Patent Application
20100012502
This invention relates to the recovery of metal-containing values from minerals and/or ores containing the same, and more particularly to the recovery of metal-containing values from minerals and/or ores by a process of reduction. (By “metal-containing values” is meant metals and/or metal compounds). In a preferred embodiment this invention relates to the recovery of values containing copper, i.e. metallic copper and/or copper compounds, from chalcopyrite and other ores containing copper and also preferably containing sulfur.
Approximately 70% of the world's copper reserves are contained in the mineral chalcopyrite (CuFeS2). The structure of this mineral renders it especially refractory to direct oxidative leaching. For that reason, to date the most economical method to extract its metallic values is through smelting, a high-temperature process that produces toxic sulfur dioxide, which must be stabilized, generally as sulfuric acid, in order for the process to meet environmental norms.
For the last forty years or so, scientists and. engineers have researched various moderate-temperature, aqueous phase alternatives to smelting. However, the only viable techniques have involved bacterial attack of the mineral, which oxidizes the sulfur to sulfate, liberating the metallic ions into solution. Unfortunately, the technique is extremely slow and is cost-effective only for low grade minerals. Attempts to oxidize the chalcopyrite in the aqueous phase with moderate to high pressures of oxygen at temperatures near or at the boiling point of the solution have encountered difficulties due to passivation by sulfur originated from the chalcopyrite itself. Other work on recovering copper values from chalcopyrite and metal values from other ores using hydrolytic techniques has been carried out, but again none of the results of such work have been commercialized.
For instance, Australian researcher Thomas Biegler and his colleagues reported electrolytic reduction of chalcopyrite in a slurry or particle bed through contacting with a copper, lead, mercury or graphite cathode with the electrolytic chamber, as typical in such processes, divided into anodic and cathode compartments. This work is described, for instance in Biegler et al., “Continuous electrolytic reduction of a chalcopyrite slurry”, J. Appl. Electrochemistry 7:175 (1975), Biegler et al., “Upgrading and activation of chalcopyrite concentrate by slurry electrolysis”, Transactions of the Institution of Mining and Metallurgy, Section C 23:23 (1976) and Australian patent 495,175 (published application 1975-80050). Some work on electrolytic reduction of zinc ferrite is reported in Bhat et al., “Electroleaching of zinc leach residues”, Hydrometallurgy 18:287 (1987). More recent work is discussed, and others were summarized, by Dreisinger et al., “A fundamental study of the reductive leaching of chalcopyrite using metallic iron, part 1: kinetic analysis”, Hydrometallurgy 66:37 (2002).
Prior efforts at hydrolytic reduction of such ores often called for somewhat elevated temperatures of about 60-70° C. for best results, with much poorer results at lower temperatures. In addition, these reductive leaching techniques require a large excess of a reducing agent, usually metallic iron, lead or copper and temperatures near the aqueous solution boiling point, using sulfuric acid. This renders the processes uneconomical and, in the case of sulfide-containing ores, produces toxic hydrogen sulfide gas.
The present invention proves a novel method to reduce chalcopyrite (CuFeS2) and other minerals that are susceptible to reduction, to a less refractory mineral phase (for example, CuxS) that can be conducted at near ambient temperatures and pressures, without the production of toxic gases.
In brief, the invention comprises a process for the recovery of metal-containing values from an ore, or from one or more minerals, that are not readily oxidizable and that are susceptible to reduction, comprising:
The present invention builds on the concept, as shown for instance by Dreisinger et al. supra, of first reducing chalcopyrite to resulting mineral phases (chalcocite (Cu2S) or djurleite (Cu1.96S) that are more easily oxidized to produce copper. The inefficiencies of the previous techniques, however, are overcome by the process of this invention, in which the reduction of chalcopyrite, as well as of other ores mentioned below, is carried out using a cathode comprising a material that promotes the generation of monatomic hydrogen (e.g., via the H.E.R., or Hydrogen Evolution Reaction) and in which the reduction is carried out in an electrolysis chamber having a single undivided acidic electrolytic medium.
In an electrolytic process there are always two reactions taking place simultaneously—reduction at the cathode and oxidation at the anode. The overall reactions occurring during the classical electrolysis of water are the production of hydrogen gas at the cathode and oxygen gas at the anode. In the present process, when chalcopyrite is used as the starting ore or mineral, the generation of hydrogen at the cathode results in reduction of chalcopyrite to form chalcocite, hydrogen sulfide and ferrous ion. The process is carried out in a single undivided acidic electrolytic medium, that is, in an acidic electrolytic medium as described below that is not divided into cathodic and anodic chambers by a membrane or other means. Because there is no division of the electrolytic chamber, the soluble hydrogen sulfide and ferrous ions are then transported to the anode, where they are oxidized: the sulfide to sulfur and the ferrous ion to ferric, virtually eliminating the problem of toxic gas production.
The Hydrogen Evolution Reaction (H.E.R.) is well known in the art, and is described, for instance in “Corrosion: Volume 1; Metal/Environmental Reactions” (L. L. Shrier, ed., Newnes-Butterworths, London/Boston, 1976) (hereinafter referred to as “Shrier”), pp. 9:41-9:52, the contents of which are hereby incorporated by reference herein. Overall, the reaction is described in Shrier as
H3O++e→½H2+H2O
which in turn is broken down into two steps: a discharge step and either a chemical desorption step or an electrochemical desorption step. The discharge step involves adsorption of hydrogen atoms at available sites on the cathode surface and is represented in Shrier as:
H3O++M+e→M−H+H2O
This is described as being followed by a transfer of adsorbed hydrogen across the interface into interstitial sites within the metal, represented as:
M−Hads.→Hinst.−M
In the chemical desorption step the adsorbed hydrogen atoms diffuse about on the metal surface until two of them collide, forming a hydrogen molecule which then escapes into the solution:
M−H+M−H→2M+H2
Electrochemical desorption is said to be more complex, involving reaction between an adsorbed hydrogen atom, a hydrated proton and an electron, in which for desorption to occur the proton must discharge onto a hydrogen atom adsorbed onto the metal surface. This is generally represented in Shrier as:
M−H+H3O++e→M+H2+H2O.
The preferred materials for the cathode in this invention comprise those metals that promote a relatively fast discharge step but that also promote a relatively slow chemical or electrochemical desorption step, and are described more specifically below. In working with a sulfur-containing ore or mineral such as chalcopyrite, in order to absorb the majority of the electrons produced by oxidizing H2S to S°, excess hydrogen must be produced at the cathode. The overall reaction of the process in the reactor, with respect to chalcopyrite, is:
2CuFeS2+6H+→4Cu2S+3S°+2Fe3++3H2
If excess hydrogen gas is evolved, for example during the treatment of chalcopyrite, the off-gases could be used as a source of energy to offset the electrical costs of the process. After the reduction process, the solid product may be treated with an additional oxidant, such as, but not limited to, hydrogen peroxide, or it may be simply heated, allowing the ferric ion to oxidize the chalcocite, dissolving the copper.
The ores and minerals for which the process may be used are those that contain one or more reducible phases but are not easily oxidizable. In general, the term “mineral” as used in the art means a (single) mineral phase (chalcopyrite or zinc ferrite) and the term “ore” is used to mean a mixture or aggregate of minerals. The process of this invention can be used with an ore, a mineral, or a mixture of minerals as the treated material. Metals whose values may be recovered by this process include copper, zinc, manganese, silver, and nickel. The ores and/or minerals may include sulfides. Some examples are chalcopyrite, bornite, pentlandite, pyrargyrite, zinc ferrites and manganese oxides.
The cathode electrode is comprised of materials that enhance the HER (hydrogen evolution reaction) at the cathode, preferably those metals that promote a relatively fast discharge step but that also promote a relatively slow chemical or electrochemical desorption step. Such metals include, for example, titanium, nickel, tantalum, molybdenum, aluminum, platinum, palladium, and iridium, including alloys, composites, and the like thereof. Such materials are described, for instance, in Shrier, Table 9.3, p. 9:48. In addition, the electrode materials must be resistant to sulfate ion or other anions that may be included in the acid media, and when treating sulfide ores such as chalcopyrite, promote sulfide oxidation at the anode. In the experiments that resulted in this invention, aluminum foam was used for the cathode and graphite [reticulated glassy carbon (RVC) or Dimensionally Stable™ Anodes] was used for the anode. However, the invention extends to the use of different materials that serve the same purpose, and the anode may comprise any suitable material that meets the above requirements. The same is the case for the material of the reactor. In the examples, a cylindrical glass reactor was employed; however the actual material need only be resistant to the medium and should be physically designed to maximize solid-liquid mixing.
The process is conducted in an acidic medium. This medium preferably comprises sulfuric acid or hydrochloric acid; however other acids such as glacial acetic acid may be used as long as they do not attack the electrodes or the reactor materials. High acid and sulfate concentrations (>1 M) may be needed to attain high reaction velocities in the reduction of chalcopyrite; however, the process may be generally conducted at lower acid concentrations.
Preferably the process is operated so as to maximize the contact between the solution and the ore or mineral particles as well as between the solution and the electrodes. The process may be operated at ambient temperature and pressure. As compared to prior techniques, the process of this invention produces excellent results under such conditions. However, it is not limited to these conditions; higher or lower temperature and/or pressure are not precluded and their use may increase the velocity of the process, as long as the aqueous solution remains a liquid. Evaporation is not prohibited, although fumes can be corrosive and thus should be controlled.
The product of the process of the invention comprises a reduced form of the metal contained in the ore and/or mineral fed to the process, and from that form the final metal values may be readily recovered. For instance, chalcopyrite (CuFeS2) is reduced to chalcocite (Cu2S) and/or Djurleite (Cu1.96S), which can readily be oxidized by known methods to produce copper. In the case of some ores or minerals, for example manganese oxides, the reduced-form product of the process of the invention will be metallic ions (e.g., Mn+2); in the case of others such as silver-containing materials the product may be the metal itself.
The following examples are representative of processes according to the invention but are not intended to limit the inventive concept in any way.
2.5 grams of the −200+270 mesh fraction of a chalcopyrite concentrate (21% Cu, 24% Fe, 7% Zn, 5% Pb and 32% S) was placed in a glass reactor containing 250 ml of 1 M H2SO4 in water at room temperature (23° C.). Current was passed through the stirred solution using an aluminum foam cathode (20 ppi) and a reticulated vitreous carbon (RVC) anode at a rate of 0.4 Amps for a period of 3 hours (a total of 4720 coulombs). After this time, the solution contained 2160 ppm Fe and 75 ppm Cu with an unquantified but sizeable amount of sulfur floating on the surface. The solid residue, containing 8.6% Fe and 24.4% Cu, was identified principally as chalcocite, copper and unreacted pyrite. From reconstructive head calculations, the conversion of chalcopyrite to chalcocite and copper was 99% and the current efficiency with respect to the total cell reaction and the iron reduction were 79 and 20%, respectively.
12.5 grams of chalcopyrite concentrate (21% Cu, 18% Fe and 32% S) was placed in a glass reactor containing 250 ml of 1 M H2SO4 in water at room temperature (23° C.). Current was passed through the stirred solution using an aluminum foam cathode (20 ppi) and a reticulated vitreous carbon (RVC) anode at a rate of 0.4 Amps for a period of 12 hours (a total of 17,280 coulombs). After this time, the solution contained 8158 ppm Fe and 220 ppm Cu with an unquantified but sizeable amount of sulfur floating on the surface. The solid residue, containing 7.7% Fe and 30.4% Cu, was identified principally as chalcocite, copper and unreacted pyrite and chalcopyrite. From reconstructive head calculations, the conversion of chalcopyrite to chalcocite and copper was at least 75% and the current efficiency with respect to the total cell reaction and the iron reduction were 81% and 20%, respectively.
2.5 grams of the −200+270 mesh fraction of a chalcopyrite concentrate (21% Cu, 24% Fe, 7% Zn, 5% Pb and 32% S) was placed in a glass reactor containing 250 ml of 1 M H2SO4 in water at room temperature (23° C.). Current was passed through the stirred solution using a platinum mesh cathode (2.5 cm×2.5 cm) and a reticulated vitreous carbon (RVC) anode at a rate of 0.4 Amps for a period of 2 hours and 40 minutes (a total of 4296 coulombs). After this time, the solution contained 1430 ppm Fe and negligible Cu with an unquantified but sizeable amount of sulfur floating on the surface. The conversion of chalcopyrite to chalcocite and copper ions was estimated at 66%. These values were lower than those obtained in Example 1 due to the greatly reduced cathodic surface area in this case.
2.5 grams of the −150+300 mesh fraction of a chalcopyrite concentrate (18% Cu, 23% Fe, and 30% S) was placed in a glass reactor containing 250 ml of 1 M H2SO4 in water at room temperature (22° C.). Current was passed through the stirred solution using an aluminum foam cathode (20 ppi) and a Dimensionally Stable Anode® at a rate of 0.7 Amps for a period of 5 hours (a total of 12600 coulombs). After this time, the solution contained 2120 ppm Fe and negligible Cu with an unquantified but sizeable amount of sulfur floating on the surface. The solid residue, containing 4.5% Fe and 35.9% Cu, was identified principally as chalcocite, copper and unreacted pyrite. From reconstructive head calculations, the conversion of chalcopyrite to chalcocite and copper was at least 98%.
2.5 grams of the −150+300 mesh fraction of a chalcopyrite concentrate (17% Cu, 22% Fe, and 30% S) was placed in a glass reactor containing 250 ml of 0.5 M H2SO4 in water at room temperature (22° C.). Current was passed through the stirred solution using an aluminum foam cathode (20 ppi) and a Dimensionally Stable Anode® at a rate of 0.7 Amps for a period of 6 hours (a total of 15,120 coulombs). After this time, the solution contained 1950 ppm Fe and negligible Cu with an unquantified but sizeable amount of sulfur floating on the surface. The solid residue, containing 5.9% Fe and 33% Cu, was identified principally as chalcocite, copper and unreacted pyrite. From reconstructive head calculations, the conversion of chalcopyrite to chalcocite and copper was at least 90%.
2.5 grams of the −150+300 mesh fraction of a chalcopyrite concentrate (17% Cu, 24% Fe, and 30% S) was placed in a glass reactor containing 250 ml of 1 M H2SO4 in water immersed in a controlled water bath at 40° C. Current was passed through the stirred solution using an aluminum foam cathode (20 ppi) and a Dimensionally Stable Anode at a rate of 0.7 Amps for a period of 3 hours (a total of 7,560 coulombs). After this time, the solution contained 2420 ppm Fe and negligible Cu with an unquantified but sizeable amount of sulfur floating on the surface. The solid residue, containing 2.7% Fe and 36% Cu, was identified principally as chalcocite, copper and unreacted pyrite. From reconstructive head calculations, the conversion of chalcopyrite to chalcocite and copper was at least 98%.
10 grams of manganese nodule DH-2 (18% Mn, 10% Fe, 0.15% Cu, 0.77% Ni and 0.11% Co) was placed in a glass reactor containing 250 ml of 1 M H2SO4 in water at room temperature. Current was passed through the stirred solution using an aluminum foam cathode (20 ppi) and a reticulated vitreous carbon (RVC) anode at a rate of 0.4 Amps for a period of 51/2 hours (a total of 7,920 coulombs). After this time, the solution contained 7180 ppm Mn, 3880 ppm Fe, 34 ppm Cu, 303 ppm Ni and 41 ppm Co. The solid residue contained 0.04% Mn, 0.52% Fe, 0.06% Cu, 0.01% Ni and 0.004% Co. From reconstructive head calculations, the dissolution of the different oxides was at least 99% for Mn, 95% for Fe, 58% for Cu, 99% for Ni and 96% for Co.
25 grams of zinc ferrite (18% Zn and 23% Fe) was placed in a glass reactor containing 250 ml of 1 M H2SO4 in water at room temperature (24° C.). Current was passed through the stirred solution using an aluminum foam cathode (20 ppi) and a graphite rod anode at a rate of 0.8 Amps for a period of 5 hours (a total of 14,400 coulombs). After this time, the solution contained 14,000 ppm Fe and 7,200 ppm Zn. From reconstructive head calculations, the zinc extraction was at least 40%.
10 grams of a refractory silver concentrate (only 50% silver extraction with cyanide), containing 2% Ag and 33% Fe, was placed in a glass reactor containing 250 ml of 1 M H2SO4 in water at room temperature (22° C.). Current was passed through the stirred solution using an aluminum foam cathode (20 ppi) and a graphite rod as the anode at a rate of 0.7 Amps for a period of 5 hours (a total of 12,600 coulombs). After this time, the solution contained 280 ppm Fe. The solid phase was filtered and placed in contact at room temperature (22° C.) with a 250 ml of a solution containing 0.2M thiourea of which 12% had previously been converted to formamidine disulfide. After constant stirring, the silver extraction was estimated at 54, 60 and 89% after 1, 5 and 21 hours of contact, respectively. The untreated concentrate showed a silver extraction of only 8, 20 and 81% after 1, 5 and 21 hours, respectively under the same conditions.
All publications and patent applications cited in this specification are herein incorporated by reference as if each individual publication or patent application were specifically and individually indicated to be incorporated by reference.
Although the foregoing invention has been described in some detail by way of illustration and example for purposes of clarity of understanding, it will be readily apparent to those of ordinary skill in the art in light of the teachings of this invention that certain changes and modifications may be made thereto without departing from the spirit or scope of the appended claims.
| Filing Document | Filing Date | Country | Kind | 371c Date |
|---|---|---|---|---|
| PCT/US08/54661 | 2/22/2008 | WO | 00 | 9/18/2009 |
| Number | Date | Country | |
|---|---|---|---|
| 60903195 | Feb 2007 | US |
