AU3489995A - Method and apparatus for extracting precious metals from their ores and the product thereof - Google Patents

Description

METHOD AND APPARATUS FOR EXTRACTING PRECIOUS METALS FROM THEIR ORES AND THE PRODUCT THEREOF

STATEMENT AS TO RIGHTS IN INVENTIONS

MADE UNDER FEDERALLY-SPONSORED

RESEARCH AND DEVELOPMENT

This invention was made with U.S. Government support under Small Business Innovation Research Grant No. DMI-9461234 which was awarded by the National Science Foundation (NSF), an independent agency of the U.S. Government. The U.S. Government has certain rights in the invention.

TECHNICAL FIELD

This invention relates to a method and apparatus for extracting precious metals from their ores and the product thereof. In particular, it relates to the following: (1) a biohydrometallurgical process and apparatus for extraction and recovery of gold, silver and platinum group elements from their ores; (2) the products of that process and apparatus.

BACKGROUND ART

Development of cost-effective techniques for extraction of precious metals from their ores has been a goal of metallurgists for hundreds of years. In recent years, the addition of environmental costs to the total cost of products of precious metal miners has encouraged a search for environmentally acceptable options, especially for refractory ores.

In the prior art the first step in precious metal production from ore involves preparing the ore for precious metal extraction. Preparation can take any one of a number of courses depending on the character of the ore. Gold and silver ores often contain metallic sulfides. Ores containing platinum-group elements (PGE) typically also contain metallic sulfides. For example, in the Bushveld Igneous Complex in South Africa, platinum group element values occur in solid solution in the base-metal sulfides pyrrhotite (Fe^S), pentlandite (Fe,Ni)₉S₈, pyrite (FeS₂), and as the discrete platinoid metal minerals cooperite (PtS), laurite (RuS₂), braggite (Pt,Pd,Ni)S and Pt-Fe alloys or their intergrowths (see Guilbert, J.M. & Park, C.F., Jr. "The geology of ore deposits." New York: W.H. Freeman and Company, 1986). The Stillwater Complex in Montana is a similar deposit. Typically, refractory, non-oxidized (e.g., sulfide) gold and silver ores (especially those with a relatively high carbon content) are oxidized at elevated temperatures and pressures in large autoclaves (i.e., "roasted"), prior to the extraction of precious metals by means of cyanide leaching, (see McQuiston, Jr., F.W., & Shoemaker, R.S., Gold and Silver Cyanidation Plant Practice, Vol. II, Baltimore: Port City Press, 1980).

During the last decade, heap leach processes for cost-effective bio-oxidation of pyritic and arsenopyritic sulfides in gold and silver ores have been developed to the point of commercial application (see Torma, A.E., Biotechnology: A Comprehensive Treatise in 8 Volumes, Deerfield Beach. Florida: Verlag Chemie, 1981). Recent improvements in the art are disclosed by: Pooley et al. in U.S. Patent No. 4,822,413, April 18, 1989; Hackl et al. in U.S. Patent No. 4,987,081, January 22, 1991; Hunter in U.S. Patent No. 5,076,927, December 31, 1991; Brierly et al. In U.S. Patent No. 5,127,942, July 7, 1992; and Brierly and Hill in U.S. Patent No. 5,246,486, September 21, 1993. The disclosures in the aforementioned patents are incorporated by reference herein as if fully set forth. A great variety of precious metal extraction processes have also been developed (see Gupta, C.K., & Mukherjee, T.K., Hydrometallurgy in Extraction Processes, Vol. I, Boston: CRC Press, 1990). Precious metal extraction processes are disclosed by: Pesic in U.S. Patent No. 4,778,519, October 18,. 1988; Ball et al. in U.S. Patent No. 4,902,345, February 20, 1990; and Kandemir in UK Patent No. 2,180,829, published April 8, 1987. The disclosures of the aforementioned patents are incorporated by reference herein as if fully set forth. The relatively low economic cost of cyanidation, however, has ensured its proliferation.

State-of-the-art precious metal heap leach practice varies with the nature of the ore. Biooxidation process steps may include ore crushing, acid pretreatment, inoculation with appropriate sulfide-oxidizing bacteria, addition of nutrients, recirculating the biolixiviant and cooling the heap (for 3 to 8 days), and allowing the heap to "rest" (for 3 to 8 days). Precious metal extraction by means of cyanidation may include the process steps of washing the heap for an extended period (e.g., 14 days) to remove residual acidity or iron content, breaking the heap apart in order to agglomerate it with cement and/or lime to make a new heap, leaching it with an alkaline cyanide or thiosulfate solution for 30 to 40 days, and recovery of gold and silver from the leach solution by adsorption on activated carbon or zinc dust precipitation.

A significant amount of work in the field of bio-oxidation and metals extraction has been accomplished by a variety of investigators. Tomizuka, N. & Yagisawa, M., in "Optimum conditions for leaching of uranium and oxidation of lead sulfide with thiobacillus ferrooxidans and recovery of metals from bacterial leaching solution with sulfated-reducing bacteria," (in

Metallurgical Applications of Bacterial Leaching and Related Microbiological Phenomena, Murr, L.E., Torma, A.E., & Brierly, J.A. (Eds.) New York: Academic Press, 1978), describe a two- step process for leaching of uranium and oxidation of lead sulfide where recovery of metals is accomplished by means of microbial sulfate reduction. Alper, J., in "Bacterial methods may strike it rich in refining metals, cleaning coal," (High Technology, April, 1984, pp. 32-35), describes the bio-oxidation of gold-bearing arsenopyrite/pyrite and notes that production of large amounts of arsenic and sulfurous gases is avoided. Torma, A.E., in Biotechnology: A Comprehensive Treatise in 8 Volumes, (Deerfield Beach, FI: Verlag Chemie, 1988), reviewed bioleaching processes. Livesay-Goldblatt, E., in Fundamental and Applied Biohydrometallurgy, (Proc. 6th International Symposium on Biohydrometallurgy, Vancouver, B.C. 89-96, 1986), described a process for gold recovery from arsenopyrite/pyrite ore by bacterial leaching and cyanidation.

Torma, A.E., in "Biotechnology: A comprehensive treatise in 8 volumes." (Deerfield Beach, FL: Verlag Chemie, 1988), reviews bio-oxidation of gold and silver ores. Hackl, R.P., Wright, F., & Bruynesteyn, A., in Proceedings of the Third Annual General Meeting of Biominet, (August 20- 21, 71-90, 1986), described development of the BIOTANKLEACH process for leaching pyritic materials from gold and silver ore. The results of bench-scale and pilot-scale evaluations were presented. Marchant, P.B., & Lawrence, R.W., in "Flowsheet design, process control, and operating strategies in the bio-oxidation of refractory gold ores," (Proceedings of the Third Annual General Meeting of Biominet, August 20-21, 39-51, 1986), listed considerations in the design of commercial bio-oxidation plants. Lawrence R.W., in "Biotreatment of Gold,"

(Microbial Mineral Recovery New York: McGraw-Hill edited by Ehrlich, H.L. and Brierly, CL,

1990), discussed biotreatment of gold ore. The benefits of using the BacTech moderately thermophilic cultures in bio-oxidation processes were discussed by Budden, J.R., & Spencer, P. A. in "Tolerance to temperature and water quality for bacterial oxidation: The benefits of BacTech's moderately thermophilic culture." (FEMS Microbiology Reviews, 11, 191-196, 1993). Chapman, J.T., Marchant, P.B., Lawrence, R.W., & Knopp, R., in "Biooxidation of a refractory gold bearing high arsenic sulphide concentrate: A pilot study," (FEMS Microbiology Reviews. 11, 243-252, 1993), described a modular mobile bioleach pilot plant for bio-oxidation of a refractory gold-bearing high-arsenic sulfide concentrate. Moffat, A.S., in "Microbial mining boosts the environment," (Science, 264, 778-779, 1994), disclosed how bio-oxidation can increase the efficiency of mining.

Thermophilic versus mesophilic bioleaching process performance was evaluated by Duarte, J.C., Estrada, P.C., Pereira, P.C., & Beaumont, H.P. in " FEMS Microbiology Reviews, 11, 97-102, 1993. Two years of BIOX bio-oxidation pilot plant data were analyzed by Hansford, G.S., & Miller, D.M. in "Biooxidation of a gold-bearing pyrite-arsenophyrite concentrate,"

FEMS Microbiology Reviews, 11, 175-182, 1993. Hoffman, W., Katsikaros, N., & Davis, G. , in "Design of a reactor bioleach process for refractory gold treatment," (FEMS Microbiology Reviews, 11, 221-230, 1994), described the design of a reactor bioleach process for refractory gold treatment. Liu, X., Petersson, S., & Sandstrom, A., in "Evaluation of process variables in bench-scale bio-oxidation of the Olympias concentrate," (FEMS Microbiology Reviews, 11, 207- 214, 1993), presented an evaluation of the effects of process variables on pyrite/arsenopyrite oxidation and gold extraction. Maturana, H., Lagos, U. , Flores, V., Gaeta, M., Cornejo, L., & Wiertz, J.V., in "Integrated biological process for the treatment of a Chilean complex gold ore," (FEMS Microbiology Reviews, 11, 215-220, 1993), described an integrated biological process for treatment of a complex gold ore. Mineral sulfide oxidation by enrichment cultures of a novel thermoacidophilic bacteria were described by Norris, P. R. & Owen, J.P. in " Mineral sulphide oxidation by enrichment cultures of novel thermoacidophilic bacteria," FEMS Microbiology Reviews. 11, 51-56, 1993). Rate controls on the bio-oxidation of heaps of pyritic material imposed by bacterial upper temperature limits were described by Pantelis, G. & Ritchie, A.I.M. in "Rate controls on the oxidation of heaps of pyritic material imposed by upper temperature limits on the bacterially catalysed process," (FEMS Microbiology Reviews, 11, 183-190, 1993). Biooxidation bacteria have been characterized in detail. Brierly, C.L., & Brierly, J.A., in "A chemoautotrophic and thermophilic microorganism isolated from an acid hot spring," (Canadian J. Microbiology, 19, 183-188, 1973), characterized a chemoautotrophic and thermophilic (70°C) microorganism isolated from an acid hot spring. De Rosa, M., Gambacorta, A., & Bullock, J.D., in "Extremely thermophilic acidophilic bacteria convergent with Sulfolobus acidocaldarws," (j. General Microbiology, 86, 156-164, 1975), characterized the extremely thermophilic (85°C), acidophilic (pH 1.0) bacteria Sidfolobus acidocaldarius.

While prior art has extensively studied and developed the bio-oxidation process for oxidizing metal sulfides present in gold and silver ore to expose or mobilize precious metal values, little attention has been given to biotechnologies for extracting (solubilizing) and recovering those values. Torma, A.E., in "Biotechnology: A Comprehensive Treatise in 8 Volumes," (Deerfield Beach, FL: Verlag Chemie, 1988), reviewed dissolution of gold by microorganisms. Olson, G.J., in "Microbial oxidation of gold ores and gold bioleaching," (FEMS Microbiology Letters, 119, 1-6, 1994), reviewed microbial oxidation of gold ores and gold bioleaching. He described cyanogenic microorganisms such as Chromobacteri um violaceum, and he noted that gold forms soluble sulfide and polysulfide complexes and suggested that biogenesis of partially- oxidized sulfur compounds may be a mechanism of gold dissolution. The U.S. Bureau of Mines, (Office of Technology Transfer in U.S. Bureau of Mines Cooperative Research Opportunities, Washington, DC: U.S. Bureau of Mines, 1995), disclosed that the solubility of gold in dilute polysulfide solutions at elevated temperatures and pressures and neutral pH levels is comparable to the solubility of gold in cyanide solutions.

Investigators have hypothesized natural processes for gold solubilization involving specific sulfide-oxidizing bacteria such as Thiobacillus ferroxidans and Thiobacillus denitrificans (see Lyalikova, N.N. & Mokeicheva, L.Y. in "The role of bacteria in gold migration in deposits," Microbiology, 38: 805-810, 1969; Kulibakin, V.G, Roslyakow, N.A., Tsimbalist, V.G.,

Mel'nikova, R.D., and Nepeina, L. A., in "Role of sulfur bacteria in supergene migration and concentration of gold," Trans. Inst. Geol. Geofiz. Akad. SSSR Sib. Otd., 370: 75-86, 1977 (in Russian)) and sulfate-reducing bacteria such as Desulfovibrio (see Meyers, W.B. "An hypothesis of the chemical environment of the Rand Goldfield, South Africa," U.S. Geological Survey Open- File Report 1389, 1970; Meyers W.B. "Precambrian pyritic gold-and uranium-bearing

conglomerates," Geological Society of America, Abstracts with Programs, 3: 656-657, 1971; Myers, W.B. ,"Genesis of Uranium and Gold-Bearing Precambrian Quartz-Pebble

Conglomerates," Geological Survey Professional Paper 1161-AA, 1981; and Mossman, D.J. & Dyer, B.D., "The geochemistry of Witwatersrand-type gold deposits and the possible influence of ancient prokaryotic communities on gold dissolution and precipitation," Precambrian Research, 30: 303-319, 1985).

Speculation concerning the impact of biological organisms on the solubilization of gold and silver by bisulfide complexes is based on the findings of those who have studied the formation of precious metal-bisulfide complexes abiotically. Krauskopf, K B., in "The solubility of gold," (Economic Geology, 46, 858-870, 1951), disclosed that "gold may be transported in alkaline sulfide solutions, even in dilute solutions near the neutral point" and "experimentally, one of the most perplexing facts about the chemistry of gold is its ability to dissolve in solutions of HS- of moderate concentration even at room temperature, whereas it dissolves in S^-2 (i.e., more alkaline solutions) only in concentrated solutions at high temperature." Barnes, H.L. in "Geochemistry of hvdrothermal ore deposits," (New York: Hold, Rinehart & Winston, Inc., 1967), disclosed that "gold is known to be soluble in alkaline solutions containing bisulfide or sulfide ior" and that

"when the pH is increased to 7.0 at constant∑S, HS- increases relative to H₂S and the solubilities (of silver) summarized by Anderson rise abruptly by a factor of 20." (See Anderson, G.M., "The solubility of PbS in H₂S-water solutions," Economic Geology, 57, 809-829, 1962). Weissberg, B.G, in "Solubility of gold in hydrothermal alkaline sulfide solutions," (Economic Geology, 65, 551-556, 1970), disclosed that "in less alkaline solutions, where the HS- ion predominates, the experimentally determined solubility of gold ranges from 100 to 200 ppm Au in solutions containing from 0.2 to 0.3 moles NaHS/Kg solution at temperatures between 150 and 250°C..." and that "the present results are in good agreement with results given by Ogryzlo and by Lindner and Gruner and substantiate the high solubility of gold in near neutral pH bisulfide solutions." (See Ogryzlo, S.P., "Hydrothermal experiments with gold," Economic Geology, 30, 400-424, 1935; and Linder, J.L. & Gruner, J.W., "Action of alkali sulphide solutions on minerals at elevated temperatures," Economic Geology, 34, 537-560, 1939). Seward, T.M., in "Thio complexes of gold and the transport of gold in hydrothermal ore solution," (Geochimica et cosmochimica Acta, 37, 379-399, 1973), disclosed that "...an increase in the bisulfide ion concentration at constant pH (or HS-/H₂S ratio) leads to higher gold solubilities" and that "considerable quantities of gold may be transported in hydrothermal ore solutions as thio complexes, particularly in the near neutral pH region where the Au(HS)₂- complex predominates." Seward, T.M., in "The stability of chloride complexes of silver in hydrothermal solutions up to

350°C," (Geochimica et Cosmochimica Acta, 40, 1329-1341, 1976), disclosed that "the solubility data (up to 180°C) of Melent'yev et al. suggest that Ag(HS)₂- will probably be important in near neutral hydrothermal solutions.. ." (See Melent'yev, B.N., Ivanenko, V.V., and Pamfilova, L.A. "Solubility of some ore-forming sulfides under hydrothermal conditions," Rastvorimost' nekotorvkh rudoobrazuyushkikh sul'fidov v gidrotermal' nykh usloviyakh' Moskva, 27-102.

1968). Mountain, B.W. & Wood, S.A., in "Chemical controls on the solubility, transport, and deposition of platinum and palladium in hydrothermal solutions: A thermodynamic approach," (Economic Geology, 83, 492-510, 1988), disclosed that "Westland states that Pt is soluble in alkaline sulfide solutions, possibly a [Pt(HS)_x(OH)_4-x]- complexes. Recent experiments ... have yielded Pt concentrations of about 1 ppm after one month for Pt metal in contact with a 1.0 m Na₂S solution." (See Westland, A.D., "Inorganic chemistry of the platinum-group elements," Canadian Inst. Mining Metallurgy Spec. Vol. 23, 7-18). Gammons, C.H. & Barnes, H.L., in "The solubility of Ag₂S in near-neutral aqueous sulfide solutions at 25 to 300°C," (Geochimica et Cosmochimica Acta, 53, 279-290, 1989), disclosed that "Ag(HS)₂- is the dominant silver species in hydrothermal fluids with near-neutral to alkaline pH, relatively low oxidation state, high total sulfide, and T < 300°C...X Shenberger, D.M. & Barnes, H.L., in "Solubility of gold in aqueous sulfide solutions from 150 to 350°C," (Geochimica et Cosmochimica Acta. 53, 269-278, 1989), disclosed that "the fact that gold is soluble in alkaline sulfide solutions has been known since at least the 17th century..." and that "the high stability of Au(HS)₂- indicates that geologically significant quantities of gold can be transported in typical hydrothermal solutions." Wood, S.A. & Mountain, B.W., in "Thermodynamic constraints on the solubility of platinum and palladium in hydrothermal solutions: Reassessment of hydroxide, bisulfide, and ammonia complexing," (Economic Geology, 84, 2020-2028, 1989), disclosed "... Pt solubilities on the order of 10 to 100 ppb are attainable in bisulfide solutions at alkaline pH at 25 degrees C" and that "Pt and Pd bisulfide complexes show a strong similarity to Ag and Au bisulfide complexes, where the Au(HS)₂- complex predominates over a wide range of conditions." Gammons, C.H., Bloom, M.S., & Yu, Y., in "Experimental investigation of the hydrothermal geochemistry of platinum and palladium: I. Solubility of platinum and palladium sulfide minerals in NaCl/H₂SO₄ solutions at 300°C," (Geochimca et Cosmochimica Acta, 56, 3881-3894, 1992), disclosed "... a broad similarity in the chemical behavior of Au and the PGE elements."

In evaluating the potential for transport of precious metals in natural systems as bisulfide complexes in hydrothermal fluids, investigators have assumed the hydrogen fugacity in their abiotic systems is set by the mineral assemblages through which the fluids would move in nature For example, Weissberg, B.G., in "Solubility of gold in hydrothermal alkaline sulfide solutions," (Economic Geology, 65, 551-556, 1970), disclosed that "... the solubility of gold in natural systems depends on the hydrogen fugacity, which is controlled principally by equilibria between the minerals pyrrhotite, pyrite, magnitite and hematite." Seward, T.M., in "Thio complexes of gold and the transport of gold in hydrothermal ore solutions," (Geochimica et Cosmochimica

Acta, 37, 379-399, 1973), disclosed that "since the dissolution of gold is a function of hydrogen fugacity (see, for example, Raymahashay, B.C. & Holland, H.D., "Redox reactions accompanying hydrothermal wall rock alteration," Economic Geology, 64, 291-305, 1969), a pyrite-pyrrhotite 'redox' buffer was present in all experiments in order that the f_H2 was maintained at a known value."

In the field of biological waste degradation, investigators have long understood and utilized the bioprocessing opportunities presented by "interspecies hydrogen transfer" between hydrogen-producing and hydrogen-consuming anaerobic microorganisms. A

biohydrometallurgical application of this knowledge was disclosed by Hunter, R.M., in

"Biocatalyzed Partial Demineralization of Acidic Metal-Sulfate Solutions," (Ph.D. Thesis,

Montana State University, 1989) and by Hunter, R.M. in "Biocatalyzed Partial Demineralization of Acidic Metal-Sulfate Solutions," (U.S. Patent No. 5,076,927, December 31, 1991), the disclosure of which patent is incorporated herein as if fully set forth. Microbial hydrogen management techniques are disclosed by Harper, S.R. & Pohland, F G. in "Recent developments in hydrogen management during anaerobic biological wastewater treatment," (Biotechnology and Bioengineering, 28, 585-602, 1986). The following reactions illustrate the consumption of hydrogen by acetogens (ACET), methanogens (METH), sulfate-reducing bacteria (SRB) and nitrate-reducing bacteria that produce ammonium (AMM) and nitrogen (NRB):

Acetogens (acetogenic bacteria):

2HCO₃- + 4H₂ + H⁺→ CH₃COO- + 4H₂O ΔG_o ^' = - 104.6 kJ

Methanogens (methanogenic bacteria):

HCO₃- + 4H₂ + H⁺→ CH₄ + 3H₂O ΔG_o ^' = -135.6 kJ Sulfate-reducing bacteria:

SO₄ ^-2 + 4H₂ + H⁺→ HS- + 4H₂O ΔG₀ ^' = - 151.9 kJ

Nitrate-reducing bacteria that produce ammonium:

NO₃- + 4H₂ + 2H⁺→ NH₄- + 3H₂O ΔG_o ^' = -599.6 kJ

Nitrate-reducing bacteria that produce nitrogen gas:

2NO₃- + 5H₂ + 2H⁺→ N₂ + 6H₂O ΔG_o ^' = -1,120.5 kJ

The negative free energies of these reactions at pH 7.0 (ΔG_o ^' ) indicate it is thermodynamically feasible for oxyanion-reducing, hydrogen-consuming bacteria to reduce hydrogen gas fugacities in reactor environments to very low levels under anaerobic conditions.

No single prior art reference or combination of references have suggested combining the knowledge of the above lines of inquiry: the art of solubilization and transport of precious metals in hydrothermal fluids, and the arts of aerobic and anaerobic bioprocessing. The prior art does not teach the use of bio-oxidation to liberate (mobilize) platinum-group elements from their ores. The prior art does not teach the use of sulfate-reducing bacteria to increase the fugacity of hydrogen sulfide gas and the activity of bisulfide ions in a reactor in order to increase the solubility of precious metal-bisulfide complexes. Neither does the prior art teach the use of microorganisms capable of biological reduction of an oxyanion to lower the hydrogen gas fugacity in a reactor in order to increase the solubility of precious metal-bisulfide complexes in the reactor. In fact, the prior art teaches away from the present invention toward aerobic processes for leaching of precious metals from their ores. Such aerobic processes are disclosed in the following recently published books on the subject: Ehrlich, H.L. (1990), Microbial Mineral Recovery, New York: McGraw-Hill; Gupta, C.K., & Mukherjee, T.K. (1990), Hvdrometallurgy in

Extraction Processes, Vols. I and II, Boston: CRC Press; Yannopoulos, J.C. (1991), The Extractive Metallurgy of Gold, New York: Van Nostrand Reinhold; Marsden, J. & House, I. (1993), The Chemistry of Gold Extraction, New York: Ellis Horwood.

DISCLOSURE OF INVENTION

For the purposes of this disclosure, the term "ore" refers to a composition that comprises precious metal values. Thus, ore may be a mineral assemblage that is being mined in-situ (in place) or that has been mined conventionally; or it may be a waste product, such as obsolete or damaged electronic components. The term "precious metals" refers to gold(Au), silver(Ag) and/or platinum-group elements (PGE). The term "platinum-group elements" refers to platinum (Pt), palladium (Pd), rhodium (Rh), ruthenium (Ru), osmium (Rh) and iridium (Ir). The term "bisufide lixiviant" refers to an aqueous solution comprising HS- ions, and may also comprise dissolved H,S gas (H₂S_(aq)). The term "bisulfide complex" refers to a complex comprising a precious metal and bisulfide.

The present invention provides method and apparatus for leaching of precious metals from their ores by means of a leaching solution comprising a sulfide ion and having a low fugacity of hydrogen gas. Leaching is accomplished by formation of precious metal complexes. For example, in gold leaching at neutral pH's, the complex Au(HS)₂- predominates. At a pH above about 10, the solubility of gold is increased by the formation of the complex Au₂S₂ ^-2. Below a pH of about 3, the solubility of gold is increased by the formation of the complex AuHS⁰. Thus, formation of a variety of precious metal-sulfide complexes is possible.

The invention may be practiced on oxidized ore, sulfide ore, or otherwise refractory ore in a tank reactor or heap leach operation. Preferably, a bio-oxidation step for removing base-metal sulfides from precious metal ores is coupled with a bisulfide precious metal leaching step, but conventional roasting may also be used to remove base-metal sulfides and produce an acidic, sulfate stream. Preferably, the leaching solution is essentially neutral or alkaline. In a preferred embodiment, the process of producing the leaching solution is biocatalyzed.

In a preferred embodiment, a first process step of bio-oxidation of ore particles is accomplished to free (liberate) precious metals dispersed or occluded within the ore. A portion of the acidic, base-metal sulfate leach solution produced by the bio-oxidation step is introduced to an anaerobic reactor. In a heap leach embodiment of the process, the anaerobic reactor is a side- stream reactor or a series of such reactors in series. In one alternative slurry (e.g., vat or tank) leaching embodiment, the anaerobic process may occur on-line. One or more preferably non- toxic electron donors (such as hydrogen gas, formate, acetate and/or methanol—which does not bind effectively to activated carbon) and growth requirements (such as vitamins and/or salts), are added to the anaerobic reactor to enrich within it a culture of at least one oxyanion-reducing bacterium (e.g., a sulfate-reducing bacterium). In an alternative embodiment, the electron donors and/or growth requirements are derived from organic material deposited on the ore by sulfide- oxidizing bacteria during the bio-oxidation step. The hydrogen fugacity in theϊeactor, or at least in the last reactor in a series of such reactors, is maintained at a low level by at least one hydrogen-consuming bacterium. The anaerobic reactor may be operated in a pH-stat mode by adding sufficient acidic sulfate solution to maintain a neutral pH in the reactor (see Hunter, R.M., Biocatalyzed Partial Demineralization of Acidic Metal-Sulfate Solutions, Ph.D. Thesis, Montana State University, 1989). In an alternative embodiment, the anaerobic reactor may be operated in a sulfide-stat mode by adding sufficient sulfate solution to maintain a constant dissolved sulfide concentration in the reactor in response to signals from a sulfide sensor (e.g., sulfide ion selective electrode). Base metals are preferably precipitated and removed and a portion of the hydrogen sulfide gas (H₂S) produced in the anaerobic reactor is preferably removed. In this way, oxyanion- reducing bacteria are used to create an essentially neutral leaching solution comprising a relatively high concentration bisulfide ions, a high fugacity of hydrogen sulfide gas, a low concentration of dissolved base metals and a low fugacity of hydrogen gas.

In an alternative embodiment, the precious metal leaching solution is produced in an anaerobic environment by contacting a stream of gas comprising hydrogen sulfide gas and essentially no hydrogen gas with the solution until the environment has an appropriately high concentration of hydrogen sulfide gas and an appropriately low fugacity of hydrogen gas. The gas may be produced biotically by a culture of sulfate-reducing bacteria, or it may be produced abiotically by purifying H₂S gas to remove H₂ gas.

In a second process step, the oxidized ore (possibly in a heap that is covered and submerged to exclude oxygen) is leached (by recirculating the neutral or alkaline bisulfide lixiviant comprising, or saturated with, H₂S) in a leaching reactor. In one embodiment, the H₂S partial pressure is increased by introducing the lixiviant under pressure at the bottom of a heap submerged in water, causing ion concentrations to increase in direct proportion to the increase in H₂S partial pressure.

In a preferred embodiment, the anaerobic reactor and the leaching reactor are operated together as a single, essentially completely-mixed reactor. A completely mixed reactor is one that produces an effluent concentration of a conservative tracer (e.g., a non-reactive dye) equal to 37 ±3 percent of the initial tracer concentration (i.e., tracer mass divided by liquid volume) one detention time (i.e., liquid volume divided by liquid volumetric flow rate) after an impulse input (i.e., slug addition) of the tracer.

The complexed precious metal (e.g., gold and silver) is recovered (preferably

continuously) from the lixiviant solution. Recovery may be accomplished in a conventional manner by adsorption on activated carbon or by modifying either the solution pH, hydrogen fugacity, or oxidation-reduction potential (ORP).

Recovered precious metals are converted into products. This may include the operations of separating, smelting and casting of each precious metal into bars, bullion or other forms.

The present invention offers a variety of advantages not provided by the prior art. One object of the invention is to lower the monetary cost of gold, silver and platinum-group element production. By utilizing a waste product (excess sulfuric acid from a roasting or bio-oxidation pretreatment step) as the starting material for preparation of a bisulfide lixiviant, the lixiviant (a neutral bisulfide solution) would be produced biologically instead of being purchased. Another object of the invention is to use both inorganic (salts) and organic (biofilm carbonaceous compounds) byproducts of bioxidization as inputs to a precious-metal solubilization process.

Another object of the invention is to lower the environmental risk of precious metal mining. This is the case because the actual and perceived environmental risk of maintaining a large inventory of a neutral bisulfide solution is much lower than that associated with maintaining an equivalent volume of caustic cyanide solution. Another object of the invention is to provide a method and apparatus for both in-situ or ex-situ (conventional) mining. Further objects and advantages of the invention will become apparent from consideration of the drawings and the ensuing description.

BRIEF DESCRIPTION OF DRAWINGS

The features of the invention will be better understood by referring to the accompanying drawings which illustrate presently preferred embodiments of the invention.

In the drawings:

Fig. 1 is a highly schematic block diagram illustrating a first representative embodiment of the present invention.

Fig. 2 is a highly schematic block diagram illustrating a second representative embodiment of the present invention.

Fig. 3 is a highly schematic block diagram illustrating a third representative embodiment of the present invention. Fig. 4 is a highly schematic block diagram illustrating a fourth representative embodiment of the present invention.

The following reference numerals are used to indicate the parts of the invention on the drawings:

2 ore

4 bio-oxidation reactor

6 sulfate ions

7 electron donor

8 sulfate reduction reactor

10 bisulfide lixiviant

20 oxidized ore

22 bisulfide leaching reactor

24 pregnant solution

26 precious metals recovery reactor

28 leached ore

30 ore

32 crushing

34 crushed ore

36 acid leaching

37 aerobic reactor

38 air

40 acid-leach solution

42 pump

44 acid-leached ore

46 bisulfide leaching

47 essentially completely-mixed, anaerobic reactor

48 bisulfide lixiviant

50 pump

60 pH controller

62 valve

64 valve

66 pregnant bisulfide lixiviant

68 gold and silver recovery 70 spent lixiviant

76 bisulfide lixiviant recirculation loop

78 activated carbon column

80 leached ore

82 sensor/controller

84 electron donor

90 dewatering

92 contained bisulfide lixiviant

94 waste ore

96 acid-leach solution portion

98 base metal removal

100 base metal removal reactor

102 acid leach solution portion

104 iron and other base metals

110 excess hydrogen sulfide gas

112 excess hydrogen sulfide gas portion

114 sulfur recovery

116 sulfur recovery reactor

120 elemental sulfur

200 heap

202 heap, second heap

204 crushed ore

205 crushed ore, oxidized ore

206 air

208 plenum

210 acidic, base-metal sulfate leach solution

212 pump

214 portion

216 distributor

220 portion

230 anaerobic, sulfate-reduction reactor

232 pH controller

234 valve 238 bisulfide leach solution

240 non-toxic electron donor

244 base metals

250 settling tank

252 portion

254 bisulfide lixiviant

260 headspace

262 headspace

264 conduit

266 excess hydrogen sulfide gas

270 sulfur recovery

272 elemental sulfur

282 bisulfide lixiviant

284 plenum

286 pump

290 distributor

292 portion

294 portion

300 pregnant portion

302 reactor

306 barren lixiviant solution

312 aerobic, stirred batch reactor

314 air pump

318 pH monitor

320 liquid supernatant

322 ore

324 portion of dried, bio-oxidized ore

326 continuously stirred tank reactor

328 media reservoir

330 pump

332 pump

334 H₂S canister

336 pH monitor/controller 338 acidic supernatant

340 pump

342 effluent storage contai

344 pump

346 pump

348 160-ml serum bottles

MODE(S) FOR CARRYING OUT THE INVENTION

Reference is now made to Fig. 1 which is a schematic block diagram illustrating a preferred embodiment of the invention, with the dashed lines representing possible variations in the process and apparatus. Ore 2 is the input to the process and, under certain conditions, may be the only input to the process. In a preferred embodiment, ore 2 is crushed and may be otherwise treated to optimize bio-oxidation. In bio-oxidation reactor 4, oxidation of metal sulfides is accomplished to free or mobilized precious metals dispersed or occluded within metallic sulfides in ore 2.

Bio-oxidation reactor 4 produces a sidestream comprising sulfate ions 6 and acidity. In some instances, the sidestream also comprises biofilm carbonaceous compounds. In an alternative embodiment, bio-oxidation does not occur and sulfate ions 6 are an input to the process. Sulfate ions 6 may be a component of a waste stream, such as acid mine drainage, or by-product of ore roasting.

In a preferred embodiment, electron donor 7 is added to sulfate reduction reactor 8 so that sulfate ions 6 are biologically reduced therein. In a preferred embodiment, sulfate reduction reactor 8 is operated at a mean cell residence time low enough to cause essentially-complete (99+ percent) utilization of electron donor 7. In a preferred embodiment, sulfate reduction reactor 8 is operated in a pH-stat mode so as to maintain an essentially constant pH (± 0.1 pH unit) in reactor

8 and in bisulfide lixiviant 10 that it produces.

Oxidized ore 20 is introduced to bisulfide leaching reactor 22. In reactor 22, precious metal values in oxidized ore 20 are dissolved and complexed by means of bisulfide lixiviant 10. Pregnant solution 24 comprising precious metal values is introduced to precious metals recovery reactor 26 for precious metals recovery in a conventional manner by adsorption on activated carbon; or by modifying either the solution pH, hydrogen fugacity, or oxidation-reduction potential (ORP). A product (e.g., gold bullion) is formed from said precious metal by smelting and casting. In one embodiment, leached ore 28 is disposed of in a conventional manner (e.g., permanent storage) and need not be treated for removal of lixiviant. In a preferred embodiment, leached ore 28 is washed and/or dewatered to remove residual lixiviant 10 prior to disposal. Lixiviant 10 removed from leached ore 28 is used to wet and/or neutralize the acidic pH of incoming oxidized ore 20 and/or it is returned to leaching reactor 22.

It is well known in the art that the composition of ores varies widely, requiring optimization of the leaching step based on ore compositions and other local conditions. For this reason, design and operation of reactors 8 and 26 are preferably optimized for precious metal dissolution and complex formation. Depending on ore composition, design and/or operation are varied to achieve the following conditions in the reactor environment:

1. Maximize dissolved bisulfide concentration and dissolved H₂S_(aq) concentration at the precious metal surface;

2. Optimize pH;

3. Minimize hydrogen fugacity at the precious metal surface;

4. Maximize pressure;

5. Maximize temperature.

For gold and/or silver leaching, information on the aqueous chemistry of gold and silver bisulfide complexes and other chemical species likely to be present in a bisulfide lixiviant is used (see Barnes, H.L. (ed.), Geochemistry of Hydrothermal Ore Deposits, 2nd ed., New York: John Wiley & Sons, 1979). Published information on the aqueous chemistry of gold and silver bisulfide complexes and other chemical species likely to be present in a bisulfide lixiviant are used to produce a mathematical model of the solubilization step. The model incorporates the data presented in Tables 1 and 2. In the model, stability and equilibrium constants are used to predict the direction of a reversible chemical reaction under certain standard conditions and under other conditions. The standard conditions are 1.0 molar (M) concentrations of dissolved reactants and products and 1.0 atmosphere (Atm) pressure of gaseous reactants and products. The temperature is usually taken as 25°C (298°K), but stability and equilibrium constants are reported at other temperatures as well. →

→

-

→

- ^e →

→

→

-

Equilibrium constants can be derived in a number of ways. The stability constant for a reaction is related to the standard free energy change the reaction as follows:

ΔG° = -RT In K

Ln K = -ΔG7RT

log K = -ΔG72.30*RT

where ΔG° = free energy change under standard conditions, kJ

R = gas law constant = 0.008314 kJ/(mol*°K)

T = absolute temperature in degrees Kelvin (°K)

2.30 = ln 10

The value of the product (2.3*RT) is 4.1840*1.3636 K cal/reaction = 5.705 kJ/reaction at 25° C and 5.935 kJ/reaction at 37°C.

An opportunity for the solubilization of precious metal (e.g., gold, silver, platinum and palladium) is created by the complex-forming reactions:

Au_(s)+ H₂S_(aq) + HS- --> Au(HS)₂- + 0.5 H_2(g)

Ag_(s) + H₂S_(aq) + HS- --> Ag(HS)₂- + 0.5H_2(g)

Pt_(s) + 2H₂S_(aq) + 2HS- --> Pt(HS)₄ ^-2 + H_2(g)

Pd_(s) + 2H₂S_(aq) + 2HS- --> Pd(HS)₄ ² + H_2(g)

The equilibrium and stability constants for the gold and silver reactions have been determined experimentally and are dependent on the temperature at which they occur.

Shenberger, D.M. & Barnes, H.L., in "Solubility of gold in aqueous sulfide solutions from 150 to

350°C," (Geochimica et Cosmochimica Acta, 53, 269-278, 1989), have derived the following equation for the temperature dependence of the first reaction (between 150° to 350°C):

log K_Au(HS)2 = -9.383 *10⁵/T² + 2170.4/T - 2.2816

where T = temperature in degrees Kelvin

The following equation applies between 25°C and 150°C:

log K_AU(HS)2 = 3.32 + (-2,420/T)

Similarly, Gammons, CH. & Barnes, H.L., in "The solubility of Ag₂S in near-neutral aqueous sulfide solutions at 25 to 300°C," (Geochimica et Cosmochimica Acta, 53, 279-290, 1989), have derived the following equation for the temperature dependence of the second reaction:

log K_Ag(HS)2 = 0.439 - 943/T

The equilibrium and stability constants for the platinum group element reactions can be estimated using the methods disclosed by Hancock, R.D., Finkelstein, N.P., & Evers, A., in "A linear free-energy relation involving the formation constants of palladium (II) and platinum (II)," (Journal of Inorganic and Nuclear Chemistry, 39, 1031-1034, 1977) and Mountain, B.W. & Wood, S.A., in "Chemical controls on the solubility, transport, and deposition of platinum and palladium in hydrothermal solutions: A thermodynamic approach," (Economic Geology, 83, 492- 510, 1988). They have demonstrated that, for metals in the group Au, Ag, Pt and Pd, plots of the logarithms of the stability constants of one metal versus another are linear for a variety of ligands.

These equilibrium and stability constants may be used to determine the equilibrium molar concentration of each complex using the following equations:

[Au(HS)₂-] = K_Au(HS)2 * [H₂S_(aq)] * [HS-]/[H_2(g)] 0° 5

[Ag(HS)₂-] - K_Au(HS)2 * [H₂S_(ag)] * [HS- ]/[H_2(g)]0.5

[Pt(HS)₄ ^-2] = K_Au(HS)4 * [H₂S_(aq)]² * [HS- ]²/[H_2(g)]

[Pd(HS)₄ ^{- 2}] = K_Pd(HS)4 * [H₂S_(ag)]² * [HS-] ²/[H_2(g)]

The above equations are used to estimate the hydrogen fugacity required to solubilize gold and silver at various temperatures and at two reactant (H₂S + HS-) concentrations. Those estimates for gold and silver are presented in Table 3. In preparing the estimates presented in Table 3, the following activity and fugacity coefficients are used: Au (HS),- = 0.7,

Ag (HS)₂- = 0.7, H,S = 0 97, HS- = 0.7, and H₂ = 1.0.

Since anaerobic digesters must be operated at H_2(g) fugacities below about 10^-4 atmosphere (atm), achievement of the fugacities indicated in Table 3 are feasible. This would produce gold and silver concentrations in a pregnant bisulfide lixiviant of 0.1 to 1.0 mg/l and 1.0 to 10 mg/l, respectively, which are the same concentrations that occur in pregnant cyanide heap leach solutions during conventional cyanidation.

1

In a preferred embodiment, bisulfide ions are generated biologically (by naturally- occurring sulfate-reducing bacteria) at very low cost using an acidic waste product (bio- oxidation heap leach effluent) as the sulfate source. For example, with formate ion as the electron donor, the following reaction occurs:

8HCOO- + 2SO₄ ^-2 + 3H⁺→ 8HCO₃ + H₂S + HS-

With acetate ion as the electron donor, the following reaction occurs:

CH₃COO- + SO₄ ^-2→ 2HCO₃- + HS

A mass balance on typical heap leach effluents has shown that biological reduction of contained sulfate ions produces bisulfide and sulfide ions in excess of the concentration required for essentially complete base metals precipitation (see Hunter thesis, 1989). Thus, production of a high-concentration bisulfide lixiviant is possible. Moreover, H₂S gas can be recovered from spent lixiviant and/or leached ore by reducing the H₂S gas partial pressure in the gas mixture in contact with said spent lixiviant and/or leached ore using a vacuum pump. More complete H₂S gas recovery can be achieved by acidifying the spent lixiviant and/or leached ore to a pH below 7.0 and/or by increasing gas/liquid interfacial area, (e.g., by forming the liquid into droplets).

A variety of techniques may be used to maximize bisulfide concentrations in the lixiviant. The solubility of hydrogen sulfides in water decreases with increasing temperature (from about 7,100 mg/1 at 0°C to 3,000 mg/l at 30°C under a partial pressure of one atmosphere),

(Environmental Protection Agency, Process Design Manual for Sulfide Control in Sanitary Sewerage Systems. EPA 625/1 -74-005,October 1974). In accordance with Henry's law, the saturation concentration of H₂S in water is directly proportional to the partial pressure of the gas in the atmosphere in contact with the liquid. Removal of H₂S increases the pH of the solution (by removing protons). Moreover, the proportion of bisulfide ion (relative to H₂S) increases with pH over at least the range pH 5-9.

The optimal pH for the bisulfide lixiviant solution for precious metal recovery is the pH that maximizes the solubility of target precious metal compounds and the stability of their complexes. For example, Krauskopf, K.B., in "The solubility of gold" Economic Geology, (46, 858-870, 1951), noted that "one of the most perplexing facts about the chemistry of gold is its ability to dissolve in solutions of HS- of moderate concentration even at room temperature, whereas it dissolves in S^-2 (i.e., more alkaline solutions) only in concentrated solutions at high temperature." Schwarzenbach, von G. & Widmer, M., in "Die loslichkeit von Metallsulfiden," (Helvetica Chimica Acta, 49, 111-123, 1966), found that the solubility of silver was greatest at pH 7 at a temperature of 20°C in the presence of excess sulfide in the form of H₂S, HS-, and S^-2. Melent'yev, B.N., Ivanenko, V.V., and Pamfilova, L.A., in "Solubility of some ore-forming sulfides under hydrothermal conditions," (Rastvorimost' nekotorykh rudoobrazuyushkikh sul'fidov v gidrotermal' nykh usloviyakh' Moskva, 27-102, 1968), found that the solubility of Ag₂S increases with pH in the range pH 4-8 in the temperature range 100-300°C. Seward reported that for gold in solutions of reduced sulfur "a pronounced solubility maximum occurs in the region of pH about 7." (Seward, T.M., "Thio complexes of gold and the transport of gold in hydrothermal ore solutions," Geochimica et cosmochimica Acta, 37, 379-399, 1973).

Options for reducing hydrogen fugacity include bioprocessing hydrogen-management techniques. Hydrogen-consuming bacteria (hydrogenotrophs) include such anion-reducing bacteria as acetogens, methanogens, sulfate-reducing bacteria and denitrifying (nitrate-reducing) bacteria. In natural ecosystems, these bacteria participate in "interspecies hydrogen transfer." Examples of acetogens include Acetobacterium woodi (ATCC 29683, DSM 1030, or DSM 2396) and Clostridium aceticum (ATCC 35044 or DSM 1496). Examples of hydrogen- consuming methanogens are numerous and include the mesophiles Methanobrevibacter ruminantmm (ATCC 35063 or DSM 1093) and Methanosarcina barken (ATCC 29786 or DSM 805), and the thermophile Methanobacterium thermoautotrophicum (ATCC 29096 or DSM 1053). Examples of hydrogen-consuming sulfate-reducing bacteria are shown in Table 4.

.

As noted, some of the sulfate-reducing bacteria listed in Table 4 are also nitrate-reducing bacteria because they can also reduce nitrate to produce ammonium. In a preferred embodiment, leaching solution is produced in an anaerobic reactor by culturing in the reactor sulfate-reducing bacteria capable of using formate or acetate, as well as hydrogen as electron donors, and both sulfate and nitrate as electron acceptors. Since anions, such as sulfate and nitrate, are reduced, such bacteria are oxyanion-reducing bacteria. Examples of such bacteria include mesophilic, fresh- water species such as Desulfobacterium catecholicum DSM 3882 (acetate and formate) and Desulfovibrio simplex DSM 4141 (formate); mesophilic, salt-water species, such as Desulfovibrio salexigens DSM 2638 (formate); and thermophilic, fresh- water species such as Desidfomacidum kuznetsovii DSM 6115 or VKM B-1805 (acetate and formate). Microorganisms with ATCC accession numbers can be obtained from the American Type Culture Collection, 12301 Parklawn Drive, Rockville, Maryland 20852-1776, tel 1-800- 638-6597, fax 1-301-231-5826. Microorganisms with DSM accession numbers can be obtained from the Deutsche Sammlung von Mikroorganismen und Zellkulturen GmbH, Mascheroder Weg lb, D-38124 Braunschweig, Germany, tel 011-49 (0)531-2616-336, fax 011-49 (0)531-

2616-418. Microorganisms with VKM accession numbers can be obtained from the Institute of Biochemistry and Physiology of Microorganisms of the Russian Academy of Science,

Pushchino-na-Oke, 142292 Moscow Region, Russian Federation.

In an alternative embodiment, preferred especially for laboratory (process optimization) studies, additional hydrogen consumption is accomplished by purging the headspace of bisulfide leaching reactor 22 through aH₂S-scrubbing means (e.g., a zinc acetate solution "bubbler") into a nitrate-fed reactor containing a culture of sulfate-reducing bacteria that are also capable of nitrate reduction, operated in parallel (side-stream) or in series with reactor 22. In this way, "hydrogen-scrubbed" headspace gas is recycled back to reactor 22. Alternatively, a zinc acetate bubbler would not be required if the H,S concentration in reactor 22 were controlled independently by a side-stream bubbler controlled by a sulfide ion-selective electrode that would turn on a H₂S-scrubbing bubbler loop when a high H₂S setpoint was reached.

In an alternative embodiment, the H₂S_(aq) concentration and the HS- concentration may be increased to an appropriate level, and the H₂ fugacity may be reduced to an appropriate level in the environment provided by reactor 22 by contacting the contents of reactor 22 with a stream of gas having an appropriate H₂S fugacity and effectively no H₂. This stream of gas may be produced biologically by a culture comprising sulfate-reducing bacteria or it may be produced abiotically using conventional means. An equilibrium will be reached that partitions the constituents of reactor 22 of limited solubility between the gas and liquid phases in reactor 22. Henry's Law can be used to predict equilibrium and steady state constituent levels.

While gold, silver and platinum-group elements are soluble in bisulfide solutions at ambient (atmospheric) pressures and at room temperature, their solubilities generally increase with pressure and temperature (Krauskopf, K. B., Economic Geology, 46, 858-870, 1951; Weissberg, B.C., Economic Geology, 65, 551-556, 1970). For this reason, in an alternative embodiment, sulfate-reduction reactor 8 is operated in the thermophilic (50-100°C) and barophilic (over one atmosphere) ranges (e.g., in a submerged, covered heap).

If sulfate-reduction reactor 8 is operated at steady-state at relatively high total dissolved sulfide (H₂S _(aq) + HS- + S^-2) concentrations (say over about 1,000 mg/l in the liquid), then sulfate-reducing bacteria will be enriched in reactor 8 that are relatively resistant to growth rate inhibition by such total sulfide concentrations. Many investigators have reported that common sulfate-reducing bacteria can grow in media containing over 2,700 mg/1 of total sulfides (See Miller, L.P. (1950). Formation of metal sulfides through the activities of sulfate-reducing bacteria. Contributions from Boyce Thompson Institute, 16, 85-89; Saleh, A.M., MacPherson, R., & Miller, J.D.A. (1964). The effect of inhibitors on sulphate reducing bacteria: a compilation. Journal of Applied Bacteriology, 27, 281-293).

In a preferred embodiment, sulfate-reduction reactor 8 is operated at a relatively low total sulfide concentration (say less than about 1,000 mg/1 in the liquid) in order to minimize inhibition of the sulfate-reducing bacteria growing in it. This may be achieved by using a vacuum pump or purging gas stream to transfer H₂S gas from the headspace of reactor 8 to the liquid in leaching reactor 22.

In one embodiment, a H₂S gas pump is used to increase the H₂S partial pressure in the transferred gas stream. With this embodiment, the H₂S gas removed from reactor 8 is absorbed in a basic (pH >7) solution as dissolved HS ions during the intake portion of the pumping cycle.

During a subsequent discharge portion of the pumping cycle, the solution containing dissolved HS^' ions is acidified to convert the dissolved HS- to H₂S gas and the H₂S is pumped into leaching reactor 22. In one embodiment, waste sulfuric acid produced by oxidation of metal sulfides is used to acidify the solution containing the dissolved HS- ions.

In an alternative embodiment, H₂S gas pumping is accomplished by dissolving it in a liquid solution at a relatively low temperature (e.g., 10°C). The H₂S is then driven out of the solution by heating the liquid to a relatively higher temperature (e.g., 60°C). This form of H₂S pumping is made possible by the significant change in Henry's law coefficient for H₂S gas with temperature.

Precious metals recovery options include adsorption on activated carbon; adsorption on ion-exchange resin; and modification of the solution pH, hydrogen fugacity, or oxidation- reduction potential (ORP). In an alternative embodiment, precious metals are adsorbed on the cell walls of bacteria and the bacteria are separated from the liquid in which they are suspended by settling and/or filtration of the liquid after settling of the ore particles. Options that do not otherwise modify lixiviant solution chemistry are preferable. For this reason, in preferred embodiments, at least reactors 8 and 22, and preferably also reactor 26, are operated together as a single, essentially completely-mixed reactor.

In an alternative embodiment, pregnant solution 24 is degassed to reduce its total dissolved sulfide concentration before and/or concurrent with contacting it with granular activated carbon in precious metals recovery reactor 26. Degassing may be accomplished by pumping gas from the headspace of reactor 26 into the liquid in leaching reactor 22. Precious metals that have absorbed to the activated carbon are eluted into a concentrated solution that is a solvent for the precious metals. Precious metals are recovered from the concentrated solution by conventional means.

Recovered precious metals are converted into products. This may include the operations of separating, smelting and casting of each precious metal into bars or bullion.

Reference is now made to Fig. 2 which is a schematic diagram illustrating a second alternative representative embodiment of the invention, with dashed lines representing possible variations in the process and apparatus. In this embodiment, ore 30 preferably undergoes crushing 32 to facilitate exposure of precious metal values in the ore to processing solutions. Crushed ore 34 then undergoes acid leaching 36 in aerobic reactor 37. If necessary, air 38 containing oxygen and carbon dioxide is added in the acid leaching step. Acid-leach solution 40 is recirculated through the ore undergoing acid leaching by means of pump 42.

Acid-leached ore 44 then undergoes bisulfide leaching 46 in essentially completely- mixed, anaerobic reactor 47. Bisulfide lixiviant 48 is recirculated through the ore undergoing bisulfide leaching by means of pump 50. The pH of bisulfide lixiviant 48 is established at an optimum pH by pH controller 60 which controls the rate of addition of acid-leached ore 44 and acid-leach solution 40 to reactor 47 by means of valves 62 and 64. The sulfate and/or the sulfide concentration in bisulfide lixiviant recirculation loop 76 is monitored by sensor/controller 82, which may comprise an ion-specific electrode. Sensor/controller 82 is programmed to add up to a stoichiometric amount of electron donor 84, which is a sulfate-reducing bacteria growth substrate such as formate, acetate or methanol, to bisulfide lixiviant recirculation loop 76.

Pregnant bisulfide lixiviant 66, which contains precious metal values is subjected to gold and silver recovery 68. Recovered gold and silver is converted into products (e.g., bars of essentially pure metal). Spent lixiviant 70 is returned to bisulfide lixiviant recirculation loop 76. In a preferred embodiment, gold and silver recovery 68 is accomplished by passing pregnant bisulfide lixiviant 66 through activated carbon column 78.

Leached ore 80 undergoes dewatering 90 by conventional means, such as settling and/or vacuum filtration. Contained bisulfide lixiviant 92 is returned to bisulfide lixiviant recirculation loop 76. Waste ore 94 is disposed of by using conventional means.

In an alternative embodiment, acid-leach solution portion 96 undergoes base metal removal 98 in base metal removal reactor 100. Excess hydrogen sulfide gas 110 removed from anaerobic reactor 47 is introduced to base metal removal reactor 100 to precipitate iron and other base metals 104. Acid-leach solution portion 102 having a reduced base metal content may be returned to reactor 37, or optionally, to reactor 47.

In an alternative embodiment, excess hydrogen sulfide gas portion 112 undergoes sulfur recovery 114 in sulfur recovery reactor 116. Recovery of element sulfur 120 may be accomplished by the conventional Claus process or by means of the process disclosed in U.S. Patent No. 4,666,852, which disclosure is incorporated herein as if fully set forth.

Reference is now made to Fig. 3 which is a schematic diagram illustrating a third alternative representative embodiment of the invention, with dashed lines representing possible variations in the process and apparatus. In this embodiment, sequential processing of heaps 200 and 202 of crushed ore 204 and 205 is accomplished. In heap 200, conventional bio-oxidation of crushed ore particles 200 is accomplished to free precious metals dispersed or occluded within the ore. Air 206 may be introduced to heap 200 via plenum 208. Acidic, base-metal sulfate leach solution 210 is collected from the bottom of heap 200 through plenum 208 by means of pump 212. Portion 214 of leach solution is recirculated by means of pump 212 and distributor 216 to the top of heap 200.

As was noted above, bio-oxidation of heap 200 may include ore crushing, acid pretreatment, inoculation with appropriate sulfide-oxidizing bacteria, addition of nutrients, recirculating the biolixiviant and cooling the heap (for 3 to 8 days), and allowing the heap to "rest" (for 3 to 8 days). Additional process steps may include washing heap 200 for an extended period (e.g., 14 days) to remove residual acidity or iron content, and breaking heap 200 apart in order to agglomerate ore 202 with cement and/or lime to make a new heap, such as heap 202.

Portion 220 of acidic, base-metal sulfate leach solution 210 produced by the bio- oxidation step is introduced to anaerobic, sulfate-reduction reactor 230. In this embodiment of the process, reactor 230 is a side-stream reactor. The rate of addition of portion 220 to reactor 230 may be controlled by pH controller 232 which operates valve 234 to create an optimum pH for precious metals leaching in bisulfide leach solution 238 produced by reactor 230. Preferably, non-toxic electron donor 240 (such as formate, acetic acid (e.g., vinegar), acetate, or methanol- which does not bind effectively to activated carbon), is added to anaerobic reactor 230 to enrich within reactor 230 a microbial culture comprising sulfate-reducing bacteria. Anaerobic reactor 230 is preferably operated in a pH-stat mode by adding a sufficient portion 220 of acidic sulfate solution to maintain a neutral pH in reactor 230. In some embodiments, the concentration of dissolved sulfide (H,S, HS-, and S^-2) in the anaerobic reactor is maintained below about 2,500 mg/1 to prevent inhibition of the microbial culture comprising sulfate-reducing bacteria.

In a preferred embodiment, base metals 244 (such as iron) are precipitated in

downstream settling tank 250, and portion 252 of clarified bisulfide lixiviant 254 is recirculated to reactor 230. The rate of recirculation of portion 252 is preferably chosen so that reactor 230 and settling tank 250 are operated together as a single, essentially completely-mixed reactor. Headspace 260 of reactor 230 and headspace 262 of settling tank 250 are preferably connected by conduit 264. Excess hydrogen sulfide gas (H₂S) 266 produced in anaerobic reactor 230 (e.g., that amount over about 2,700 mg/l) and tank 250 is preferably removed. In some embodiments, excess hydrogen sulfide gas undergoes sulfur recovery 270 to produce elemental sulfur 272. By means of reactor 230 and tank 250, sulfate-reducing bacteria are used to create clarified, approximately neutral (pH 6 to 9) leaching solution 254 comprising bisulfide ions and a low concentration of dissolved and suspended base metals. Bisulfide lixiviant 254 and headspace 260 comprise the reactor environment of reactor 230.

In another preferred embodiment, excess H₂S gas produced in reactor 230 is removed from headspace 260 and/or headspace 262 by means of a H₂S gas pump (not shown) and transferred into clarified bisulfide lixiviant 254 downstream from settling tank 250. In this was the concentrations of H₂S_(aq) and HS- in the lixiviant are increased after most of base metals 244 are removed from it.

In a preferred embodiment, heap 200 is undergoing bio-oxidation while a second heap 202, which has previously undergone bio-oxidation, undergoes leaching with bisulfide lixiviant. In a second process step, oxidized ore 205 is preferably covered with cover 208 and submerged in bisulfide lixiviant 282 to exclude oxygen. Heap 202 is leached by recirculating portion 292 of neutral bisulfide lixiviant 282 saturated with H₂S through it by means of plenum 284, pump 286, and distributor 290. In an alternative embodiment, the H₂S partial pressure is increased by introducing the lixiviant [and/or H₂S gas having a low concentration (less than 1,000 parts per million by volume) of H₂ gas] under pressure at the bottom of a heap via plenum 284 which is submerged in lixiviant 282, causing HS- ion concentrations to increase in direct proportion to the increase in H₂S partial pressure. This may increase the concentration of dissolved sulfide (H₂S, HS-, and S^-2) in heap 202 above 2,500 mg/l. In a preferred embodiment, anaerobic reactor 230, settling tank 250, and heap 202 are operated together as a single, essentially completely- mixed reactor by recirculating portion 294, from heap 202, to reactor 230.

In an alternative embodiment, pressure sensors are placed at multiple points throughout the system for safety reasons. This provides a warning system for users of the system, since releases of H₂S_(g) can be toxic. Low pressures sound an alarm, indicating a leak somewhere, while high pressures indicate unsafe operation. The use of multiple gauges pinpoints the source of the problem quickly. The pressure gauges are also used to monitor and regulate the H₂S_(g) pressures to optimize the solubility of the gold and silver.

Additionally, conductivity and total dissolved solids meters are placed in the effluent streams of the sulfate-reducing reactor in order to measure the ionic strength of the solvent. The meters are used to monitor the ionic strength of the solvent, which controls the activity coefficients of the gold and silver complexes, H₂S_(aq), HS-, and H_2(g). Control of the activities of these compounds increases the efficiency of solubilizing the gold and silver.

Complexed gold and silver in pregnant portion 300 of lixiviant 282 is recovered continuously from the lixiviant solution in reactor 302. Recovery may be accomplished in a conventional manner by adsorption on activated carbon or by precipitation on zinc dust or by modifying either the solution pH, hydrogen fugacity, or oxidation-reduction potential (ORP). Metal that has been recovered from activated carbon eluent by electrowinning or zinc dust may be smelted to recover precious metal values as products such as jewelry or electronic system components. Barren lixiviant solution 306 is recycled to heap 202. Working Example No. 1

A chemostat having a working (liquid) volume of 5 liters and a headspace volume of 2.5 liters was operated at a dilution rte of 0.006 per hour for over 6 hydraulic detention times so that steady state conditions were achieved. A sulfate-reducing bacteria growth medium comprising formate ions was pumped into the chemostat at a constant rate. The pH of the liquid in the chemostat was maintained at 7.0 by means of a pH controller that added bio- oxidation process effluent (acidic metal sulfate solution) to the reactor as required.

Sufficient formate ions and sulfate ions were introduced to the chemostat to produce a headspace H₂S partial pressure of about 1 atmosphere. Achievement of this partial pressure of H₂S was assured by purging the chemostat with a gas containing 99.5+ percent H₂S at the beginning of the experiment.

The concentration of H₂ gas in the chemostat headspace before it was purged and in the gas used to purge the chemostat was measured by means of a gas chromatograph with a thermal conductivity detector. The concentration of H₂ in the headspace was about 300 parts per million (ppm) by volume and the H₂ concentration in the purging gas was about 200 ppm.

The liquid level and the liquid volume in the chemostat was kept constant by

withdrawing liquid and headspace gas from the chemostat at a greater rate than liquid was added to the chemostat. The chemostat effluent contained about 200 mg/l of formate. The effluent was discharged to a reservoir, the headspace of which was connected to the headspace of the chemostat.

A square of gold foil about 0.1 inch on a side and 0.025 inch thick was placed in a 160- milliliter (ml) serum bottle and a Teflon ® septum stopper was crimped on the bottle mouth. The bottle was purged with oxygen-free nitrogen gas and 100 ml of chemostat effluent was transferred to the bottle without exposing it to air.

The contents of the bottle were then purged three times with the afore described H₂S gas mixture at about 3-day intervals. Within four hours of the initial purging, the liquid in the bottle took on a bright yellow color. Testing of a six-ml sample of the liquid plus two ml of aqua regia revealed that the liquid contained about 0.3 mg/l of gold.

A 10-ml sample of the liquid was withdrawn from the bottle and introduced

anaerobically to a similar serum bottle containing washed granular activated carbon and the H₂S gas mixture. The liquid in contact with the activated carbon immediately became colorless indicating adsorption of the gold on the activated carbon granules. Assays of the activated carbon revealed that gold was adsorbed on the carbon.

Working Example No. 2

Reference is now made to Fig. 4 which is a schematic diagram illustrating a fourth alternative representative embodiment of the invention, with dashed lines representing possible variations in the process and apparatus. In this embodiment, an experiment was conducted to illustrate the disclosed method and apparatus on low-grade samples of gold ore. Experimental procedures and results are presented below.

Each 1.5 kilogram ore sample was ground to a fine powder that could pass through a 150-mesh sieve (sieve opening = 0.004 inch). The entire sample was split into three representative samples of approximately 500 grams each. The first representative aliquot of the ore was assayed three times for gold and silver content, as well as for the presence of trace elements. The second and third representative aliquots were bio-oxidized to oxidize (and solubilize) metal sulfides and mobilize gold and silver values.

Bio-oxidation was accomplished in aerobic, stirred, batch reactor 312 having a working volume of 5 liters. Batch reactor 312 was placed in a water bath (not shown) having a temperature of 35°C About 1,000 grams of the ground ore was suspended in about 5 liters of an acidic Thiobacillus ferrooxidans medium in the reactor. The acidic medium as described in ASTM Standard E 1357 contained the constituents shown in Table 5 and its pH was adjusted to pH 2 with concentrated sodium hydroxide (NaOH). Air and carbon dioxide were introduced into the suspension by pumping air into it at a relatively high rate with air pump 314. The suspension was inoculated with an active culture of Thiobacillus ferrooxidans, ATCC 13661, obtained from the American Type Culture Collection at the address given above. The progress of bio-oxidation was monitored by measuring pH (with first pH monitor 318) and dissolved iron concentrations in the acidic medium.

After a period often to twelve days, when the rate of increase in iron concentrations in the acidic medium slowed and the pH stabilized, ore 322 was separated from liquid supernatant 320 by settling and drying in an oven at 140°C A first representative portion of dried, bio- oxidized ore 324 was subjected to conventional cyanide extraction, and then assayed for gold content to provide a basis of comparison with the bisulfide extraction. A leaching solution comprising dissolved hydrogen sulfide gas and bisulfide ions was produced in continuously stirred tank reactor (CSTR or chemostat) 326 having a working (liquid) volume of five liters and a headspace volume of 2.5 liters. Chemostat 326 was placed in a water bath (not shown) having a temperature of 35°C. Chemostat 326 was started in a batch mode by placing a sulfate-reducing bacteria medium in the chemostat, inoculating the chemostat with wild sulfate-reducing bacteria and allowing the culture to acclimate for 5-7 days.

After an acclimation period, sulfate-reduction medium 328 containing the constituents shown in Table 6 and formate as a carbon source was pumped into reactor 326 by pump 330 at a rate that produced a dilution rate of about 0.005 per hour. Liquid effluent was removed from the chemostat by pump 332 at the rate required to maintain the liquid level in the chemostat at a set level and discharged to effluent storage container 342. This dilution rate produced a mean cell residence time in the reactor that was much less than the maximum specific growth rate of the sulfate-reducing bacteria used to inoculate it. Chemostat 326 was operated in a pH-stat mode at pH 7.0 by continuously monitoring the pH of the liquid in chemostat 326 with pH monitor/controller 336, and by intermittently pumping acidic supernatant 338 produced by the bio-oxidation step into chemostat 326 with pump 340.. Addition of acidic supernatant 338 to chemostat 326 increased the dilution rate to about 0.006 per hour. The chemostat headspace was periodically purged with hydrogen sulfide from canister 334 to maintain positive pressure within the reactor. After chemostat 326 had been operating for about three hydraulic detention times and had reached steady state, the effluent from the reactor was used as solvent in leaching experiments.

Pyrex serum bottles 348 with a capacity of 160 ml were used as batch leaching reactors. The reactors had previously been washed with aqua regia because Pyrex is known to adsorb gold complexes under certain conditions. Representative four-gram portions of the bio-oxidized ore were added to the reactors. The reactors were augmented with 4 gram portions of prewashed 4-12 mesh activated carbon. Effluent collected from the chemostat was dispensed in 100 ml aliquots into the leaching reactors by pump 346. The reactors were immediately capped, sealed, purged and pressurized to 1 atmosphere absolute with 99.5 percent pure hydrogen sulfide gas. The reactors were then placed in both 35 and 65 °C incubators. The experiments were mixed by hand about two times daily and purged and pressurized with hydrogen sulfide gas at least every 48 hours.

Samples of the liquid phase from the reactors were taken several times throughout the experiment. The 6 ml liquid samples were taken from well settled reactors and filtered with a 0.2 μm millipore filter into 10 ml serum bottles. The samples were preserved in 2 ml aqua regia and analyzed for gold concentrations by inductively-coupled plasma atomic emission spectroscopy (ICP) analysis.

Upon completion of the experiments, all the components of the experiments were analyzed for gold and silver concentrations. Liquid samples were taken from the reactors, then the bottles were opened and the components were separated. The reactors were shaken and poured through a small mesh sieve (to collect the activated carbon) into a vacuum filter funnel containing a Whatman glass fiber filter. The liquid passed through and the ore portion was collected on the filter. The activated carbon was washed with distilled water, blotted dry with filter paper and dispensed into serum bottles. The ore and the filter were removed from the apparatus and inserted into serum bottles. Both the carbon and the ore samples, were analyzed by fire assay for gold and silver concentrations. The liquid samples were prepared as described above and analyzed for gold by ICP analysis.

Bisulfide leaching results are presented in Table 7. The relatively low recovery percentages were attributed to infrequent mixing of serum bottle contents and infrequent purging with H₂S gas. Because the rates of metal dissolution reactions are controlled by the rates at which reactants can reach and products can leave the metal surface as well as by reactant and product concentrations at the metal surface, subsequent experimental designs addressed these factors.

INDUSTRIAL APPLICABILITY

The invention has utility as a means of extracting precious metals from ore that is being mined in situ or ex situ. The invention can also be used to recover precious metals from scrap.

Many variations in configurations have been discussed and others will occur to those skilled in the art. Some variations within the scope of the claims include biotic and abiotic means for producing the bisulfide lixiviant and other H₂S gas pumping schemes. All such variations within the scope of the claims are intended to be within the scope and spirit of the present invention.

Claims (20)

We claim:

1. A process of making a product, said product comprising a precious metal that can form a complex comprising a sulfide, said complex having an aqueous solubility, said process comprising the steps of:

introducing an ore of said precious metal into an anaerobic reactor environment, said reactor environment having an activity of dissolved hydrogen sulfide gas, an activity of dissolved sulfide ions, and a fugacity of hydrogen gas;

increasing said fugacity of hydrogen sulfide gas and said activity of sulfide ions and reducing said fugacity of hydrogen gas, whereby the solubility of the precious metal-sulfide complex is increased;

removing the precious metal-sulfide complex from said reactor environment;

separating said precious metal from said metal-sulfide complex; and

forming said product from said precious metal;

wherein said increasing step and said reducing step are accomplished by at least one step selected from the group consisting of:

contacting said reactor environment with a culture of an oxyanion-reducing bacterium, and

contacting said reactor environment with a stream of gas comprising hydrogen sulfide gas and less than 1,000 parts per million by volume of hydrogen gas.

2. The process of claim 1 wherein said dissolved sulfide ions comprise dissolved bisulfide ions.

3. The process of claim 1 wherein said oxyanion-reducing bacterium is a sulfate- reducing bacterium.

4. A method for extracting a precious metal from an ore comprising the steps of producing a leaching solution, said solution comprising dissolved hydrogen sulfide gas and bisulfide ions, and having a fugacity of hydrogen gas;

reducing the fugacity of hydrogen gas in said leaching solution by means of an oxyanion-reducing bacterium;

exposing said ore to said leaching solution; and

recovering said precious metal from said leaching solution.

5. The method of claim 4 wherein the producing step is carried out by means of a sulfate-reducing bacterium.

6. The method of claim 4 wherein the oxyanion-reducing bacterium is selected from the group consisting of:

an acetogen,

a methanogen,

a sulfate-reducing bacterium, and

a nitrate-reducing bacterium.

7. The method of claim 4 wherein the oxyanion-reducing bacterium is selected from the group consisting of:

Desulfobacterium catecholicum (DSM 3882),

Desulfovibrio simplex (DSM 4141),

Desulfovibrio salexigens (DSM 2638), and

Desulfomacidum kuznetsovii (DSM 6115 or VKM B-1805).

8. A process for precious metal leaching comprising the steps of

biologically reducing a dissolved sulfate under anaerobic conditions to produce a bisulfide lixiviant having an approximately neutral pH and a hydrogen gas fugacity of less than

0.001 atmospheres,

leaching a first portion of an ore comprising at least one precious metal by exposing said ore to said bisulfide lixiviant, and

recovering said at least one precious metal from said bisulfide lixiviant.

9. The process of claim 8 further comprising

biologically oxidizing said first portion of an ore prior to said leaching step.

10. The process of claim 8 wherein said dissolved sulfate is produced by biologically oxidizing a second portion of an ore.

11. The process of claim 8 wherein said biologically reducing step and said leaching step are accomplished in an essentially completely mixed reactor and at a pH in the range of 6 to 8 units.

12. The process of claim 10 wherein said biologically oxidizing step occurs in a first heap of ore and said leaching step occurs in a second heap of ore.

13. The process of claim 12 wherein said recovering step is comprised of exposing pregnant bisulfide solution to activated carbon and eluting said precious metals from said activated carbon.

14. An apparatus for extracting a precious metal from an ore comprising:

means for producing a leaching solution, said solution comprising dissolved hydrogen sulfide gas and bisulfide ions, and having a fugacity of hydrogen gas;

means for reducing the fugacity of hydrogen gas in said leaching solution by means of an oxyanion-reducing bacterium;

means for exposing said ore to said leaching solution; and

means for recovering said precious metal from said leaching solution.

15. The apparatus of claim 14 wherein the means for producing comprises a sulfate- reducing bacterium.

16. The apparatus of claim 14 wherein the oxyanion-reducing bacterium is selected from the group consisting of:

an acetogen,

a methanogen,

a sulfate-reducing bacterium, and

a nitrate-reducing bacterium.

17. The apparatus of claim 14 wherein the oxyanion-reducing bacterium is selected from the group consisting of:

Desulfobacterium catecholicum (DSM 3882),

Desulfovibrio simplex (DSM 4141), Desulfovibrio salexigens (DSM 2638), and

Desulfomaculum kuznetsovii (DSM 6115 or VKM B-1805).

18. An apparatus for precious metal leaching comprising:

means for biologically reducing a dissolved sulfate under conditions that produce a bisulfide lixiviant having a hydrogen gas fugacity of less than 0.001 atmospheres, whereas the conditions that produce a bisulfide lixiviant comprise

absence of oxygen,

presence of an electron donor,

presence of a source of carbon, and pH in the range of 6 to 9 units; means for leaching a first portion of an ore comprising at least one precious metal by exposing said ore to said bisulfide lixiviant; and

means for recovering said at least one precious metal from said bisulfide lixiviant wherein the means for recovering said at least one precious metal from said bisulfide lixiviant is selected from the group consisting of

means for adsorbing said at least one precious metal on activated carbon, means for changing the pH of said bisulfide lixiviant,

means for increasing the hydrogen fugacity of said bisulfide lixiviant,

means for changing the oxidation-reduction potential of said bisulfide lixiviant, means for reducing the pressure of said bisulfide lixiviant, and

means for reducing the temperature of said bisulfide lixiviant.

19. The apparatus of claim 18 further comprising

means for biologically oxidizing said first portion of an ore prior to its introduction into said means for leaching.

20. The apparatus of claim 18 wherein said means for biologically reducing a dissolved sulfate comprises a completely mixed stirred tank reactor operated at a dilution rate of about 0.005 per hour and said apparatus further comprising means for monitoring and controlling the ionic strength of the lixiviant and thereby optimizing an activity coefficient of a reactant or a product.