CA2412352A1 - Method for thiosulfate leaching of precious metal-containing materials - Google Patents
Method for thiosulfate leaching of precious metal-containing materials Download PDFInfo
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- CA2412352A1 CA2412352A1 CA002412352A CA2412352A CA2412352A1 CA 2412352 A1 CA2412352 A1 CA 2412352A1 CA 002412352 A CA002412352 A CA 002412352A CA 2412352 A CA2412352 A CA 2412352A CA 2412352 A1 CA2412352 A1 CA 2412352A1
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- thiosulfate
- precious metal
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/06—Extraction 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
- C22B3/08—Sulfuric acid, other sulfurated acids or salts thereof
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B11/00—Obtaining noble metals
- C22B11/04—Obtaining noble metals by wet processes
- C22B11/042—Recovery of noble metals from waste materials
- C22B11/044—Recovery of noble metals from waste materials from pyrometallurgical residues, e.g. from ashes, dross, flue dust, mud, skim, slag, sludge
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- Y—GENERAL 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
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
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Abstract
Processes are provided for recovering precious metals from refractory materials using thiosulfate lixiviants. The processes can employ heap leaching or lixiviants that include one or more blinding agents.
Description
~rl~T~ao I~oR T»~c~su~~l~AT~. ~~ACr~~lvc of PC10US 141L'1'A1.~-CONTAINI11G MATERIALS
FtI;LD Oh' Tlil~ t'~IV~NTION
I~ 'the present invention is directed generally to the recovery ofprecious metals from precious metal-ccrnta~ining material and specifically to the recovery ofprecious metals from pr~ious natal-containutg material using thiosulfate lixiviants.
t 5 A traditional technique for re-covering precious mctal(s} from prcaious rnetal-e~ntaining ore is by teaching the material with a cyanide lixivisnt. As used herein, a "precious metal" refers to gold, sih~er, and the platinum group metals (e.g., platinum, palladium, ruthenium, rhodium, osmium, and iridium). Many countries arc placing severe limitations an the use ofcyanide du~c to the tleteterious eiFects ofcyanide on the environment.
?U Lneidents of fish and other wildlife haring been killed by the Icalca~ of cyanide into waterways tu~~e been r~cported. The limitations being placed on cyanide: use have inc:rcaseci substantially the cost of extracting precious metals) from ore, thereby decreasing precious n~cetal reserves in ttiany countries. Cyanide is also unable to recover precious metals such as gold from refractory ores without a pretreatment step. "Refractory ores" refer to thase 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 ores that are both sulfidic and carbonaceous.
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, but not limited to, 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 arnmine 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:
FtI;LD Oh' Tlil~ t'~IV~NTION
I~ 'the present invention is directed generally to the recovery ofprecious metals from precious metal-ccrnta~ining material and specifically to the recovery ofprecious metals from pr~ious natal-containutg material using thiosulfate lixiviants.
t 5 A traditional technique for re-covering precious mctal(s} from prcaious rnetal-e~ntaining ore is by teaching the material with a cyanide lixivisnt. As used herein, a "precious metal" refers to gold, sih~er, and the platinum group metals (e.g., platinum, palladium, ruthenium, rhodium, osmium, and iridium). Many countries arc placing severe limitations an the use ofcyanide du~c to the tleteterious eiFects ofcyanide on the environment.
?U Lneidents of fish and other wildlife haring been killed by the Icalca~ of cyanide into waterways tu~~e been r~cported. The limitations being placed on cyanide: use have inc:rcaseci substantially the cost of extracting precious metals) from ore, thereby decreasing precious n~cetal reserves in ttiany countries. Cyanide is also unable to recover precious metals such as gold from refractory ores without a pretreatment step. "Refractory ores" refer to thase 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 ores that are both sulfidic and carbonaceous.
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, but not limited to, 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 arnmine 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:
-2-Cu" + 4NH, --+ Cu{NH, );' (1) Oxidation of gold by cupric ammine, gold complexation as the gold-thiosulfate anion, and reduction of cupric ammine to cuprous thiosulfate:
Au+ Cu(NH3)4+ + SS203 ~ Au(5203)2- + Cu(5203)3 + 4NH3 (2) Oxidation of the cuprous thiosulfate back to cupric ammine with oxygen:
Cu(5203)3 +4NH3 +/0Z +~HZO~Cu(NH3)4+ +35203 +0H- (3) Summing equations (2) and (3) yields the overall thiosulfate leach reaction for gold:
Au+ 25203 + '/402 + H20 -> Au(S203)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 ofproblems. 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 thiosutfate consumption:
Cu2 + 520; + 20H -~ Cu,S + ,504 + H2O (5) Rapid oxidation of thiosulfate by cupric ammine also occurs, leading to excessive degradation and loss of thiosulfate:
2Cu( NH3 )4+ + 8S2O3 -a 2Cu(5203 )3 + S4O6 + 8NH3 (6)
Au+ Cu(NH3)4+ + SS203 ~ Au(5203)2- + Cu(5203)3 + 4NH3 (2) Oxidation of the cuprous thiosulfate back to cupric ammine with oxygen:
Cu(5203)3 +4NH3 +/0Z +~HZO~Cu(NH3)4+ +35203 +0H- (3) Summing equations (2) and (3) yields the overall thiosulfate leach reaction for gold:
Au+ 25203 + '/402 + H20 -> Au(S203)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 ofproblems. 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 thiosutfate consumption:
Cu2 + 520; + 20H -~ Cu,S + ,504 + H2O (5) Rapid oxidation of thiosulfate by cupric ammine also occurs, leading to excessive degradation and loss of thiosulfate:
2Cu( NH3 )4+ + 8S2O3 -a 2Cu(5203 )3 + S4O6 + 8NH3 (6)
-3-Loss of ammonia by volatilization occurs readily, particularly in unsealed gas-sparged reactors operating at pH greater than 9.2, leading to excessive ammonia consumption:
NH4 + OH- --> NH,~aQ~ + HZD --~ NH»g~ + H20 (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 potythionates, 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
These and other needs have been addressed by the methodologies and systems of the present invention. The methodologies can recover precious metals, such as gold and silver, 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;
..rl_ (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 O.OSM 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 I S ppm (added) copper ions;
(e) an (added) sulfite concentration that is preferably no more than about 0.01 M
such that the slurry has a maximum total concentration of sulfite ofpreferably no more than about 0.02M and more preferably no more than about O.OIM; and/or (fj 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 I S about 80%.
The thiosulfate lixiviant can be derived from any suitable forms) 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 example, 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 S 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 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 S 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 to prevent conversion of thiosulfate in the pregnant leach solution into polythionate. 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 2IK", 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 thiosulfate decomposition to polythionates is significantly reduced 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 _7_ typically at least most of polythionates (particularly trithionate and tetrathiona.te) 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 ions) Sxz-, 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 SxO62-, where x = 4 to 6, to thiosulfate. The sulfite, sulfide, and/or polysulfide can be compounded with any canon, with Group 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.
I S 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 through the leach solution. As will be appreciated, the atmosphere or gas may contain one or mare 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, _g_ the inert atmosphere (or spurge 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 oxidants) (other than thiosulfate a.nd polythionates) in the atmosphere and/or (pregnant or barren) leach solution the rate or degree of oxidation of S thiosulfates to form potythionates 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 time in hours (horizontal axis);
Fig. 8 is another plot of gold extraction in percent (vertical axis) versus leach time in hours (horizontal axis);
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); and _9_ Fig. 10 is a plot of gold extraction in percent (vertical axis) and time (horizontal axis) for aerated and unaerated columns.
DETAILED DESCRIPTION
The present invention provides an improved thiosulfate leaching process for the recovery ofprecious metals from precious metal-bearing material. The precious metals) 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 ofthe 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 su~ciently 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 gulp density (after wet comminution or formation of the comminuted material into a slurry) is less than about 20%.
The thickener overflow so lution 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- ~ S203 + 2Sz + 3H20 (8) where x = 3-6, or by reaction of bisulfide with bisulfate:
2 HS- + 4 HS03 -> 352 03 + 3 H2 0 (9) or by reaction of elemental sulfur with sulfite:
S + .503 ~ -~ S2 03 (10) If the desirable temperature is above ambient, 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 ammonia under elevated oxygen partial pressure can be illustrated by the following reaction:
Au + 2S203 + 4 Oz + y H20 --~ Au(Sz03)Z + 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 so urce of oxygen. The desired oxygen partial pressure (POD
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 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 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. Blimination 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 andlor total solution) copper concentration is no more than about 20 ppm, more preferably no more than about 15 ppm, and even more preferably no more than about 10 ppm while the (added andlor 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 of thiosulfate 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 S"O6 , where n = 2-6):
Tetrathionate formation: 2S2 03 - + y U2 + H2 O --3 S4 O6 + 2 UH- ( 12) Trithionate formation: 3.54 O6 + 2 02 + HZ O -+ 4S3 D6 + 2 H+ ( 13) Sulfite formation: S3O6 + ~ 02 + 2H20 --+ 3503 + 4H+ (14) Sulfate formation: 2503 + 02 -> 2504 (15) Overall: S2O3 - + 202 + H20 -~ 2S04 + 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 in U.S.
Patent No.
6,183,706 and copending U.S. Patent Application Serial No. 09/561,256, filed April 27, 2000, which are incorporated herein by reference. 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
NH4 + OH- --> NH,~aQ~ + HZD --~ NH»g~ + H20 (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 potythionates, 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
These and other needs have been addressed by the methodologies and systems of the present invention. The methodologies can recover precious metals, such as gold and silver, 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;
..rl_ (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 O.OSM 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 I S ppm (added) copper ions;
(e) an (added) sulfite concentration that is preferably no more than about 0.01 M
such that the slurry has a maximum total concentration of sulfite ofpreferably no more than about 0.02M and more preferably no more than about O.OIM; and/or (fj 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 I S about 80%.
The thiosulfate lixiviant can be derived from any suitable forms) 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 example, 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 S 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 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 S 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 to prevent conversion of thiosulfate in the pregnant leach solution into polythionate. 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 2IK", 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 thiosulfate decomposition to polythionates is significantly reduced 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 _7_ typically at least most of polythionates (particularly trithionate and tetrathiona.te) 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 ions) Sxz-, 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 SxO62-, where x = 4 to 6, to thiosulfate. The sulfite, sulfide, and/or polysulfide can be compounded with any canon, with Group 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.
I S 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 through the leach solution. As will be appreciated, the atmosphere or gas may contain one or mare 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, _g_ the inert atmosphere (or spurge 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 oxidants) (other than thiosulfate a.nd polythionates) in the atmosphere and/or (pregnant or barren) leach solution the rate or degree of oxidation of S thiosulfates to form potythionates 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 time in hours (horizontal axis);
Fig. 8 is another plot of gold extraction in percent (vertical axis) versus leach time in hours (horizontal axis);
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); and _9_ Fig. 10 is a plot of gold extraction in percent (vertical axis) and time (horizontal axis) for aerated and unaerated columns.
DETAILED DESCRIPTION
The present invention provides an improved thiosulfate leaching process for the recovery ofprecious metals from precious metal-bearing material. The precious metals) 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 ofthe 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 su~ciently 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 gulp density (after wet comminution or formation of the comminuted material into a slurry) is less than about 20%.
The thickener overflow so lution 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- ~ S203 + 2Sz + 3H20 (8) where x = 3-6, or by reaction of bisulfide with bisulfate:
2 HS- + 4 HS03 -> 352 03 + 3 H2 0 (9) or by reaction of elemental sulfur with sulfite:
S + .503 ~ -~ S2 03 (10) If the desirable temperature is above ambient, 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 ammonia under elevated oxygen partial pressure can be illustrated by the following reaction:
Au + 2S203 + 4 Oz + y H20 --~ Au(Sz03)Z + 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 so urce of oxygen. The desired oxygen partial pressure (POD
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 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 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. Blimination 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 andlor total solution) copper concentration is no more than about 20 ppm, more preferably no more than about 15 ppm, and even more preferably no more than about 10 ppm while the (added andlor 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 of thiosulfate 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 S"O6 , where n = 2-6):
Tetrathionate formation: 2S2 03 - + y U2 + H2 O --3 S4 O6 + 2 UH- ( 12) Trithionate formation: 3.54 O6 + 2 02 + HZ O -+ 4S3 D6 + 2 H+ ( 13) Sulfite formation: S3O6 + ~ 02 + 2H20 --+ 3503 + 4H+ (14) Sulfate formation: 2503 + 02 -> 2504 (15) Overall: S2O3 - + 202 + H20 -~ 2S04 + 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 in U.S.
Patent No.
6,183,706 and copending U.S. Patent Application Serial No. 09/561,256, filed April 27, 2000, which are incorporated herein by reference. 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 material 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 optionallybe directed to RIP pretreatment 140 to reduce the concentration of polythionates 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 S 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.
I 5 By way of example, a sulfide-containing reagent can reduce the polythionates back to thiosulfate, as shown by the following reactions:
~17~
~l g~
Any reagent that releases sulfide ions on dissolution will suffice, such as sodium bisulfide, NaHS, sodium sulfide, NazS, hydrogen sulfide gas, HZS, or a polysulfide. The use of a sulfide reagent has the advantages of rapidly and e~ciently 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:
I 5 S4 06 + S'03 -~ S2 03 + u'3 06 ( 19) Sulfite treatment is effective in reducing tetrathionate quickly, but a disadvantage is it is ineffective in reducing trithionate. The sulfite can be added in any form 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 R.IP
pretreatment 140 is preferably less than 60°C to inhibit, at least substantially, the precipitation of precious ..1g_ metals) 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
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 material 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 optionallybe directed to RIP pretreatment 140 to reduce the concentration of polythionates 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 S 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.
I 5 By way of example, a sulfide-containing reagent can reduce the polythionates back to thiosulfate, as shown by the following reactions:
~17~
~l g~
Any reagent that releases sulfide ions on dissolution will suffice, such as sodium bisulfide, NaHS, sodium sulfide, NazS, hydrogen sulfide gas, HZS, or a polysulfide. The use of a sulfide reagent has the advantages of rapidly and e~ciently 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:
I 5 S4 06 + S'03 -~ S2 03 + u'3 06 ( 19) Sulfite treatment is effective in reducing tetrathionate quickly, but a disadvantage is it is ineffective in reducing trithionate. The sulfite can be added in any form 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 R.IP
pretreatment 140 is preferably less than 60°C to inhibit, at least substantially, the precipitation of precious ..1g_ metals) 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 temperature of the leach slurry at a sufficiently high value in the absence of oxygen to effect the following hydrolytic disproportionation reactions:
45406 + 5H20 ~ 75203 + 2S04 + 10H+ (20) 5306 + H20 ~ 5203 + 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 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 perfornned in any suitable vessel(s), preferably agitated.
Preferably, RIP pretreatment is performed in a series of tanks or in a multistaged vessel.
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 polythionates in the slurry into ..20-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 preferablyranges from about -100 to about 100 mV (SHE) to substantially minimize precious metal precipitation.
The exiting RIP pretreated shury 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. Patent Application Serial No. 09/452,736, filed in June, 2000, entitled "A Process for Recovering Gold from Thiosulfate Leach Solutions and Slurries with Ion Exchange Resins", to Thomas, et al.; U.S.
Patent Application Serial No. 09/034,$46, filed March 4, 1998, entitled "Method for Recovering Gold from Refractory Carbonaceous Ores' ; and U.S. Patents 5,536,297 and 5,785,736, all ofwhich are incorporated herein by reference. Preferred resins include anion exchange resins, preferably a strong base resin including a quaternary 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 ofa 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, "Amberlite IRA 410" manufactured by Rohm and Haas, "Amberlite IRA 900" manufactured by Rohm and Haas, and "Vitrokele 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 carned 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 metals) in the leach slurry will be "loaded" or adsorbed onto the resin.
To inhibit the formation ofpolythionates 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 atmosphere. 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 1b8 (which will typically contain no more than about 0.01 ppm precious metals or 1% of the precious metals) 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 liquid/solid separation, provides water balancing in the circuit, and prevents build up of impurities in the leach circuit by removing a portion of the IS leach solution with the solids.
Regeneration 184 can be performed in one or more vessels) 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 as set forth in U.S. Application Serial No. 09/452,736 above, and a nitrate, i0 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 1 S is displaced by the eluant into the pregnant solution 196.
The stripped resin 197 is recycled 1:o 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 20 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 ehiant and typically no more than about 100 ppm and more typically from about 10 to about 500 ppm solubitized precious metals, is subjected to electrowinning 198 to recover the solubilized precious metals and form a barren solution 199. Problems in electrowinning ofprecious metals out of a medium containing polythionates and/or thiosulfate have been encountered in U.S.
patent application Serial No. 09/452,736. 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 anade. 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 0.01 M 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 cells) 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 ofa 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 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 earned 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 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% ofthe precious metal and convert at least about 90% of the polythionates to thiosulfate, effectively regenerating the thiosulfate lixiviant. The effectiveness of the conversion 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 andlor 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 AtrzS. Maximum precipitation of gold and regeneration of thiosulfate is accomplished by adding at least a stoichiornetric 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 I 5 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 _28_ 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 20 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 through 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 S connection with step 198 of Fig. I .
This process is preferred in certain applications over the process ofFig. 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 teachable precious metal content. The teachable 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 teachable precious metal content that was not recovered in atmospheric leaching 300. Inpressure 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 Asia. The slurry 316 exiting pressure leaching 312 is then processed in essentially the same manner as the slurry exiting leaching 300 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 ~~re 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-dii~cult-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.
Before heap leaching can commence, the heap is formed on a leach pad. To provide desired levels of heap porosity and permeability, the precious metal-containing material is comminuted to a desired size range, typically by wet or dry crushing or may be placed as mined ore. Typically, the material is comminuted to a Pgo size of from about 2 inches to about 1/4 inch. Prior to placement on the leach pad, fine particles {particles preferably having a size of less than about 150 ~m and more preferably of less than about 500 pm) are removed from the comminuted material by suitable screens, filters, and the like.
Alternatively or additionally, the comminuted precious metal-containing material can be formed into agglomerates or particulates by known techniques. Thiosulfate lixiviant can be contacted with the comminuted material immediately before or during agglomeration to assist in agglomerate formation and "jump start" the leaching process and/or a suitable binder can be added to provide more robust agglomerates. Possible binders include NALCO 9704, water, barren thiosulfate-containing solution, and leachate. In any event, it is preferred that the Pgo size ofthe particles/agglomerates formed into the heap is at least about 150 Vim, more preferably at least about 500 Vim, and even more preferably at least about 1,000 Vim.
Thiosulfate lixiviant can be contacted with the particles of material and/or agglomerates as the particles/agglomerates are being placed on the heap. In the process configuration of Fig.
4, the comminuted precious metal-containing material 404 is directly formed into a heap without prior agglomeration.
-~2-The pad typically includes a liquid impervious liner, which is placed beneath the heap, to collect the pregnant leach solution and prevent the pregnant leach solution from being lost to the surrounding environment. The height of each lift of the heap is typically from about 4 to about 8 m and ofthe heap itself can be up to 100 m. Rather than moving the heap after thiosulfate leaching is completed (which is done in a dynamic heap configuration), further heaps or lifts can be built on top of exhausted heap{s) (which is done in a static heap configuration).
During heap construction, a network of aerating pipes can be located in a lower portion of the heap to force an oxygen-containing gas through the heap during thiosulfate leaching. The pipes can be perforated so as to provide a substantially uniform dispersion of the gas throughout the heap. The oxygen-containing gas is typically forced through the pipes using a single- or mufti-stage compressor, blower, fan, or other mechanical device. When the oxygen-containing gas is pressurized and forced through the pipes, it typically has a pressure greater than the ambient atmosphf:ric pressure, more typically of at least about 1 inch HzO, and even more typically of at least about 30 inches HZO greater than the ambient atmspheric pressure. Preferably, at least a stoichiometric amount of molecular oxygen {relative to the amount of precious metal in the precious metal-containing material) is deliberately introduced into the heap during leaching. More preferably, at least about 0.5 kg of molecular oxygen and even more preferably from about 1 to about 10 kg of molecular oxygen is introduced into the heap during leaching for each ton of material to be leached in the heap. Preferably, at least about 2 and more preferably from about 4 to about 40 m' of oxygen-containing gas is introduced into the heap for each liter of lixiviant applied to the heap.
To control evolution ofammonia gas during forced air introduction into the heap, the pH of the thiosulfate lixiviant is controlled. Preferably, the pH of the thiosulfate lixiviant is maintained (when introduced into the heap and during passage through the heap) at a pH of no more than about pH 9, more preferably of no more than about pH 8.75, and even more preferably of from about pH 6.5 to about pH 8.75. Alternatively, the concentration of free ammonia can be maintained below levels sufficient to result in evolution of significant amounts of ammonia gas. In some configurations, the concentration of free ammonia in the thiosulfate lixiviant applied to the top of the heap is maintained at a level of no more than about 2000 ppm, more preferably no more than about 1000 ppm, and even more preferably no more than about 500 ppm. This can be realized, for example, by using sodium thiosulfate alone as the lixiviant or using a mixture of ammonium and sodium thiosulfate.
The pH can be controlled by using suitable (acid or base) buffering agents to produce the desired change in pH. In one embodiment, carbonaceous compounds, such as calcium carbonates, are added to the lixiviant before or after application to the heap and/or to the heap itself. Surprisingly, the carbon component in the buffering agent has been found under suitable conditions to perform, at most, only a minimal degree preg robbing.
Typical consumption of carbonates is in the range of about 0.5 to about 10 lb/ton of material in the heap.
To apply the thiosulfate iixiviant to the heap, a number of techniques can be employed. For example, spray systems (such as spray nozzles), drip and/or trickle systems (such as drip emitters and perforated pipes), injection holes in the heap, and irrigation ditches on top of the heap can be used to apply the lixiviant. The preferred lixiviant distribution system preferably applies the lixiviant at least substantially uniformly throughout the heap.
In a preferred configuration, the applied lixiviant flows countercurrently through the heap relative to the flow of the oxygen-containing gas. Typically, the lixiviant flows from the top to the bottom of the heap while the gas flows from the bottom to the top of the heap. In a preferred configuration, at least about 0.5 and even more preferably from about 1 to about 10t solution /t ore of lixiviant is applied to the top of the heap from start-to-finish of heap leaching. In this configuration the lixiviant is applied for at least about 0.5 and even more preferably from about 1 to 48 months from the start to finish of heal leaching.
In one configuration, the lixiviant is sparged with the oxygen-containing gas before application to the heap or the gas is otherwise contacted with the lixiviant before application to the heap (such as by in-line mixing) to cause the lixiviant to have a heightened dissolved molecular oxygen content. Additional gas may be deliberately introduced into the heap separately from the lixiviant, if desired. The dissolved molecular oxygen content of the lixiviant preferably is at least about 1 mg/L and even more preferably ranges from about 5 to about 40 mg/L.
After construction of the 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 while the oxygen-containing gas is introduced to the bottom of the heap. The pregnant leach solution 408 is collected from the base of the heap. The pregnant leach solution is recycled to the top of the heap. The recycle rate is sufficient to provide an application rate of the lixiviant to the top of the heap of from about 2 and more preferably from about 4 to about 24 LmJmz of top surface area for the heap.
During recycle, at least a portion (typically at least about S vol.% and more typically from about 50 to about 100 vol.%) of the pregnant leach solution is removed and subjected to precious metal recovery, such as by any of the techniques referred to above. A particularly preferred technique is to recover the dissolved precious metal by sulfide precipitation. As noted, sulfide precipitation not only precipitates the precious metal but also regenerates the thiosulfate by converting polythionates into thiosulfate. Sulfide precipitation is typically conducted under anerobic conditions, as noted above. Such conditions can be realized and maintained by de-aerating the pregnant leach solution with a vacuum, inert gas bubbling, and/or allowing the pregnant leach solution to stand dormant for a sufficient period of time for the dissolved oxygen level to decrease to desired levels. Sufficient sulfide is added to the pregnant leach solution to precipitate at least most of the dissolved precious metals) as precious metal sulfides and to convert at least most of the polythionates to thiosulfate. The barren thiosulfate lixiviant, after precious metal recovery, can be recombined with the remaining pregnant leach solution and returned to the top of the heap.
While not wishing to be bound by any theory, it is believed that the precious metal thiosulfate complex may be unstable under certain conditions, including those set forth above, and that the precious metal can be stripped from the thiosulfate-containing solution by a number of substances commonly encountered in precious metal-containing materials.
The substances or preg robbing materials typically absorb, adsorb or precipitate the precious metal. Such preg-robbing materials include carbonaceous materials, pyrite-containing materials, chacopyrite and iron oxides. Surprisingly and unexpectedly, blinding agents may be used in the thiosulfate li5ziviant to prevent or uihibit preg robbing of the precious metal by the preg robbing material. The blinding agent itself absorbs or adsorbs (in preference to the precious metal) or other~rise neutralizes (such as by chemical reaction) the preg robbing sites on the material. Suitable blinding agents include one or more of hydrocarban-containing substances, such as petroleumproducts (e.g., kerosene, diesel fuel, and gasoline), alcohols, esters, or aldehydes; surfactants such as detergents, sodium lauryl sulfonate, or organic phosphates; and reactive metal salts such as lead, mercury, cadmium, tin, and silver salts. In such situations, the thiosulfate lixiwiant, in any of the leaching processes discussed above, typically inchxdes at least about 0.1 mg/1.,, more typically at least about 1 mg/L, and even more typically from about 2 to about 200 mg/L of the blinding agent.
'~fhile not wishing to be bound by any theory, other agents xmay also be suitable as blinding agents. Preferably, the agents do not destabilize thiosulfate in solution. Agents which act as oxidation catalysts, can destabilize thiosulfate in solution. By gray of example, copper salts under certain conditions are not preferred as a blinding agent as copper salts under these conditions can catalyze thiosulfate decomposition.
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 nwre of bioo~cidation or chemical oxidation to oxidize sulfides, ultrafxne grinding to liberate occluded precious metals, conventional roasting to destro~r carbon- and/or sulfide-containing minerals, and/or microwave roasting.
EX.A_MP'LE 1 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 m,axirnurn of 83% of the contained gold was capable of being solubili2ed while the remaining gold was inaccessible to a lrxi~riant because it was encapsulated withili 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 naturalpH ofthe slurry at ambient temperature was 8.3. The p~T
ofthe shrny 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 a hours in order to monitor gold extraction with time. Upon termination of leaching, the slurry was tittered 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 fo Mows:
Leach Temp. Leach TimeCalc'd HeadResidue Au Ext'n (C) (hours) Au (g/t) Au (g/t)(%) 20 2 33.3 4 41.9
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 temperature of the leach slurry at a sufficiently high value in the absence of oxygen to effect the following hydrolytic disproportionation reactions:
45406 + 5H20 ~ 75203 + 2S04 + 10H+ (20) 5306 + H20 ~ 5203 + 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 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 perfornned in any suitable vessel(s), preferably agitated.
Preferably, RIP pretreatment is performed in a series of tanks or in a multistaged vessel.
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 polythionates in the slurry into ..20-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 preferablyranges from about -100 to about 100 mV (SHE) to substantially minimize precious metal precipitation.
The exiting RIP pretreated shury 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. Patent Application Serial No. 09/452,736, filed in June, 2000, entitled "A Process for Recovering Gold from Thiosulfate Leach Solutions and Slurries with Ion Exchange Resins", to Thomas, et al.; U.S.
Patent Application Serial No. 09/034,$46, filed March 4, 1998, entitled "Method for Recovering Gold from Refractory Carbonaceous Ores' ; and U.S. Patents 5,536,297 and 5,785,736, all ofwhich are incorporated herein by reference. Preferred resins include anion exchange resins, preferably a strong base resin including a quaternary 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 ofa 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, "Amberlite IRA 410" manufactured by Rohm and Haas, "Amberlite IRA 900" manufactured by Rohm and Haas, and "Vitrokele 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 carned 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 metals) in the leach slurry will be "loaded" or adsorbed onto the resin.
To inhibit the formation ofpolythionates 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 atmosphere. 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 1b8 (which will typically contain no more than about 0.01 ppm precious metals or 1% of the precious metals) 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 liquid/solid separation, provides water balancing in the circuit, and prevents build up of impurities in the leach circuit by removing a portion of the IS leach solution with the solids.
Regeneration 184 can be performed in one or more vessels) 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 as set forth in U.S. Application Serial No. 09/452,736 above, and a nitrate, i0 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 1 S is displaced by the eluant into the pregnant solution 196.
The stripped resin 197 is recycled 1:o 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 20 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 ehiant and typically no more than about 100 ppm and more typically from about 10 to about 500 ppm solubitized precious metals, is subjected to electrowinning 198 to recover the solubilized precious metals and form a barren solution 199. Problems in electrowinning ofprecious metals out of a medium containing polythionates and/or thiosulfate have been encountered in U.S.
patent application Serial No. 09/452,736. 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 anade. 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 0.01 M 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 cells) 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 ofa 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 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 earned 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 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% ofthe precious metal and convert at least about 90% of the polythionates to thiosulfate, effectively regenerating the thiosulfate lixiviant. The effectiveness of the conversion 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 andlor 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 AtrzS. Maximum precipitation of gold and regeneration of thiosulfate is accomplished by adding at least a stoichiornetric 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 I 5 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 _28_ 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 20 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 through 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 S connection with step 198 of Fig. I .
This process is preferred in certain applications over the process ofFig. 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 teachable precious metal content. The teachable 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 teachable precious metal content that was not recovered in atmospheric leaching 300. Inpressure 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 Asia. The slurry 316 exiting pressure leaching 312 is then processed in essentially the same manner as the slurry exiting leaching 300 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 ~~re 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-dii~cult-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.
Before heap leaching can commence, the heap is formed on a leach pad. To provide desired levels of heap porosity and permeability, the precious metal-containing material is comminuted to a desired size range, typically by wet or dry crushing or may be placed as mined ore. Typically, the material is comminuted to a Pgo size of from about 2 inches to about 1/4 inch. Prior to placement on the leach pad, fine particles {particles preferably having a size of less than about 150 ~m and more preferably of less than about 500 pm) are removed from the comminuted material by suitable screens, filters, and the like.
Alternatively or additionally, the comminuted precious metal-containing material can be formed into agglomerates or particulates by known techniques. Thiosulfate lixiviant can be contacted with the comminuted material immediately before or during agglomeration to assist in agglomerate formation and "jump start" the leaching process and/or a suitable binder can be added to provide more robust agglomerates. Possible binders include NALCO 9704, water, barren thiosulfate-containing solution, and leachate. In any event, it is preferred that the Pgo size ofthe particles/agglomerates formed into the heap is at least about 150 Vim, more preferably at least about 500 Vim, and even more preferably at least about 1,000 Vim.
Thiosulfate lixiviant can be contacted with the particles of material and/or agglomerates as the particles/agglomerates are being placed on the heap. In the process configuration of Fig.
4, the comminuted precious metal-containing material 404 is directly formed into a heap without prior agglomeration.
-~2-The pad typically includes a liquid impervious liner, which is placed beneath the heap, to collect the pregnant leach solution and prevent the pregnant leach solution from being lost to the surrounding environment. The height of each lift of the heap is typically from about 4 to about 8 m and ofthe heap itself can be up to 100 m. Rather than moving the heap after thiosulfate leaching is completed (which is done in a dynamic heap configuration), further heaps or lifts can be built on top of exhausted heap{s) (which is done in a static heap configuration).
During heap construction, a network of aerating pipes can be located in a lower portion of the heap to force an oxygen-containing gas through the heap during thiosulfate leaching. The pipes can be perforated so as to provide a substantially uniform dispersion of the gas throughout the heap. The oxygen-containing gas is typically forced through the pipes using a single- or mufti-stage compressor, blower, fan, or other mechanical device. When the oxygen-containing gas is pressurized and forced through the pipes, it typically has a pressure greater than the ambient atmosphf:ric pressure, more typically of at least about 1 inch HzO, and even more typically of at least about 30 inches HZO greater than the ambient atmspheric pressure. Preferably, at least a stoichiometric amount of molecular oxygen {relative to the amount of precious metal in the precious metal-containing material) is deliberately introduced into the heap during leaching. More preferably, at least about 0.5 kg of molecular oxygen and even more preferably from about 1 to about 10 kg of molecular oxygen is introduced into the heap during leaching for each ton of material to be leached in the heap. Preferably, at least about 2 and more preferably from about 4 to about 40 m' of oxygen-containing gas is introduced into the heap for each liter of lixiviant applied to the heap.
To control evolution ofammonia gas during forced air introduction into the heap, the pH of the thiosulfate lixiviant is controlled. Preferably, the pH of the thiosulfate lixiviant is maintained (when introduced into the heap and during passage through the heap) at a pH of no more than about pH 9, more preferably of no more than about pH 8.75, and even more preferably of from about pH 6.5 to about pH 8.75. Alternatively, the concentration of free ammonia can be maintained below levels sufficient to result in evolution of significant amounts of ammonia gas. In some configurations, the concentration of free ammonia in the thiosulfate lixiviant applied to the top of the heap is maintained at a level of no more than about 2000 ppm, more preferably no more than about 1000 ppm, and even more preferably no more than about 500 ppm. This can be realized, for example, by using sodium thiosulfate alone as the lixiviant or using a mixture of ammonium and sodium thiosulfate.
The pH can be controlled by using suitable (acid or base) buffering agents to produce the desired change in pH. In one embodiment, carbonaceous compounds, such as calcium carbonates, are added to the lixiviant before or after application to the heap and/or to the heap itself. Surprisingly, the carbon component in the buffering agent has been found under suitable conditions to perform, at most, only a minimal degree preg robbing.
Typical consumption of carbonates is in the range of about 0.5 to about 10 lb/ton of material in the heap.
To apply the thiosulfate iixiviant to the heap, a number of techniques can be employed. For example, spray systems (such as spray nozzles), drip and/or trickle systems (such as drip emitters and perforated pipes), injection holes in the heap, and irrigation ditches on top of the heap can be used to apply the lixiviant. The preferred lixiviant distribution system preferably applies the lixiviant at least substantially uniformly throughout the heap.
In a preferred configuration, the applied lixiviant flows countercurrently through the heap relative to the flow of the oxygen-containing gas. Typically, the lixiviant flows from the top to the bottom of the heap while the gas flows from the bottom to the top of the heap. In a preferred configuration, at least about 0.5 and even more preferably from about 1 to about 10t solution /t ore of lixiviant is applied to the top of the heap from start-to-finish of heap leaching. In this configuration the lixiviant is applied for at least about 0.5 and even more preferably from about 1 to 48 months from the start to finish of heal leaching.
In one configuration, the lixiviant is sparged with the oxygen-containing gas before application to the heap or the gas is otherwise contacted with the lixiviant before application to the heap (such as by in-line mixing) to cause the lixiviant to have a heightened dissolved molecular oxygen content. Additional gas may be deliberately introduced into the heap separately from the lixiviant, if desired. The dissolved molecular oxygen content of the lixiviant preferably is at least about 1 mg/L and even more preferably ranges from about 5 to about 40 mg/L.
After construction of the 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 while the oxygen-containing gas is introduced to the bottom of the heap. The pregnant leach solution 408 is collected from the base of the heap. The pregnant leach solution is recycled to the top of the heap. The recycle rate is sufficient to provide an application rate of the lixiviant to the top of the heap of from about 2 and more preferably from about 4 to about 24 LmJmz of top surface area for the heap.
During recycle, at least a portion (typically at least about S vol.% and more typically from about 50 to about 100 vol.%) of the pregnant leach solution is removed and subjected to precious metal recovery, such as by any of the techniques referred to above. A particularly preferred technique is to recover the dissolved precious metal by sulfide precipitation. As noted, sulfide precipitation not only precipitates the precious metal but also regenerates the thiosulfate by converting polythionates into thiosulfate. Sulfide precipitation is typically conducted under anerobic conditions, as noted above. Such conditions can be realized and maintained by de-aerating the pregnant leach solution with a vacuum, inert gas bubbling, and/or allowing the pregnant leach solution to stand dormant for a sufficient period of time for the dissolved oxygen level to decrease to desired levels. Sufficient sulfide is added to the pregnant leach solution to precipitate at least most of the dissolved precious metals) as precious metal sulfides and to convert at least most of the polythionates to thiosulfate. The barren thiosulfate lixiviant, after precious metal recovery, can be recombined with the remaining pregnant leach solution and returned to the top of the heap.
While not wishing to be bound by any theory, it is believed that the precious metal thiosulfate complex may be unstable under certain conditions, including those set forth above, and that the precious metal can be stripped from the thiosulfate-containing solution by a number of substances commonly encountered in precious metal-containing materials.
The substances or preg robbing materials typically absorb, adsorb or precipitate the precious metal. Such preg-robbing materials include carbonaceous materials, pyrite-containing materials, chacopyrite and iron oxides. Surprisingly and unexpectedly, blinding agents may be used in the thiosulfate li5ziviant to prevent or uihibit preg robbing of the precious metal by the preg robbing material. The blinding agent itself absorbs or adsorbs (in preference to the precious metal) or other~rise neutralizes (such as by chemical reaction) the preg robbing sites on the material. Suitable blinding agents include one or more of hydrocarban-containing substances, such as petroleumproducts (e.g., kerosene, diesel fuel, and gasoline), alcohols, esters, or aldehydes; surfactants such as detergents, sodium lauryl sulfonate, or organic phosphates; and reactive metal salts such as lead, mercury, cadmium, tin, and silver salts. In such situations, the thiosulfate lixiwiant, in any of the leaching processes discussed above, typically inchxdes at least about 0.1 mg/1.,, more typically at least about 1 mg/L, and even more typically from about 2 to about 200 mg/L of the blinding agent.
'~fhile not wishing to be bound by any theory, other agents xmay also be suitable as blinding agents. Preferably, the agents do not destabilize thiosulfate in solution. Agents which act as oxidation catalysts, can destabilize thiosulfate in solution. By gray of example, copper salts under certain conditions are not preferred as a blinding agent as copper salts under these conditions can catalyze thiosulfate decomposition.
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 nwre of bioo~cidation or chemical oxidation to oxidize sulfides, ultrafxne grinding to liberate occluded precious metals, conventional roasting to destro~r carbon- and/or sulfide-containing minerals, and/or microwave roasting.
EX.A_MP'LE 1 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 m,axirnurn of 83% of the contained gold was capable of being solubili2ed while the remaining gold was inaccessible to a lrxi~riant because it was encapsulated withili 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 naturalpH ofthe slurry at ambient temperature was 8.3. The p~T
ofthe shrny 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 a hours in order to monitor gold extraction with time. Upon termination of leaching, the slurry was tittered 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 fo Mows:
Leach Temp. Leach TimeCalc'd HeadResidue 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 40 2 51.2 4 5s. ~
6 26.4 9.25 64.9 60 2 63.7 4 6$.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 increasing temperature and leach time in the temperature range 20-60°C.
The best results i 0 were obtained at 60°C, with about 81 % gold extraction obtained after 6 hours leaching, this representing about 98% of the teachable gold content of the ore.
1:~AMPLE 2 A second gold ore from Nevada, designated Sample B, was subjected to 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 afthe 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 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 B hours for the pH
11 and 12 tests.
The results were as follows:
Tnitial Leach Time Calc'd Residue Au Ext'n Head pH (hours) Au (g/t) Au (g/t) (%) 9 1 50.2 2 62.4 4 72.0 6 $.49 2.10 7s.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 leachimg). Hovu'ever, gold leaching was suppressed during the first 4 hours of leaching at pH
12, but then started to recorder.
A third gold ore sample from Nevada, Sample C, was subjected to thiosul~ate leaching under oxygen pressure at varying temperatures. The head analysis of the ore was as I S follo~crs:
Gold Ore Sample C
Au, g/t 9.50 C (t), % 4.45 1~e, % 2.52 C (C03), % 3.12 Cu, pprn 40 C (org), % 1.38 As, ppm 647 S (2~), % 0.35 Tag, pprn 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 Vim). Samples of the ore were slurried with water to a pulp density of 33% solids in a mechanically agitated laboratory autoclave. The initial pI~ 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 slurrywas 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 far 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 fcltered and the residue solids were washed with a dilute thivsulfate 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 G. It can be seen that the gold leached quickly at 60°C and 80°C, there being little difference inthe extractionrate 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 Teach 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 ExaMPLE a The gold ore designated Sample C was subjected to thiosuLfate leaching at varying o~rygen pressures. Samples of the ore were prepared and leached as described in Exanxple 3 except the temperature Was maintained at 60°C in each test and the oxygen partial pressuxe was varied.
Fig. 6 portrays the efr'ect 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 maixitained with pure oxygen at atmospheric pressure).
It can be seen that the rate of gold leaching was somewhat sensiti~re to oxygen pressuxe, in that the rate increased with increasing pressure, particularly during the first tvcro hours of leaching. After 6 hours leaching, gold extraction varied from a low of 78% at 0 psig OZ to a high of 83% at 200 psig 02.
An overall sum~nary~ of the results is provided below:
Parameter Test #7 Test #25 Test #22 Test #28 Test #31 200 psig 100 prig 50 psig 10 psig 0 psig Oz 02 02 Oz O
Leach tune, hours 8 6 6 6 6 Final pH NA 8.7 9.0 9.3 9.4 2~ Final ORP, tnV' (SHE)NA 242 223 216 232 Calc'd Head Au, g/t 9.78 9.43 9.40 8.95 9.08 Residue Au, glt 1,68 1.63 1.7? 1.72 2.00 Au Ext'n, % 82.8 82.7 81.1 $0.8 78.0 NA = not analysed EXAMPLE S
The gold ore designated Sample C was subjected to thiosulfate leaching under oxygen pressure at varying initial sodium thiosulfate concentrations. Samples of the ore were prepared and leached as described in Example 3 except the temperature was maintained at 60°C in each test and the initial sodium thiosulfate concentration was varied.
Fig. 7 portrays the effect ofinitial sodium thiosulfate concentration, in the range 0.05-0.2 M, on the rate of gold extraction from Sample C. It can be seen that the rate of gold IO leaching was insensitive to initial thiosulfate concentration in the 0.1-0.2 M range. At 0.05 M thiosulfate, the rate was reduced significantly, particularly during the first two hours of leaching. After 6 hours leaching gold extraction was 78% at 0.05 M thiosulfate compared to 82% achieved at both 0.1 M and 0.2 M thiosulfate concentration.
An overall summary of the results is pro~rided below:
Parameter Test #4 Test #25 Test #8 0.2 M 0.1 M 0,05 M
Leach time, hours 8 6 6 Final pH 8.'7 8,7 8.5 Final 4RP, mV (SHE) NA 242 262 Calc'd Head Au, get 8.$S 9.43 9.40 Residue Au, glt 1.50 1.63 1.87 A~ Ext~, % 83.o 8z.7 so.l IvTA --- no t anatyaed The gold ore designated S ample C eras subjected to thiosulfate leaching under oxygen pressure at two different pulp densities. Samples of the ore were prepared and leached as described in E~xnple 3, except the temperature was maintained at 60°C
in each test and the leach pulp density was either 33% or 45°~~o solids by weight, Fig. 8 portrays the effect of33% vs. 45% pulp density on the rate ofgold 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, pit 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 oxygenpressure at the naturalpH ofthe shuzy, for 8 hours. The head analysis of the ore was as follows:
Grald Ure Sample D
Au, g/t 12.15 C (t), % 4.31 Fe, % 2.09 C (C03), % 3.02 Cu, ppm 39 C (org), % 1.30 As, ppm 692 S (2-), % 0.12 Tag, ppm 27 S (t), % 0.22 Ca, % 8.9 Mg, % 1.3 From a diagnostic leaching analysis of the ore it was deternvned that 80% of the contained gold was capable of being solubi)ized.
The ore was ground to 80% passing 200 mesh {75 Vim). A sample of the ore was slurried with water to a pulp density of 40% solids in a mechanically agitated iaboratory autoclave. The autoclave was sealed and pressurized to 100 psig oxygen with pure (95%
plus) oxygen gas and the slurrywas heated to 60bC. To initiate leaching, a quantity of sodium thiosulfate stock solution eras 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 thiosulfate solution. The residue solids and leach solution ~srere assayed for gold to determine the final gold extraction.
Fig. 9 depicts pexcent gold extraction and percent remaining thiosulfate with time.
Crold extraction reached 79.3% after 8 hours, or 99% of the teachable gold content.
Thiosul,fate consumption was low, with $6.7% of the thiosulfate remaining after 8 hours and available for reuse, An overall summary of the results is pro ided below:
Parameter Test #37-01 Leach time, hours $
lnitiai 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 2.3$
Gold extraction, % 79.3 Amount of thiosulfate remaining, % 86.7 E~.A.MPLE $
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 WFthout precipitating gold. The slurry was kept under a nitrogen atmosphere to ensure the S dissolved oxygen content was maintained below 0.2 mg/I~. A single dose of a 0.26 M sodium.
bisulfide (I~aI~S) solution, adjusted to pH 9, was added and the pretreatment was allowed to proceed at 60°C and ambient pressure for 2 hours. The amount of sulfide added was 150%
of stoichiometric based on the amount required to convert the pol~ythionates back to thiosulfate in accordance with the following reactions:
2 ~4 06 J ~I- S 2 -~ ~ 112 ~ ~ ~,; S2 ~3 + 3 ~+
A summary of the results is pxovided below:
TFme Au 5203' S4O6~ S30b~' ORP pH
{~) {-) ~P~) ~~-) ~BIL) ~m~) 0 4.3f> 8.35 0,51 0.59 240 6.9 120 4.36 11.0 0.06 O.IO 5 6.7 The tetrathionate and trithionate concentrations were reduced by 88% and 83%
respectively while all of the gold remained in solution.
EX,.AMPLE 9 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 (N'a2S) solution was added and the treatment was allowed to proceed for 2 hours at ambient temperature (22°C) and pressure. The amount of suLfiide added was 100%
of stoichiometric based on the amount required to convert the polytluonates back to thiosulfate in accordance with the following reactions:
Zs4o6- + sz- + /Hzo-~ r2s2o3- + 3H+
s3 ob + s u' -~ 2s.z o3 A summary of the results is provided below:
Tune Au S2O3z Sa06~ S3O6~ ORl' p~T
() WL) W~) WL) W) ~m~) 0 4.12 7,R 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 IO.a 0.01 9.9 0.01 0.01 -230 10.2 15 60 0.01 9.8 0.01 0.01 -260 10.3 90 0.01 10.1 0.01 0.01 -2$0 10.3 120 0.01 10.2 0.01 0.01 -260 10.3 The rate of con~rersion of polythionates to thiosulfate was extremely fast under 20 ambient conditions, with essentially complete conversion achieved after about 10 minutes.
Similarly, gold precipitation was also fast and essentially complete after about 30 minutes.
A gold ore from Nevada, designatf:d sample ?~, was subjected to thiosulfate leaching in a column. The ore assayed 2.48 g/t gold, 2.2 wt.% iron, 0.2 wt.% total sulfur, 0.13 ~cxrt.%
25 sulfide sulfur, 6.1 wt.% total carbon and 1.1 wt.% organic carbon. The ore was crushed to -47~
100% passing 1 inch and then loaded into a 10 inch diameter, 8 foot tall column constructed from F'V'C fitted with a retaining plate to keep the ore in the column.
A solution containing 15 g/L ammonium thiosulfate at a pH of 9:0 was applied continuousl~r to the top of the column and was collected at the: bottom. The solution was not recycled. Solution eras applied at a rate of 12 L/h/m2. To the aerated column, air was forced through the bottom of the column at a rate of 130 L/h/m2. No air was forced through the unaerated column. The temperature of both tests averaged 22°C.
'U'pon termination of leaching, the residue solids were washed 'with water and the residue solids and collected leach solution were assayed to determine the Vital gold extraction.
The results are as folio vs:
Leach Aerated Column Unaerated Column 'I~me Au Extraction Au Extraction (~Ys) (m~) (m~) 0.00 0,00 0% 0.00 0%
0.89 1.80 0% 0.00 0%
1.93 5.20 9% 2.38 2%
2.99 3.70 2 7% 1.86 11 4.01 2.20 36% 1.40 17%
5.09 1.30 43% I.02 22%
8.00 0.62 52% 0,64 32%
11.99 0.34 S9% 0.38 39%
14.91 0.24 Ei2% 0.26 43%
18.91 0.20 64% 0.16 46%
21.81 0.16 66% 0.12 48%
25.96 0.16 68% 0.10 50%
28.78 0.10 70% 0.08 51%
33.89 0.08 i'I% 0.06 52%
The results indicate that the rate and extent of gold extraction was improved with the addition of forced aeration.
While this invention has been described in conjunction with the specific embodiments thereof, it is evident that nnany alternatives, modifications, and variations will be apparent to those skilled in the art. Accordingly, preferred embodiments of the 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 tt>iosulfate. 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 I08, between grinding 108 and thickening 116, in the thickener 1 I6 immediately before or during, leaching 132 and/or between resin in pulp 164 and CCD 1'~2. 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 ox 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.
6 26.4 9.25 64.9 60 2 63.7 4 6$.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 increasing temperature and leach time in the temperature range 20-60°C.
The best results i 0 were obtained at 60°C, with about 81 % gold extraction obtained after 6 hours leaching, this representing about 98% of the teachable gold content of the ore.
1:~AMPLE 2 A second gold ore from Nevada, designated Sample B, was subjected to 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 afthe 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 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 B hours for the pH
11 and 12 tests.
The results were as follows:
Tnitial Leach Time Calc'd Residue Au Ext'n Head pH (hours) Au (g/t) Au (g/t) (%) 9 1 50.2 2 62.4 4 72.0 6 $.49 2.10 7s.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 leachimg). Hovu'ever, gold leaching was suppressed during the first 4 hours of leaching at pH
12, but then started to recorder.
A third gold ore sample from Nevada, Sample C, was subjected to thiosul~ate leaching under oxygen pressure at varying temperatures. The head analysis of the ore was as I S follo~crs:
Gold Ore Sample C
Au, g/t 9.50 C (t), % 4.45 1~e, % 2.52 C (C03), % 3.12 Cu, pprn 40 C (org), % 1.38 As, ppm 647 S (2~), % 0.35 Tag, pprn 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 Vim). Samples of the ore were slurried with water to a pulp density of 33% solids in a mechanically agitated laboratory autoclave. The initial pI~ 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 slurrywas 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 far 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 fcltered and the residue solids were washed with a dilute thivsulfate 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 G. It can be seen that the gold leached quickly at 60°C and 80°C, there being little difference inthe extractionrate 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 Teach 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 ExaMPLE a The gold ore designated Sample C was subjected to thiosuLfate leaching at varying o~rygen pressures. Samples of the ore were prepared and leached as described in Exanxple 3 except the temperature Was maintained at 60°C in each test and the oxygen partial pressuxe was varied.
Fig. 6 portrays the efr'ect 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 maixitained with pure oxygen at atmospheric pressure).
It can be seen that the rate of gold leaching was somewhat sensiti~re to oxygen pressuxe, in that the rate increased with increasing pressure, particularly during the first tvcro hours of leaching. After 6 hours leaching, gold extraction varied from a low of 78% at 0 psig OZ to a high of 83% at 200 psig 02.
An overall sum~nary~ of the results is provided below:
Parameter Test #7 Test #25 Test #22 Test #28 Test #31 200 psig 100 prig 50 psig 10 psig 0 psig Oz 02 02 Oz O
Leach tune, hours 8 6 6 6 6 Final pH NA 8.7 9.0 9.3 9.4 2~ Final ORP, tnV' (SHE)NA 242 223 216 232 Calc'd Head Au, g/t 9.78 9.43 9.40 8.95 9.08 Residue Au, glt 1,68 1.63 1.7? 1.72 2.00 Au Ext'n, % 82.8 82.7 81.1 $0.8 78.0 NA = not analysed EXAMPLE S
The gold ore designated Sample C was subjected to thiosulfate leaching under oxygen pressure at varying initial sodium thiosulfate concentrations. Samples of the ore were prepared and leached as described in Example 3 except the temperature was maintained at 60°C in each test and the initial sodium thiosulfate concentration was varied.
Fig. 7 portrays the effect ofinitial sodium thiosulfate concentration, in the range 0.05-0.2 M, on the rate of gold extraction from Sample C. It can be seen that the rate of gold IO leaching was insensitive to initial thiosulfate concentration in the 0.1-0.2 M range. At 0.05 M thiosulfate, the rate was reduced significantly, particularly during the first two hours of leaching. After 6 hours leaching gold extraction was 78% at 0.05 M thiosulfate compared to 82% achieved at both 0.1 M and 0.2 M thiosulfate concentration.
An overall summary of the results is pro~rided below:
Parameter Test #4 Test #25 Test #8 0.2 M 0.1 M 0,05 M
Leach time, hours 8 6 6 Final pH 8.'7 8,7 8.5 Final 4RP, mV (SHE) NA 242 262 Calc'd Head Au, get 8.$S 9.43 9.40 Residue Au, glt 1.50 1.63 1.87 A~ Ext~, % 83.o 8z.7 so.l IvTA --- no t anatyaed The gold ore designated S ample C eras subjected to thiosulfate leaching under oxygen pressure at two different pulp densities. Samples of the ore were prepared and leached as described in E~xnple 3, except the temperature was maintained at 60°C
in each test and the leach pulp density was either 33% or 45°~~o solids by weight, Fig. 8 portrays the effect of33% vs. 45% pulp density on the rate ofgold 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, pit 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 oxygenpressure at the naturalpH ofthe shuzy, for 8 hours. The head analysis of the ore was as follows:
Grald Ure Sample D
Au, g/t 12.15 C (t), % 4.31 Fe, % 2.09 C (C03), % 3.02 Cu, ppm 39 C (org), % 1.30 As, ppm 692 S (2-), % 0.12 Tag, ppm 27 S (t), % 0.22 Ca, % 8.9 Mg, % 1.3 From a diagnostic leaching analysis of the ore it was deternvned that 80% of the contained gold was capable of being solubi)ized.
The ore was ground to 80% passing 200 mesh {75 Vim). A sample of the ore was slurried with water to a pulp density of 40% solids in a mechanically agitated iaboratory autoclave. The autoclave was sealed and pressurized to 100 psig oxygen with pure (95%
plus) oxygen gas and the slurrywas heated to 60bC. To initiate leaching, a quantity of sodium thiosulfate stock solution eras 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 thiosulfate solution. The residue solids and leach solution ~srere assayed for gold to determine the final gold extraction.
Fig. 9 depicts pexcent gold extraction and percent remaining thiosulfate with time.
Crold extraction reached 79.3% after 8 hours, or 99% of the teachable gold content.
Thiosul,fate consumption was low, with $6.7% of the thiosulfate remaining after 8 hours and available for reuse, An overall summary of the results is pro ided below:
Parameter Test #37-01 Leach time, hours $
lnitiai 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 2.3$
Gold extraction, % 79.3 Amount of thiosulfate remaining, % 86.7 E~.A.MPLE $
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 WFthout precipitating gold. The slurry was kept under a nitrogen atmosphere to ensure the S dissolved oxygen content was maintained below 0.2 mg/I~. A single dose of a 0.26 M sodium.
bisulfide (I~aI~S) solution, adjusted to pH 9, was added and the pretreatment was allowed to proceed at 60°C and ambient pressure for 2 hours. The amount of sulfide added was 150%
of stoichiometric based on the amount required to convert the pol~ythionates back to thiosulfate in accordance with the following reactions:
2 ~4 06 J ~I- S 2 -~ ~ 112 ~ ~ ~,; S2 ~3 + 3 ~+
A summary of the results is pxovided below:
TFme Au 5203' S4O6~ S30b~' ORP pH
{~) {-) ~P~) ~~-) ~BIL) ~m~) 0 4.3f> 8.35 0,51 0.59 240 6.9 120 4.36 11.0 0.06 O.IO 5 6.7 The tetrathionate and trithionate concentrations were reduced by 88% and 83%
respectively while all of the gold remained in solution.
EX,.AMPLE 9 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 (N'a2S) solution was added and the treatment was allowed to proceed for 2 hours at ambient temperature (22°C) and pressure. The amount of suLfiide added was 100%
of stoichiometric based on the amount required to convert the polytluonates back to thiosulfate in accordance with the following reactions:
Zs4o6- + sz- + /Hzo-~ r2s2o3- + 3H+
s3 ob + s u' -~ 2s.z o3 A summary of the results is provided below:
Tune Au S2O3z Sa06~ S3O6~ ORl' p~T
() WL) W~) WL) W) ~m~) 0 4.12 7,R 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 IO.a 0.01 9.9 0.01 0.01 -230 10.2 15 60 0.01 9.8 0.01 0.01 -260 10.3 90 0.01 10.1 0.01 0.01 -2$0 10.3 120 0.01 10.2 0.01 0.01 -260 10.3 The rate of con~rersion of polythionates to thiosulfate was extremely fast under 20 ambient conditions, with essentially complete conversion achieved after about 10 minutes.
Similarly, gold precipitation was also fast and essentially complete after about 30 minutes.
A gold ore from Nevada, designatf:d sample ?~, was subjected to thiosulfate leaching in a column. The ore assayed 2.48 g/t gold, 2.2 wt.% iron, 0.2 wt.% total sulfur, 0.13 ~cxrt.%
25 sulfide sulfur, 6.1 wt.% total carbon and 1.1 wt.% organic carbon. The ore was crushed to -47~
100% passing 1 inch and then loaded into a 10 inch diameter, 8 foot tall column constructed from F'V'C fitted with a retaining plate to keep the ore in the column.
A solution containing 15 g/L ammonium thiosulfate at a pH of 9:0 was applied continuousl~r to the top of the column and was collected at the: bottom. The solution was not recycled. Solution eras applied at a rate of 12 L/h/m2. To the aerated column, air was forced through the bottom of the column at a rate of 130 L/h/m2. No air was forced through the unaerated column. The temperature of both tests averaged 22°C.
'U'pon termination of leaching, the residue solids were washed 'with water and the residue solids and collected leach solution were assayed to determine the Vital gold extraction.
The results are as folio vs:
Leach Aerated Column Unaerated Column 'I~me Au Extraction Au Extraction (~Ys) (m~) (m~) 0.00 0,00 0% 0.00 0%
0.89 1.80 0% 0.00 0%
1.93 5.20 9% 2.38 2%
2.99 3.70 2 7% 1.86 11 4.01 2.20 36% 1.40 17%
5.09 1.30 43% I.02 22%
8.00 0.62 52% 0,64 32%
11.99 0.34 S9% 0.38 39%
14.91 0.24 Ei2% 0.26 43%
18.91 0.20 64% 0.16 46%
21.81 0.16 66% 0.12 48%
25.96 0.16 68% 0.10 50%
28.78 0.10 70% 0.08 51%
33.89 0.08 i'I% 0.06 52%
The results indicate that the rate and extent of gold extraction was improved with the addition of forced aeration.
While this invention has been described in conjunction with the specific embodiments thereof, it is evident that nnany alternatives, modifications, and variations will be apparent to those skilled in the art. Accordingly, preferred embodiments of the 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 tt>iosulfate. 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 I08, between grinding 108 and thickening 116, in the thickener 1 I6 immediately before or during, leaching 132 and/or between resin in pulp 164 and CCD 1'~2. 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 ox 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 (12)
1. A process for recovering a precious metal from a precious metal-containing material, comprising:
(a) providing a heap of the precious metal-containing material; and (b) passing a thiosulfate lixiviant and molecular oxygen through the heap to form a pregnant leach solution comprising dissolved precious metals, wherein the molecular oxygen is at a pressure greater than ambient atmospheric pressure before introduction into the heap.
(a) providing a heap of the precious metal-containing material; and (b) passing a thiosulfate lixiviant and molecular oxygen through the heap to form a pregnant leach solution comprising dissolved precious metals, wherein the molecular oxygen is at a pressure greater than ambient atmospheric pressure before introduction into the heap.
2. The process of Claim 1, wherein the molecular oxygen is in the form of a gas and the thiosulfate lixiviant and molecuar oxygen flow countercurrently through the heap.
3. The process of Claim 1, wherein the molecular oxygen is introduced under pressure into the heap by a network of conduits positioned in the base of the heap.
4. The process of Claim 1, wherein the thiosulfate lixiviant has a pH of no more than about 9 before introduction to the heap.
5. The process of Claim 1, wherein the thiosulfate lixiviant has a free ammonia content of no more than about 2,000 ppm.
6. The process of Claim 1, wherein at least about 0.5 kg of molecular oxygen/ton of heap material is introduced into the heap during leaching.
7. The process of Claim 1, wherein a dissolved molecular oxygen content of the lixiviant is at least about 1 mg/L.
8. 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-containing lixiviant, wherein the lixiviant contains a blinding agent.
contacting the carbonaceous precious metal-containing material with a thiosulfate-containing lixiviant, wherein the lixiviant contains a blinding agent.
9. The process of Claim 8, wherein the blinding agent includes one or more of hydrocarbons, alcohols, esters, aldehydes, surfactants, lauryl sulfonates, phosphates; and metal salts.
10. The process of Claim 8, wherein the thiosulfate lixiviant includes at least about 0.1 mg/L blinding agent.
11. The process of Claim 8, wherein the thiosutfate lixiviant includes at least about 1 mg/L blinding agent.
12. The process of Claim 8, wherein the thiosulfate lixiviant includes from about 2 to about 200 mg/L blinding agent.
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CA002412352A CA2412352A1 (en) | 2002-11-18 | 2002-11-21 | Method for thiosulfate leaching of precious metal-containing materials |
CA2424714A CA2424714C (en) | 2002-11-15 | 2003-04-01 | Method for thiosulfate leaching of precious metal-containing materials |
CA2850217A CA2850217A1 (en) | 2002-11-15 | 2003-11-14 | Method for thiosulfate leaching of precious metal-containing materials |
CA2850216A CA2850216A1 (en) | 2002-11-15 | 2003-11-14 | Method for thiosulfate leaching of precious metal-containing materials |
CA2691607A CA2691607C (en) | 2002-11-15 | 2003-11-14 | Method for thiosulfate leaching of precious metal-containing materials |
CA2449467A CA2449467C (en) | 2002-11-15 | 2003-11-14 | Method for thiosulfate leaching of precious metal-containing materials |
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WO2012141607A1 (en) | 2011-03-09 | 2012-10-18 | Malusel Viorel Alexandru | Gold and silver extraction technology |
EP2721185A4 (en) * | 2011-06-15 | 2015-06-17 | Barrick Gold Corp | Method for recovering precious metals and copper from leach solutions |
US10161016B2 (en) | 2013-05-29 | 2018-12-25 | Barrick Gold Corporation | Method for pre-treatment of gold-bearing oxide ores |
US10415116B2 (en) | 2010-12-07 | 2019-09-17 | Barrick Gold Corporation | Co-current and counter current resin-in-leach in gold leaching processes |
US11639540B2 (en) | 2019-01-21 | 2023-05-02 | Barrick Gold Corporation | Method for carbon-catalysed thiosulfate leaching of gold-bearing materials |
CN116858991A (en) * | 2023-09-04 | 2023-10-10 | 济宁华晟服装股份有限公司 | Cotton desizing treatment monitoring method |
-
2002
- 2002-11-21 CA CA002412352A patent/CA2412352A1/en not_active Abandoned
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