon termination of leaching, the slurry was filtered and the residue solids were washed with a dilute thiosulfate solution.
The residue solids and leach solution were assayed for gold to determine the final gold extraction.

The results were as follows:
Leach temp. Leach time Calc'd head Residue Au ext'n ( c) (Hours) Au (g/t) Au (g/t) (%) 20 2 33.3 4 41.9 6 22.8 9.44 58.5 5 40 2 51.2 4 55.1 6 26.4 9.25 64.9 60 2 63.7 4 68.5 6 22.8 4.26 81.3 60 (repeat) 2 65.2 4 73.0 6 80.9 The results indicate that the rate and extent of gold extraction was improved with 10 increasing temperature and leach time in the temperature range 20-60 C. The best results were obtained at 60 C, with about 81% gold extraction obtained after 6 hours leaching, this representing about 98% of the leachable gold content of the ore.

A second gold ore from Nevada, designated Sample B, was subjected to 15 thiosulfate leaching under oxygen pressure at varying initial pH's. The ore assayed 9.45 g/t gold, 2.50% iron, 0.39% total sulfur, 0.36% sulfide sulfur, 4.20% total carbon, 1.46%
organic carbon and 0.05% graphitic carbon. From a diagnostic leaching analysis of the ore it was determined that 82% of the contained gold was capable of being solubilized.
Samples of the ore were prepared and leached as described in Example 1, except the 20 temperature was 60 C in each test, the autoclave was pressurized with 50 psig oxygen, the initial pH was adjusted to either 9, 11 or 12, and the leach retention time was extended to 8 hours for the pH 11 and 12 tests.

The results were as follows:

Initial Leach Time Calc'd Head Residue Au Ext'n pH (hours) Au (g/t) Au (g/t) (%) 9 1 50.2 2 62.4 4 72.0 6 8.49 2.10 75.3 11 1 41.3 2 63.0 4 69.3 8 8.61 2.00 76.8 12 1 6.4 2 1.0 4 13.6 8 8.61 3.34 61.2 The results indicate that there was not much difference in gold leaching behaviour over the initial pH range of 9-11 (it should be noted that the pH
tended to decline during leaching). However, gold leaching was suppressed during the first 4 hours of leaching at pH 12, but then started to recover.

A third gold ore sample from Nevada, Sample C, was subjected to thiosulfate leaching un.der oxygen pressure at varying temperatures. The head analysis of the ore was as follows:
Gold Ore Sample C
Au, g/t 9.50 C(t), % 4.45 Fe, % 2.52 C (C03)1% 3.12 Cu, ppm 40 C (org), % 1.38 As, ppm 647 S(2-), % 0.35 Hg, ppm 14 S(t), % 0.27 Ca, % 9.0 Mg, % 1.5 From a diagnostic leaching analysis of the ore it was determined that 83% of the contained gold was capable of being solubilized.

The ore was ground to 80% passing 200 mesh (75 m). Samples of the ore were slurried with water to a pulp density of 33% solids in a mechanically agitated laboratory autoclave. The initial pH of the slurry was adjusted to approximately 11 with sodium hydroxide, after which the autoclave was sealed and pressurized to 100 psig oxygen with pure (95% plus) oxygen gas and the slurry was heated to the desired temperature. To initiate leaching, a quantity of sodium thiosulfate stock solution was injected to adjust the leach solution thiosulfate concentration to 0.1 M. Leaching was continued for 6 to 10 hours, during which no additional reagents were added. Pulp samples were taken at set intervals during leaching in order to monitor gold extraction with time.
Upon termination of leaching, the slurry was filtered and the residue solids were washed with a dilute thiosulfate solution. The residue solids and leach solution were assayed for gold to determine the final gold extraction.

Fig. 5 depicts graphically the effect of leach temperature, in the range 40-80 C, on the rate of gold extraction from Sample C. It can be seen that the gold leached quickly at 60 C and 80 C, there being little difference in the extraction rate at the two temperatures. The gold extraction peaked at approximately 83%, the maximum extractable, after 6 hours leaching. Gold leaching was slowed if the temperature was lowered to 40 C, but 80% gold extraction was still obtained after 10 hours leaching at 40 C.

An overall summary of the results is provided below:

Parameter Test #6 Test #25 Test #15 Leach time, hours 8 6 10 Final pH 7.0 8.7 9.3 Final ORP, mV (SHE) 307 242 225 Calc'd Head Au, g/t 9.48 9.43 9.27 Residue Au, g/t 1.59 1.63 1.81 Au Ext'n, % 83.2 82.7 80.5 The gold ore designated Sample C was subjected to thiosulfate leaching at varying oxygen pressures. Samples of the ore were prepared and leached as described in Example 3 except the temperature was maintained at 60 C in each test and the oxygen partial pressure was varied.

Fig. 6 portrays the effect of oxygen partial pressure, in the range 0-200 psig, on the rate of gold extraction from Sample C (in the 0 psig 02 test, the autoclave was not pressurized but the head space was maintained with pure oxygen at atmospheric pressure). It can be seen that the rate of gold leaching was somewhat sensitive to oxygen pressure, in that the rate increased with increasing pressure, particularly during the first two hours of leaching. After 6 hours leaching, gold extraction varied from a low of 78%
at 0 psig 02 to a high of 83% at 200 psig O2.

An overall summary of the results is provided below:

Parameter Test #7 Test #25 Test #22 Test #28 Test #31 200 psig 02 100 psig 02 50 psig 02 10 psig 02 0 psig 02 Leach time, hours 8 6 6 6 6 Final pH NA 8.7 9.0 9.3 9.4 Final ORP, mV (SHE) NA 242 223 216 232 Calc'd Head Au, g/t 9.78 9.43 9.40 8.95 9.08 Residue Au, g/t 1.68 1.63 1.77 1.72 2.00 Au Ext'n, % 82.8 82.7 81.1 80.8 78.0 NA = not analyzed The gold ore designated Sample C was subjected to thiosulfate leaching under oxygen pressure at varying initial sodium thiosulfate concentrations. Samples of the ore were prepared and leached as described in Example 3 except the temperature was maintained at 60 C in each test and the initial sodium thiosulfate concentration was varied.

Fig. 7 portrays the effect of initial sodium thiosulfate concentration, in the range 0.05-0.2 M, on the rate of gold extraction from Sample C. It can be seen that the rate of gold leaching was insensitive to initial thiosulfate concentration in the 0.1-0.2 M range.
At 0.05 M thiosulfate, the rate was reduced significantly, particularly during the first two hours of leaching. After 6 hours leaching gold extraction was 78% at 0.05 M
thiosulfate compared to 82% achieved at both 0.1 M and 0.2 M thiosulfate concentration.

An overall summary of the results is provided below:

Parameter Test #4 Test #25 Test #8 0.2 M 0.1 M 0.05 M
Leach time, hours 8 6 6 Final pH 8.7 8.7 8.5 Final ORP, mV (SHE) NA 242 262 Calc'd Head Au, g/t 8.85 9.43 9.40 Residue Au, g/t 1.50 1.63 1.87 Au Ext'n, % 83.0 82.7 80.1 NA = not analysed The gold ore designated Sample C was subjected to thiosulfate leaching under oxygen pressure at two different pulp densities. Samples of the ore were prepared and leached as described in Example 3, except the temperature was maintained at 60 C in each test and the leach pulp density was either 33% or 45% solids by weight.

Fig. 8 portrays the effect of 33% vs. 45% pulp density on the rate of gold extraction from Sample C. The rate of gold leaching was found to be essentially insensitive to pulp density in this range.

An overall summary of the results is provided below:
Parameter Test #26 Test #25 45% pulp 33% pulp density density Leach time, hours 6 6 Final pH 8.5 8.7 Final ORP, mV (SHE) 286 242 Calc'd Head Au, g/t 9.29 9.43 Residue Au, g/t 1.68 1.63 Au Ext'n, % 81.9 82.7 A fourth gold ore sample from Nevada, Sample D, was subjected to thiosulfate leaching at 60 C and 10 psig oxygen pressure at the natural pH of the slurry, for 8 hours.
5 The head analysis of the ore was as follows:
Gold Ore Sample D
Au,g/t 12.15 C(t),% 4.31 Fe, % 2.09 C (C03)1% 3.02 Cu, ppm 39 C (org), % 1.30 10 As, ppm 692 S(2-), % 0.12 Hg, ppm 27 S(t), % 0.22 Ca, % 8.9 Mg, % 1.3 From a diagnostic leaching analysis of the ore it was determined that 80% of the 15 contained gold was capable of being solubilized.

The ore was ground to 80% passing 200 mesh (75 m). A sample of the ore was slurried with water to a pulp density of 40% solids in a mechanically agitated laboratory autoclave. The autoclave was sealed and pressurized to 100 psig oxygen with pure (95%
plus) oxygen gas and the slurry was heated to 60 C. To initiate leaching, a quantity of 20 sodium thiosulfate stock solution was injected to adjust the leach solution thiosulfate concentration to 0.1 M. Leaching was continued for 8 hours, during which no additional reagents were added. Pulp samples were taken at set intervals during leaching in order to monitor gold extraction and remaining thiosulfate with time. Upon termination of leaching, the slurry was filtered and the residue solids were washed with a dilute 25 thiosulfate solution. The residue solids and leach solution were assayed for gold to determine the final gold extraction.

Fig. 9 depicts percent gold extraction and percent remaining thiosulfate with time.
Gold extraction reached 79.3% after 8 hours, or 99% of the leachable gold content.
Thiosulfate consumption was low, with 86.7% of the thiosulfate remaining after 8 hours 30 and available for reuse.

An overall summary of the results is provided below:

Parameter Test #37-01 Leach time, hours 8 Initial pH 7.9 Final pH 9.0 Initial ORP, mV (SHE) 411 Final ORP, mV (SHE) 397 Calc'd head Au, g/t 11.50 Residue Au, g/t 2.38 Gold extraction, % 79.3 Amount of thiosulfate remaining, % 86.7 A thiosulfate leach discharge slurry was heated to 60 C in an agitated reactor in preparation for RIP pre-treatment, the objective being to reduce the polythionate content without precipitating gold. The slurry was kept under a nitrogen atmosphere to ensure the dissolved oxygen content was maintained below 0.2 mg/L. A single dose of a 0.26 M sodium bisulfide (NaHS) solution, adjusted to pH 9, was added and the pretreatment was allowed to proceed at 60 C and ambient pressure for 2 hours. The ainount of sulfide added was 150% of stoichiometric based on the amount required to convert the polythionates back to thiosulfate in accordance with the following reactions:

2S4O6 + SZ + 32H2O-> 92S2O3 + 3H+

S3 6 + S2 ~ 2S2 O3 -A suminary of the results is provided below:

Time Au S2032 S4062S3062 ORP pH
(min) (mg/L) (g/L) (g/L) (g/L) (mV) 0 4.36 8.38 0.51 0.59 240 6.9 120 4.36 11.0 0.06 0.10 5 6.7 The tetrathionate and trithionate concentrations were reduced by 88% and 83%
respectively while all of the gold remained in solution.

A pregnant thiosulfate leach solution was adjusted to pH 10 with sodium hydroxide in an agitated reactor in preparation for sulfide treatment, the objective being to regenerate thiosulfate and precipitate the gold. The solution was kept under a nitrogen atmosphere to ensure the dissolved oxygen content was maintained below 0.2 mg/L. A
single dose of a 0.26 M sodium sulfide (Na2S) solution was added and the treatment was allowed to proceed for 2 hours at ambient temperature (22 C) and pressure. The amount of sulfide added was 100% of stoichiometric based on the amount required to convert the polythionates back to thiosulfate in accordance with the following reactions:

2S4O6 + S2 + 32 H2O -> 92 S2O3 + 3H+
S3O6 + S2 -4 2S2O3-A summary of the results is provided below:

Time Au S2032 S4062 S3062 ORP pH
(min) (mg/L) (g/L) (g/L) (g/L) (mV) 0 4.12 7.8 0.84 1.47 200 10.0 10 0.05 9.9 0.01 0.01 -210 11.0 0.02 9.9 0.01 0.01 -220 10.4 0.01 9.9 0.01 0.01 -230 10.2 60 0.01 9.8 0.01 0.01 -260 10.3 20 90 0.01 10.1 0.01 0.01 -260 10.3 120 0.01 10.2 0.01 0.01 -260 10.3 The rate of conversion of polythionates to thiosulfate was extremely fast under ainbient conditions, with essentially complete conversion achieved after about 25 minutes. Similarly, gold precipitation was also fast and essentially complete after about 30 minutes.
While this invention has been described in conjunction with the specific embodiments thereof, it is evident that many alternatives, modifications, and variations will be apparent to those skilled in the art. Accordingly, preferred embodiments of the 30 invention as set forth herein are intended to be illustrative, not limiting. By way of example, any source of sulfur species with an oxidation state less than +2 may be used in any of the above process steps to convert polythionates to thiosulfate. The regeneration step 184 in Fig. 1 can be performed in a variety of locations.
For example, regeneration 184 can be performed in the recycle loop after CCD 172 and before grinding 108, between grinding 108 and thickening 116, in the thickener 116 immediately before or during, leaching 132 and/or between resin in pulp 164 and CCD
172. Fresh thiosulfate can also be added in a number of locations. For example, fresh thiosulfate can be added in any of the locations referred to previously for the regeneration step 184 and can be added after or during regeneration 184 as noted above or in a separate tank or location. In Fig. 3, a lixiviant other than thiosulfate, such as cyanide, can be used in the atmospheric leach 300 with thiosulfate being used in the pressure leach 312. These and other changes may be made without departing from the spirit and scope of the present invention.

Claims

THE EMBODIMENTS OF THE INVENTION FOR WHICH AN EXCLUSIVE
PROPERTY OR PRIVILEGE IS CLAIMED ARE DEFINED AS FOLLOWS:

1. A process for recovering a precious metal from a precious metal-containing material, comprising:
(a) contacting the precious metal-containing material with a thiosulfate lixiviant and a molecular oxygen-containing gas at a superatmospheric pressure in the substantial or complete absence of added ammonia to solubilize the precious metal and form a pregnant thiosulfate leach solution containing the solubilized precious metal; and (b) thereafter recovering the solubilized precious metal from the pregnant thiosulfate leach solution, wherein a total ammonia concentration in the thiosulfate lixiviant is less than about 0.05M.

2. The process of Claim 1, wherein, in the contacting step (a), the thiosulfate lixiviant has a pH of less than 9 and a temperature of from about 40 to about 80°C, and the total ammonia concentration in the thiosulfate lixiviant is less than about 0.03M.

3. The process of Claim 1, wherein, in the contacting step (a), the thiosulfate lixiviant contains less than about 20 ppm of copper ion.

4. The process of Claim 1, wherein the thiosulfate lixiviant includes dissolved molecular oxygen and wherein the thiosulfate lixiviant in step (a) is free of added copper.

5. The process of Claim 1, wherein the thiosulfate lixiviant contains no more than about 0.01 M of added sulfite.

6. The process of Claim 1, wherein the thiosulfate lixiviant includes sulfite and a total concentration of sulfite in the thiosulfate lixiviant is no more than about 0.02M.

7. The process of Claim 1, wherein the thereafter recovering step (b) comprises:
(i) contacting the pregnant thiosulfate leach solution with a sulfide-containing reagent to convert most or all of polythionates to thiosulfates, wherein, in the contacting step (i), the pregnant thiosulfate leach solution has a dissolved molecular oxygen content of no more than about 1 ppm; and (ii) thereafter recovering a precious metal precipitate from the thiosulfate leach solution.

8. The process of Claim 7, further comprising before the contacting step (i):
(iii) separating most or all of the precious metal-containing material from most or all of the pregnant thiosulfate leach solution.

9. The process of Claim 7, wherein the sulfide-containing reagent is selected from the group consisting of polysulfides, bisulfides, sulfides, and mixtures thereof.

10. The process of Claim 7, wherein the pregnant thiosulfate leach solution in contacting step (i) has a pH ranging from about pH 5.5 to about pH 12.

11. The process of Claim 7, wherein in the contacting step (i), most or all of the solubilized precious metal is precipitated by reaction with the sulfide-containing reagent.

12. The process of Claim 11, wherein the thereafter recovering step (b) includes separating the precious metal precipitate from a barren thiosulfate leach solution.
13. The process of Claim 12, further comprising:
(c) adjusting the barren thiosulfate leach solution to a pH of from about pH 7 to about pH 12;

(d) contacting the barren thiosulfate leach solution with a gas including at least about 5 vol. % molecular oxygen to oxidize any remaining sulfide-containing reagent; and (e) recovering the precious metal from at least a portion of the barren thiosulfate leach solution.

14. The process of Claim 1, further comprising, after the contacting step (a):

(c) converting most or all of at least one of trithionate and tetrathionate in the pregnant thiosulfate leach solution into thiosulfate while maintaining most or all of the solubilized precious metal dissolved in the pregnant thiosulfate leach solution.

15. The process of Claim 14, wherein, the converting step (c) includes:
contacting the pregnant thiosulfate leach solution with at least one of a sulfite and a sulfide.

16. The process of Claim 14, wherein, the converting step (c) includes:
heating the pregnant thiosulfate leach solution to a temperature of more than 60°C.

17. The process of Claim 1, further comprising:
(c) contacting the pregnant thiosulfate leach solution and/or a barren thiosulfate leach solution with a sulfide and/or a bisulfide and/or a polysulfide to convert polythionates in the pregnant thiosulfate leach solution and/or the barren thiosulfate leach solution into thiosulfate; and (d) thereafter contacting the pregnant thiosulfate leach solution and/or the barren thiosulfate leach solution with an oxidant to solubilize any precipitated precious metal precipitates.

18. The process of Claim 17, wherein the oxidant is molecular oxygen, the concentration of dissolved molecular oxygen in the pregnant thiosulfate leach solution and/or the barren thiosulfate leach solution is at least about 1ppm, and the pregnant thiosulfate leach solution and/or the barren thiosulfate leach solution has a pH of from about pH 5.5 to about pH 12.

19. A process for recovering a precious metal from a carbonaceous precious metal-containing material, comprising:
(a) contacting the carbonaceous precious metal-containing material with a thiosulfate lixiviant in substantial or complete absence of added copper and added ammonia to solubilize the precious metal and form a pregnant thiosulfate leach solution containing the solubilized precious metal; and (b) thereafter recovering the solubilized precious metal from the pregnant thiosulfate leach solution, wherein, in the contacting step (a), the thiosulfate lixiviant contains less than about 20 ppm of copper ion.

20. The process of Claim 19, wherein, in the contacting step (a) the thiosulfate lixiviant has a pH of less than 9 and a temperature of from about 40 to about 80°C, and a total ammonia concentration in the thiosulfate lixiviant is less than about 0.05M.

21. The process of Claim 19, wherein, in the contacting step (a), the thiosulfate lixiviant contains less than about 15 ppm of copper ion.

22. The process of Claim 19, wherein the thiosulfate lixiviant includes molecular oxygen.

23. The process of Claim 19, wherein the thiosulfate lixiviant contains no more than about 0.01M of added sulfite.

24. The process of Claim 19, wherein the thiosulfate lixiviant includes sulfite and a total concentration of sulfite in the thiosulfate lixiviant is no more than about 0.02M.

25. The process of Claim 19, wherein the pregnant thiosulfate leach solution comprises polythionate and wherein the thereafter recovering step (b) comprises:
(i) contacting the pregnant thiosulfate leach solution with a sulfide-containing reagent to convert most or all of the polythionate to thiosulfate, wherein, in the contacting step (i), the pregnant thiosulfate leach solution has a dissolved molecular oxygen content of no more than about 1 ppm; and (ii) thereafter recovering a precious metal precipitate from the thiosulfate leach solution.

26. The process of Claim 25, further comprising before the contacting step (i):
(iii) separating most or all of the precious metal-containing material from most or all of the pregnant thiosulfate leach solution.

27. The process of Claim 25, wherein the sulfide-containing reagent is selected from the group consisting of polysulfides, bisulfides, sulfides, and mixtures thereof.

28. The process of Claim 25, wherein the pregnant thiosulfate leach solution in contacting step (i) has a pH ranging from about pH 5.5 to about pH 12.

29. The process of Claim 25, wherein in the contacting step (i), most or all of the solubilized precious metal is precipitated by reaction with the sulfide-containing reagent.

30. The process of Claim 29, wherein the thereafter recovering step (b) includes separating the precious metal precipitate from a barren thiosulfate leach solution.
31. The process of Claim 30, further comprising:
(c) adjusting the barren thiosulfate leach solution to a pH of from about pH 7 to about pH 12;

(d) contacting the barren thiosulfate leach solution with a gas including at least about 5 vol. % molecular oxygen to oxidize any of the sulfide-containing reagent remaining; and (e) recovering the precious metal from at least a portion of the barren thiosulfate leach solution.

32. The process of Claim 19, further comprising, after the contacting step (a):
(c) converting most or all of at least one of trithionate and tetrathionate in the pregnant thiosulfate leach solution into thiosulfate while maintaining most or all of the solubilized precious metals dissolved in the pregnant thiosulfate leach solution.

33. The process of Claim 32, wherein, the converting step (c), includes:
contacting the pregnant thiosulfate leach solution with at least one of a sulfite and a sulfide.

34. The process of Claim 32, wherein, the converting step (c), includes:
heating the pregnant thiosulfate leach solution to a temperature of more than 60°C.

35. The process of Claim 19, further comprising:
(c) contacting the pregnant thiosulfate leach solution and/or a barren thiosulfate leach solution with a sulfide and/or a bisulfide and/or a polysulfide to convert polythionates in the pregnant thiosulfate leach solution and/or the the barren thiosulfate leach solution into thiosulfate; and (d) thereafter contacting the pregnant thiosulfate leach solution and/or the barren thiosulfate leach solution with an oxidant to solubilize precipitated precious metal precipitates.

36. The process of Claim 35, wherein the oxidant is molecular oxygen, the concentration of dissolved molecular oxygen in the pregnant thiosulfate leach solution and/or the barren thiosulfate leach solution is at least about 1 ppm, and the pregnant thiosulfate leach solution and/or the barren thiosulfate leach solution has a pH of from about pH 5.5 to about pH 12.

37. A process for recovering a precious metal from a precious metal-containing material, comprising:
(a) contacting the precious metal-containing material with a thiosulfate lixiviant and a molecular oxygen-containing gas at a superatmospheric pressure in the substantial or complete absence of at least one of added copper and added ammonia to solubilize the precious metal and form a pregnant thiosulfate leach solution containing the solubilized precious metal; and (b) thereafter recovering the solubilized precious metal from the pregnant thiosulfate leach solution, wherein, in the contacting step (a) the thiosulfate lixiviant contains less than about 20 ppm of copper ion.

38. The process of Claim 37, wherein, in the contacting step (a) the thiosulfate lixiviant has a pH of less than 9 and a temperature of from about 40 to about 80 C, and the total ammonia concentration in the thiosulfate lixiviant is less than about 0.05M.

39. The process of Claim 37, wherein the lixiviant includes dissolved molecular oxygen.

40. The process of Claim 37, wherein the thiosulfate lixiviant contains no more than about 0.01 M of added sulfite and wherein, in the contacting step, the thiosulfate lixiviant contains less than about 15 ppm of copper ion.

41. The process of Claim 37, wherein the thiosulfate lixiviant includes sulfite and a total concentration of sulfite in the thiosulfate lixiviant is no more than about 0.02M.

42. The process of Claim 37, further comprising:
maintaining a dissolved molecular oxygen content in at least one of the thiosulfate lixiviant and the pregnant thiosulfate leach solution at or below about 1 ppm to inhibit formation of trithionate and tetrathionate.

44. The process of Claim 37, wherein the thereafter recovering step (b) comprises:
(i) contacting the pregnant thiosulfate leach solution with a sulfide-containing reagent to convert most or all of polythionates to thiosulfates, wherein, in the contacting step (i), the pregnant thiosulfate leach solution has a dissolved molecular oxygen content of no more than about 1 ppm; and (ii) thereafter recovering a precious metal precipitate from the thiosulfate leach solution.

45. The process of Claim 44, further comprising before the contacting step (i):
(iii) separating most or all of the precious metal-containing material from most or all of the pregnant thiosulfate leach solution.

46. The process of Claim 44, wherein the sulfide-containing reagent is selected from the group consisting of polysulfides, bisulfides, sulfides, and mixtures thereof.

47. The process of Claim 44, wherein the pregnant thiosulfate leach solution in contacting step (i) has a pH ranging from about pH 5.5 to about pH 12.

48. The process of Claim 44, wherein in the contacting step (i), most or all of the solubilized precious metal is precipitated by reaction with the sulfide-containing reagent.

49. The process of Claim 48, wherein the thereafter recovering step (b) includes separating the precious metal precipitate from a barren thiosulfate leach solution.

50. The process of Claim 49, further comprising:
(c) adjusting the barren thiosulfate leach solution to a pH of from about pH 7 to about pH 12;
(d) contacting the barren thiosulfate leach solution with a gas including at least about 5 vol. % molecular oxygen to oxidize any of the sulfide-containing reagent remaining; and (e) recovering the precious metal from at least a portion of the barren thiosulfate leach solution.

51. The process of Claim 37, further comprising, after the contacting step (b):
(c) converting most or all of at least one of trithionate and tetrathionate in the pregnant thiosulfate leach solution into thiosulfate while maintaining most or all of the solubilized precious metals dissolved in the pregnant thiosulfate leach solution.

52. The process of Claim 51, wherein, the converting step (c) includes:
contacting the pregnant thiosulfate leach solution with at least one of a sulfite and a sulfide.

53. The process of Claim 51, wherein, the converting step (c), includes:
heating the pregnant thiosulfate leach solution to a temperature of more than 60°C.

54. The process of Claim 37, further comprising:
(c) contacting the pregnant thiosulfate leach solution and/or a barren thiosulfate leach solution with a sulfide and/or a sulfide and/or a polysulfide to convert polythionates in the pregnant thiosulfate leach solution and/or the barren thiosulfate leach solution into thiosulfate; and (d) thereafter contacting the pregnant thiosulfate leach solution and/or the barren thiosulfate leach solution with an oxidant to solubilize precipitated precious metal precipitates.

55. The process of Claim 54, wherein the oxidant is molecular oxygen, the concentration of dissolved molecular oxygen in the pregnant thiosulfate leach solution and/or the barren thiosulfate leach solution is at least about 1ppm, and the pregnant thiosulfate leach solution and/or the barren thiosulfate leach solution has a pH of from about pH 5.5 to about pH 12.

56. A process for recovering a precious metal from a carbonaceous precious metal-containing material, comprising:
(a) contacting the carbonaceous precious metal-containing material with a thiosulfate lixiviant in the substantial or complete absence of added copper and added ammonia to solubilize the precious metal and form a pregnant thiosulfate leach solution containing the solubilized precious metal; and (b) thereafter recovering the solubilized precious metal from the pregnant thiosulfate leach solution, wherein a total ammonia concentration in the thiosulfate lixiviant is less than about 0.05M.

57. The process of Claim 56, wherein, in the contacting step (a), the thiosulfate lixiviant has a pH of less than 9 and a temperature of from about 40 to about 80°C.

58. The process of Claim 56, wherein, in the contacting step (a), the thiosulfate lixiviant contains less than 20 ppm of copper ion.

59. The process of Claim 56, wherein the lixiviant includes molecular oxygen.
60. The process of Claim 56, wherein the thiosulfate lixiviant contains no more than about 0.01 M of added sulfite.

61. The process of Claim 56, wherein the thiosulfate lixiviant includes sulfite and a total concentration of sulfite in the thiosulfate lixiviant is no more than about 0.02M.

62. The process of Claim 56, wherein the solubilized precious metal is recovered from the pregnant thiosulfate leach solution by electrowinning.

63. The process of Claim 56, further comprising:
maintaining a dissolved molecular oxygen content in at least one of the thiosulfate lixiviant and pregnant thiosulfate leach solution at or below about 1 ppm to inhibit the formation of the trithionate and tetrathionate.

64. The process of Claim 56, wherein the thereafter recovering step (b) comprises:
(i) contacting the pregnant thiosulfate leach solution with a sulfide-containing reagent to convert most or all of polythionates to thiosulfates, wherein, in the contacting step (i), the pregnant thiosulfate leach solution has a dissolved molecular oxygen content of no more than about 1 ppm; and (ii) thereafter recovering a precious metal precipitate from the thiosulfate leach solution.

65. The process of Claim 64, further comprising before the contacting step (i):
(iii) separating most or all of the precious metal-containing material from most or all of the pregnant thiosulfate leach solution.

66. The process of Claim 64, wherein the sulfide-containing reagent is selected from the group consisting of polysulfides, bisulfides, sulfides, and mixtures thereof.

67. The process of Claim 64, wherein the pregnant thiosulfate leach solution in contacting step (i) has a pH ranging from about pH 5.5 to about pH 12.

68. The process of Claim 64, wherein in the contacting step (i), most or all of the solubilized precious metal is precipitated by reaction with the sulfide-containing reagent.

69. The process of Claim 68, wherein the thereafter recovering step (b) includes separating the precious metal precipitate from a barren thiosulfate leach solution.
70. The process of Claim 69, further comprising:
(c) adjusting the barren thiosulfate leach solution to a pH of from about pH 7 to about pH 12;
(d) contacting the barren thiosulfate leach solution with a gas including at least about 5 vol. % molecular oxygen to oxidize any of the sulfide-containing reagent remaining; and (e) recovering the precious metal from at least a portion of the barren thiosulfate leach solution.

71. The process of Claim 56, further comprising, after the contacting step (a):
(c) converting most or all of at least one of trithionate and tetrathionate in the pregnant thiosulfate leach solution into thiosulfate while maintaining most or all of the solubilized precious metals dissolved in the pregnant thiosulfate leach solution.

72. The process of Claim 71, wherein, the converting step (c), includes:
contacting the pregnant thiosulfate leach solution with at least one of a sulfite and a sulfide.

73. The process of Claim 71, wherein, the converting step (c) includes:
heating the pregnant thiosulfate leach solution to a temperature of more than 60°C.

74. The process of Claim 56, further comprising:
(c) contacting the pregnant thiosulfate leach solution and/or a barren thiosulfate leach solution with a sulfide and/or a bisulfide and/or a polysulfide to convert polythionates in the pregnant thiosulfate leach solution and/or the barren thiosulfate leach solution into thiosulfate; and (d) thereafter contacting the pregnant thiosulfate leach solution and/or the barren thiosulfate leach solution with an oxidant to solubilize precipitated precious metal precipitates.

75. The process of Claim 74, wherein the oxidant is molecular oxygen, the concentration of dissolved molecular oxygen in the pregnant leach solution and/or the barren thiosulfate leach solution is at least about 1ppm, and the pregnant thiosulftate leach solution and/or the barren thiosulfate leach solution has a pH of from about pH 5.5 to about pH 12.
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Families Citing this family (63)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
US6660059B2 (en) 2000-05-19 2003-12-09 Placer Dome Technical Services Limited Method for thiosulfate leaching of precious metal-containing materials
US6451089B1 (en) * 2001-07-25 2002-09-17 Phelps Dodge Corporation Process for direct electrowinning of copper
US20050126923A1 (en) * 2001-07-25 2005-06-16 Phelps Dodge Corporation Process for recovery of copper from copper-bearing material using medium temperature pressure leaching, direct electrowinning and solvent/solution extraction
US6451088B1 (en) * 2001-07-25 2002-09-17 Phelps Dodge Corporation Method for improving metals recovery using high temperature leaching
CA2363762A1 (en) * 2001-11-23 2003-05-23 Golden Wave Resources Inc. Electromagnetic pyrolysis metallurgy
AUPS334402A0 (en) * 2002-07-02 2002-07-25 Commonwealth Scientific And Industrial Research Organisation Process for recovering precious metals
JP4144311B2 (en) * 2002-10-08 2008-09-03 住友金属鉱山株式会社 Methods for separating and recovering platinum group elements
AU2007211912B2 (en) * 2002-11-15 2008-01-31 Placer Dome Technical Services Limited Method for thiosulfate leaching of precious metal-containing materials
US7722840B2 (en) * 2002-11-15 2010-05-25 Placer Dome Technical Services Limited Method for thiosulfate leaching of precious metal-containing materials
EP1433860A1 (en) * 2002-12-23 2004-06-30 Paques B.V. Process for regenerating thiosulphate from a spent thiosulphate gold leachant
US7285256B2 (en) * 2003-04-04 2007-10-23 Newmont Usa Limited Precious metal recovery using thiocyanate lixiviant
US20040237721A1 (en) * 2003-05-29 2004-12-02 Morteza Baghalha Anoxic leaching of precious metals with thiosulfate and precious metal oxidants
MXPA03006955A (en) * 2003-08-04 2005-02-09 Univ Autonoma Metropolitana Silver and gold leaching and recovery process with electro-oxidised thiourea solutions.
EP1805336B1 (en) 2004-10-29 2012-08-15 Freeport-McMoran Corporation Process for recovery of copper from copper-bearing material using pressure leaching, direct electrowinning and solvent/solution extraction
US7736487B2 (en) 2004-10-29 2010-06-15 Freeport-Mcmoran Corporation Process for recovery of copper from copper-bearing material using pressure leaching, direct electrowinning and solution extraction
US7604783B2 (en) 2004-12-22 2009-10-20 Placer Dome Technical Services Limited Reduction of lime consumption when treating refractor gold ores or concentrates
US7572317B2 (en) * 2005-11-10 2009-08-11 Barrick Gold Corporation Thiosulfate generation in situ in precious metal recovery
WO2007053947A1 (en) * 2005-11-10 2007-05-18 Barrick Gold Corporation Thiosulfate generation in situ in precious metal recovery
CN101421426B (en) * 2006-03-03 2015-05-27 鑫亚国际有限公司 Process for extracting gold from gold-bearing ore
US8061888B2 (en) 2006-03-17 2011-11-22 Barrick Gold Corporation Autoclave with underflow dividers
US7972413B2 (en) * 2006-04-07 2011-07-05 Metal Asia International Ltd. Precious metal recovery from solution
WO2007137325A1 (en) * 2006-05-25 2007-12-06 Commonwealth Scientific & Industrial Research Organisation Process for recovering metals from resins
US8252254B2 (en) 2006-06-15 2012-08-28 Barrick Gold Corporation Process for reduced alkali consumption in the recovery of silver
US20090056501A1 (en) * 2007-08-29 2009-03-05 Vale Inco Limited Hydrometallurgical process using resin-neutralized-solution of a heap leaching effluent
EA020950B1 (en) 2007-09-17 2015-03-31 Баррик Гольд Корпорейшн Method to improve recovery of gold from double refractory gold ores
WO2009037594A2 (en) * 2007-09-18 2009-03-26 Barrick Gold Corporation Process for recovering gold and silver from refractory ores
US8262770B2 (en) 2007-09-18 2012-09-11 Barrick Gold Corporation Process for controlling acid in sulfide pressure oxidation processes
CA2705487A1 (en) * 2007-11-28 2009-06-04 Barrick Gold Corporation Microbial pre-treatment of double refractory gold ores
DE102007061806A1 (en) * 2007-12-19 2009-06-25 Mettler-Toledo Ag Process for the regeneration of amperometric sensors
WO2011067680A2 (en) 2009-12-04 2011-06-09 Barrick Gold Corporation Separation of cooper minerals from pyrite using air-metabisulfite treatment
FR2966470B1 (en) * 2010-10-25 2012-12-14 Christian Queyroix PROCESS AND TREATMENT SYSTEM FOR EXTRACTING GOLD CONTENT FROM GOLDEN ORE.
US8753513B2 (en) 2010-11-09 2014-06-17 International Business Machines Corporation Ammonia-peroxide wastewater treatment system
JP5146522B2 (en) 2010-11-26 2013-02-20 東京エレクトロン株式会社 Substrate processing apparatus, substrate processing method, and storage medium
NZ611643A (en) * 2010-12-07 2014-06-27 Barrick Gold Corp Co-current and counter current resin-in-leach in gold leaching processes
JP5554285B2 (en) * 2011-03-30 2014-07-23 Jx日鉱日石金属株式会社 Gold leaching method
AR086933A1 (en) * 2011-06-15 2014-01-29 Barrick Gold Corp METHOD FOR RECOVERING PRECIOUS METALS AND COPPER OF LIXIVIATE SOLUTIONS
WO2013023297A1 (en) * 2011-08-15 2013-02-21 Her Majesty The Queen In Right Of Canada As Represented By The Minister Of Natural Resources Canada Process of leaching precious metals
EP2836617A4 (en) * 2012-04-09 2016-01-13 Process Res Ortech Inc Chloride process for the leaching of gold
WO2014110518A1 (en) * 2013-01-14 2014-07-17 Simmons William D Flotation circuit for oxide and sulfide ores
US10161016B2 (en) 2013-05-29 2018-12-25 Barrick Gold Corporation Method for pre-treatment of gold-bearing oxide ores
GB2536781B (en) 2013-08-15 2018-10-10 Hatch Ltd Multi-compartment reactor and method for controlling retention time in a multi-compartment reactor
EP2907899B1 (en) * 2014-05-30 2016-07-27 Nicolae Costache Method for recovering metal and nonmetal elements from objects comprising metal and organic compounds
CN104032138B (en) * 2014-06-05 2016-04-20 东北大学 The method of a kind of dump leaching-mercurous tailings of precipitation stabilization treatment
US10030906B2 (en) 2014-12-18 2018-07-24 Electrolux Home Products, Inc. Refrigerator
US10814306B2 (en) 2015-07-27 2020-10-27 6th Wave Innovations Corp. Molecularly imprinted polymer beads for extraction of metals and uses thereof
US9777346B2 (en) 2015-09-03 2017-10-03 Battelle Energy Alliance, Llc Methods for recovering metals from electronic waste, and related systems
GB2551980A (en) * 2016-06-30 2018-01-10 Commw Scient Ind Res Org Method and system for low level metal analysis of mineral samples
CA3042533A1 (en) * 2016-11-03 2018-05-11 6th Wave Innovations Corp. Molecularly imprinted polymer beads for extraction of lithium, mercury, and scandium
DE102016015661A1 (en) 2016-12-31 2018-07-19 Thomas Helle Process for the purification of slag from heavy metals
CN108193049B (en) * 2017-11-29 2020-04-07 贵州省地质矿产中心实验室 Method for chemical pre-oxidation synchronous cyanide-free leaching of primary gold ore
CN109499625B (en) * 2018-11-06 2021-09-28 长春黄金研究院有限公司 High-efficiency elution method for gold-loaded resin
PE20211512A1 (en) * 2019-01-21 2021-08-11 Barrick Gold Corp METHOD FOR CARBON-CATALYZED THOSULFATE LEACHING OF MATERIALS CONTAINING GOLD
CN109852816B (en) * 2019-02-27 2020-05-05 武汉理工大学 Method for adsorbing noble metal complex ions in thiosulfate leaching solution by sulfide ore
CA3141559A1 (en) 2019-06-03 2020-12-10 Barrick Gold Corporation Method for recovering precious metals from thiosulfate leach solutions
EP4100554A4 (en) * 2020-02-07 2024-03-20 Univ Kentucky Res Found Extraction of copper, gold and other elements from waste materials
EP4136266A1 (en) * 2020-04-17 2023-02-22 Glencore Technology Pty Limited Sulphide oxidation in leaching of minerals
CN111647743B (en) * 2020-04-28 2021-11-09 西北矿冶研究院 Gold and silver mineral leaching combined medicament and preparation method thereof
US11236407B1 (en) * 2020-07-31 2022-02-01 Rio Tinto Technological Resources Inc. Metal recovery by leaching agglomerates of metal-containing material/pyrite
US11286540B2 (en) * 2020-07-31 2022-03-29 Rio Tinto Technological Resources Inc. Method of processing a pyrite-containing slurry
CA3192359A1 (en) 2020-08-18 2022-02-24 Enviro Metals, LLC Metal refinement
CN113546446A (en) * 2021-07-28 2021-10-26 江西东江环保技术有限公司 Method for recovering copper in BCC synthetic mother liquor by using cationic resin
CN115232964B (en) * 2022-08-15 2023-06-09 云南大学 Method for extracting gold from alkaline solution of gold thiosulfate complex based on natural eutectic solvent
CN115323381A (en) * 2022-09-18 2022-11-11 长春黄金研究院有限公司 Shiny silverweed herb agent and preparation method thereof

Family Cites Families (116)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
US496951A (en) 1893-05-09 Henry parkes
US1627582A (en) 1922-09-07 1927-05-10 Thomas H Sherian Process for treatment of ores
US1627583A (en) * 1925-10-23 1927-05-10 George E Thum Cone carrier
GB954435A (en) * 1960-08-25 1964-04-08 Nat Res Dev Anion exchange resins for the recovery of gold and silver from gold and silver bearing aqueous cyanide liquors
US3524724A (en) 1966-10-12 1970-08-18 Continental Oil Co Method for making ammonium thiosulfate and ammonium sulfate
US3644087A (en) * 1970-02-09 1972-02-22 Universal Oil Prod Co Process for scrubbing sulfur dioxide from a gas stream
GB1423342A (en) 1971-12-09 1976-02-04 Iws Nominee Co Ltd Polymeric compounds and process for their preparation
BE792337A (en) 1971-12-13 1973-03-30 Falconbridge Nickel Mines Ltd PROCESS FOR THE RECOVERY OF PRECIOUS METALS FROM MATERIALS CONTAINING COPPER
US3833351A (en) * 1973-02-15 1974-09-03 Univ Eng Inc Continuous preparation of pure metals by hydrogen reduction
GB1497534A (en) * 1973-12-13 1978-01-12 Matthey Rustenburg Refines Refining of metals
US3902896A (en) 1974-05-22 1975-09-02 Int Nickel Co Cementation of metals from acid solutions
CA1073681A (en) 1976-02-20 1980-03-18 Roman M. Genik-Sas-Berezowsky Recovery of precious metals from metal sulphides
CA1106617A (en) 1978-10-30 1981-08-11 Grigori S. Victorovich Autoclave oxidation leaching of sulfide materials containing copper, nickel and/or cobalt
US4296075A (en) 1978-11-24 1981-10-20 Mobil Oil Corporation Method for protecting an ion-exchange resin from chemical poisoning
CA1139569A (en) 1979-04-13 1983-01-18 Harold J. Heinen Leaching gold - silver ores
US4289532A (en) 1979-12-03 1981-09-15 Freeport Minerals Company Process for the recovery of gold from carbonaceous ores
US4369061A (en) * 1979-12-28 1983-01-18 Kerley Jr Bernard J Recovery of precious metals from difficult ores
US4269622A (en) 1979-12-28 1981-05-26 Kerley Jr Bernard J Recovery of precious metals from difficult ores
US4411873A (en) * 1980-12-31 1983-10-25 Mobil Oil Corporation In-line regeneration of polythionate poisoned ion exchange resins
US4384889A (en) 1981-04-15 1983-05-24 Freeport Research & Development Company Simultaneous leaching and cementation of precious metals
CA1198080A (en) 1981-04-15 1985-12-17 Freeport Minerals Company Simultaneous leaching and electrodeposition of precious metals
US4411612A (en) 1981-04-16 1983-10-25 Neha International Apparatus for recovering precious metals from their ores
RO81261A2 (en) 1981-12-01 1983-02-01 Institutul De Cercetari Si Proiectari Pentru Epurarea Apelor Reziduale,Ro PROCESS OF RECOVERY OF SILVER AND SUPPORT FROM WASTE OF PHOTOSENSIBLE MATERIALS
US4489984A (en) * 1982-04-22 1984-12-25 Mobil Oil Corporation In-situ uranium leaching process
ZA845087B (en) * 1983-07-08 1985-03-27 Kalocsai Guy Imre Z Dissolution of noble metals
US4585561A (en) 1983-07-25 1986-04-29 Agfa-Gevaert Aktiengesellschaft Flotation process for the continuous recovery of silver or silver compounds from solutions or dispersions
DE3347165A1 (en) 1983-12-27 1985-07-04 Skw Trostberg Ag, 8223 Trostberg METHOD FOR EXTRACTION OF PRECIOUS METALS
US4552589A (en) 1984-01-27 1985-11-12 Getty Oil Company Process for the recovery of gold from refractory ores by pressure oxidation
ES530538A0 (en) 1984-03-13 1985-07-16 Nunez Alvarez Carlos PROCEDURE TO IMPROVE THE PERFORMANCE OF GOLD AND SILVER EXTRACTION IN THE GOSSAN MINERALS
JPS60208434A (en) 1984-04-03 1985-10-21 Nippon Mining Co Ltd Method for recovering silver from precipitate of copper electrolysis
CA1234290A (en) 1984-09-27 1988-03-22 Donald R. Weir Recovery of gold from refractory auriferous iron- containing sulphidic material
CA1229487A (en) 1984-09-27 1987-11-24 Roman M. Genik-Sas-Berezowsky Process for the recovery of silver from a residue essentially free of elemental sulphur
CA1234991A (en) 1984-09-27 1988-04-12 Donald R. Weir Recovery of gold from auriferous refractory iron- containing sulphidic ore
US4634187A (en) * 1984-11-21 1987-01-06 Isl Ventures, Inc. Method of in-situ leaching of ores
ZA853701B (en) 1984-11-26 1986-05-28 Pm Mineral Leaching Tech Inc Bioleaching process
JPS61127834A (en) 1984-11-27 1986-06-16 日本鉱業株式会社 Recovery of mercury in iron sulfide concentrate
JPS61127833A (en) 1984-11-27 1986-06-16 日本鉱業株式会社 Recovery of mercury in iron sulfide concentrate
SU1284942A1 (en) 1984-12-03 1987-01-23 Предприятие П/Я М-5400 Method of producing sodium thiosulfate
US4740243A (en) 1984-12-31 1988-04-26 Ensci, Inc. Metal value recovery from metal sulfide containing ores
SU1279954A1 (en) 1985-04-08 1986-12-30 Березниковский химический завод Method of producing sodium thiosulfate
US4816234A (en) 1985-05-10 1989-03-28 Kamyr, Inc. Utilization of oxygen in leaching and/or recovery procedures employing carbon
US4654078A (en) 1985-07-12 1987-03-31 Perez Ariel E Method for recovery of precious metals from difficult ores with copper-ammonium thiosulfate
PH21188A (en) * 1985-09-10 1987-08-12 Butler Dean Leaching process
GB2180829B (en) 1985-09-20 1989-08-16 Aurotech N L Precious metal extraction
US4738718A (en) 1985-10-28 1988-04-19 Freeport Minerals Company Method for the recovery of gold using autoclaving
US5232490A (en) * 1985-11-27 1993-08-03 Leadville Silver And Gold Oxidation/reduction process for recovery of precious metals from MnO2 ores, sulfidic ores and carbonaceous materials
US4723998A (en) 1985-11-29 1988-02-09 Freeport Minerals Co Recovery of gold from carbonaceous ores by simultaneous chlorine leach and ion exchange resin adsorption process
US4721526A (en) 1986-08-13 1988-01-26 Kamyr, Inc. Heap leaching with oxygen
US4801329A (en) 1987-03-12 1989-01-31 Ensci Incorporated Metal value recovery from carbonaceous ores
US4765827A (en) 1987-01-20 1988-08-23 Ensci, Inc. Metal value recovery
US4778519A (en) * 1987-02-24 1988-10-18 Batric Pesic Recovery of precious metals from a thiourea leach
US4816235A (en) 1987-02-24 1989-03-28 Batric Pesic Silver and manganese recovery using acidified thiourea
CA1306613C (en) 1987-05-15 1992-08-25 Guy Deschenes Recovery of gold from aqueous solutions
GB8726158D0 (en) 1987-11-07 1987-12-09 British Petroleum Co Plc Separation process
US5607619A (en) 1988-03-07 1997-03-04 Great Lakes Chemical Corporation Inorganic perbromide compositions and methods of use thereof
US4923510A (en) 1988-10-31 1990-05-08 Gopalan Ramadorai Treatment of refractory carbonaceous sulfide ores for gold recovery
US4902345A (en) 1989-01-12 1990-02-20 Newmont Gold Co. Treatment of refractory carbonaceous and sulfidic ores or concentrates for precious metal recovery
MY105658A (en) 1989-03-07 1994-11-30 Butler Dean R Noble metal recovery
JP2632576B2 (en) 1989-05-12 1997-07-23 日鉱金属株式会社 Desorption method of gold iodine complex from ion exchange resin
GB9002311D0 (en) 1990-02-02 1990-04-04 Rio Tinto Minerals Dev Separation process
US5071477A (en) 1990-05-03 1991-12-10 American Barrick Resources Corporation of Toronto Process for recovery of gold from refractory ores
US5244493A (en) 1990-09-21 1993-09-14 Newmont Gold Co. Biometallurgical treatment of precious metal ores having refractory carbon content
US5127942A (en) 1990-09-21 1992-07-07 Newmont Mining Corporation Microbial consortium treatment of refractory precious metal ores
US5114687A (en) 1990-12-14 1992-05-19 South Dakota School Of Mines & Technology Ammonia extraction of gold and silver from ores and other materials
US6248301B1 (en) 1991-04-12 2001-06-19 Newmont Mining Corporation And Newmont Gold Company Process for treating ore having recoverable metal values including arsenic containing components
US5147617A (en) 1991-05-21 1992-09-15 Freeport-Mcmoran Inc. Process for recovery of gold from gold ores using a complexing pretreatment and sulfurous acid leaching
US5147618A (en) 1991-05-21 1992-09-15 Freeport-Mcmoran Inc. Process for recovery of gold from refractory gold ores using sulfurous acid as the leaching agent
US5332559A (en) 1991-07-10 1994-07-26 Newmont Gold Co. Biooxidation process for recovery of metal values from sulphur-containing ore materials
US5246486A (en) 1991-07-10 1993-09-21 Newmont Gold Co. Biooxidation process for recovery of gold from heaps of low-grade sulfidic and carbonaceous sulfidic ore materials
US5354359A (en) 1992-04-01 1994-10-11 Newmont Gold Co. Hydrometallurgical process for the recovery of precious metal values from precious metal ores with thiosulfate lixiviant
US5236492A (en) 1992-07-29 1993-08-17 Fmc Gold Company Recovery of precious metal values from refractory ores
US5364453A (en) 1992-09-22 1994-11-15 Geobiotics, Inc. Method for recovering gold and other precious metals from carbonaceous ores
US5338338A (en) 1992-09-22 1994-08-16 Geobiotics, Inc. Method for recovering gold and other precious metals from carbonaceous ores
US5308381A (en) 1993-04-15 1994-05-03 South Dakota School Of Mines & Techology Ammonia extraction of gold and silver from ores and other materials
US5484470A (en) 1994-07-28 1996-01-16 E. I. Du Pont De Nemours And Company Enhancement of gold lixiviation using nitrogen and sulfur heterocyclic aromatic compounds
US5405430A (en) 1994-04-12 1995-04-11 Groves; William D. Recovery of precious metals from evaporite sediments
US5449397A (en) 1994-06-24 1995-09-12 Hunter; Robert M. Biocatalyzed leaching of precious metal values
US5489326A (en) 1994-10-04 1996-02-06 Barrick Gold Corporation Gold recovery using controlled oxygen distribution pressure oxidation
US5536480A (en) 1994-11-29 1996-07-16 Santa Fe Pacific Gold Corporation Method for treating mineral material having organic carbon to facilitate recovery of gold and silver
US5683490A (en) 1994-12-23 1997-11-04 The United States Of America As Represented By The Secretary Of The Interior Solution mining of precious metals using aqueous, sulfur-bearing solutions at elevated temperatures
US5536297A (en) 1995-02-10 1996-07-16 Barrick Gold Corporation Gold recovery from refractory carbonaceous ores by pressure oxidation and thiosulfate leaching
US5785736A (en) * 1995-02-10 1998-07-28 Barrick Gold Corporation Gold recovery from refractory carbonaceous ores by pressure oxidation, thiosulfate leaching and resin-in-pulp adsorption
US5837210A (en) * 1995-04-18 1998-11-17 Newmont Gold Company Method for processing gold-bearing sulfide ores involving preparation of a sulfide concentrate
US5653945A (en) 1995-04-18 1997-08-05 Santa Fe Pacific Gold Corporation Method for processing gold-bearing sulfide ores involving preparation of a sulfide concentrate
GB2310424B (en) 1996-02-22 1999-09-29 Finch Limited Process for recovering gold from oxide-based refractory ores
AU727691B2 (en) 1996-06-26 2000-12-21 Cognis Ip Management Gmbh Process for the recovery of precious metal values from aqueous ammoniacal thiosulfate leach solutions
US6197214B1 (en) 1996-06-26 2001-03-06 Henkel Corporation Ammonium thiosulfate complex of gold or silver and an amine
CA2209559C (en) 1996-07-16 2001-12-18 Barrick Gold Corporation Gold recovery from refractory carbonaceous ores by pressure oxidation, thiosulfate leaching and resin-in-leach adsorption
US5733431A (en) 1996-08-21 1998-03-31 Hw Process Technologies, Inc. Method for removing copper ions from copper ore using organic extractions
CA2193305A1 (en) 1996-12-18 1998-06-18 Jean-Marc Lalancette Process for removing and recovering copper, silver and zinc from sulfide ores
US5961833A (en) 1997-06-09 1999-10-05 Hw Process Technologies, Inc. Method for separating and isolating gold from copper in a gold processing system
GB2341602A (en) 1997-06-09 2000-03-22 Hw Process Technologies Inc Method for separating and isolating precious metals from non precious metals dissolved in solutions
AUPO900097A0 (en) 1997-09-05 1997-10-02 Arton (No 001) Pty Ltd Process
CA2307500C (en) 1997-10-30 2010-01-12 Hw Process Technologies, Inc. Method for removing contaminants from process streams in metal recovery processes
US6251163B1 (en) 1998-03-04 2001-06-26 Placer Dome, Inc. Method for recovering gold from refractory carbonaceous ores
US6368381B1 (en) 1998-03-11 2002-04-09 Placer Dome Technical Services, Ltd. Autoclave using agitator and sparge tube to provide high oxygen transfer rate to metal-containing solutions
US6183706B1 (en) 1998-03-11 2001-02-06 Placer Dome, Inc. Autoclave having an agitator with an aerating impeller for high oxygen transfer rate to metal-containing slurries and method of use
KR100615773B1 (en) 1998-05-08 2006-08-25 쉘 오일 캄파니 Process to recover molybdenum and vanadium metals from spent catalyst by alkaline leaching
CA2315480A1 (en) 1999-08-13 2001-02-13 Antonio T. Robles Process for removing metals from a sorbent
AUPQ315799A0 (en) 1999-09-29 1999-10-21 Murdoch University Improved process for the elution of gold from anion exchange resins
AUPQ456299A0 (en) 1999-12-09 2000-01-13 Geo2 Limited Recovery of precious metals
US20030154822A1 (en) * 1999-12-09 2003-08-21 John Hall Recovery of precious metals
US6451275B1 (en) * 2000-03-10 2002-09-17 Lakefield Research Limited Methods for reducing cyanide consumption in precious metal recovery by reducing the content of intermediate sulfur oxidation products therein
US6344068B1 (en) 2000-04-04 2002-02-05 Barrick Gold Corporation Process for recovering gold from thiosulfate leach solutions and slurries with ion exchange resin
US6660059B2 (en) 2000-05-19 2003-12-09 Placer Dome Technical Services Limited Method for thiosulfate leaching of precious metal-containing materials
AU9330601A (en) 2000-09-29 2002-04-08 Newmont Mining Corp Method and apparatus for chemical processing
US6500231B1 (en) 2001-03-29 2002-12-31 Newmont Usa Limited Recovery of precious metals from thiosulfate solutions
AU783904B2 (en) 2001-04-10 2005-12-22 Grd Minproc Limited Improved processing of precious metal-containing materials
US6632264B2 (en) 2001-04-17 2003-10-14 The University Of British Columbia Gold recovery from thiosulfate leaching
US6641642B2 (en) 2001-12-21 2003-11-04 Newmont Usa Limited High temperature pressure oxidation of ores and ore concentrates containing silver using controlled precipitation of sulfate species
AU2002246310A1 (en) 2002-03-26 2003-10-08 Council Of Scientific And Industrial Research Process for the recovery of gold and silver from used refractory bricks
US6602319B1 (en) 2002-04-01 2003-08-05 Council Of Scientific And Industrial Research Process for the recovery of gold and silver from used refractory bricks
AUPS334402A0 (en) 2002-07-02 2002-07-25 Commonwealth Scientific And Industrial Research Organisation Process for recovering precious metals
US7722840B2 (en) 2002-11-15 2010-05-25 Placer Dome Technical Services Limited Method for thiosulfate leaching of precious metal-containing materials
EP1433860A1 (en) 2002-12-23 2004-06-30 Paques B.V. Process for regenerating thiosulphate from a spent thiosulphate gold leachant
AU2003904385A0 (en) 2003-08-18 2003-08-28 Murdoch University Improved Thiosulphate Leach Process

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