BR112019008080B1 - USE OF A REAGENT WITH A THIOCARBONIL FUNCTIONAL GROUP, AND, METHOD OF RECOVERING AT LEAST ONE BASE METAL ION FROM A MATERIAL COMPRISING AT LEAST ONE BASE METAL SULPHIDE - Google Patents

Abstract

This application relates to methods of recovering metals from metal sulfides that involve contacting the metal sulfide with an acidic sulfate solution containing ferric sulfate and a reagent having a thiocarbonyl functional group, where the concentration of the reagent in the solution acid sulfate is sufficient to increase the rate of metal ion extraction relative to an acidic sulfate solution that does not contain the reagent, to produce a turgid solution containing the metal ions.

Description

FIELD OF THE INVENTION

[001] This description relates to methods for leaching metals from ores containing metal sulfides. More particularly, it relates to a hydrometallurgical process for extracting base metals from ores containing base metal sulfides using reagents having a thiocarbonyl functional group. This disclosure further relates to the recovery of reagents having a thiocarbonyl functional group from a leach solution charged for recirculation to a hydrometallurgical process for extracting base metals from ores containing base metal sulfides using such reagents. This disclosure further relates to methods for recovering catalysts from spent leach materials, and in particular the recovery of reagents with a thiocarbonyl functional group from spent leach materials containing base metal sulfides from which the base metal has been leached.

Description of the Related Technique

[002] The aqueous processing of minerals has several advantages over pyrometallurgical approaches, particularly when dealing with complex and/or low-grade ores. The main disadvantage of hydrometallurgical processes, when applied to various metal sulfide minerals, is the low extraction rates that are observed. It is desirable to develop a process in which high metal extractions can be achieved at timescales that are of industrial interest.

[003] Chalcopyrite, for example, is a semiconductor and therefore corrodes electrochemically in oxidizing solutions. In ferric sulfate medium, the general leaching reaction is as follows: This reaction can be represented as a combination of half-cell anodic and cathodic reactions: Half-cell anodic reaction: Half-cell cathode reaction:

[004] A fundamental problem with chalcopyrite oxidation is that chalcopyrite mineral surfaces become resistant to electrochemical degradation at solution potentials above a certain level (generally considered to be about 550 to 600 mV vs Ag/AgCl). It is widely held that this results from the formation of some type of passivating film on the mineral surface which probably consists of an altered form of chalcopyrite, partially depleted of Fe. It is desirable to provide leaching processes in which such passivation is reduced or avoided.

[005] Some work has been carried out in extractive hydrometallurgy to recover precious metals, such as gold and silver, from copper concentrates or chalcopyrite residues after copper extraction. Deschênes and Ghali (Hydrometallurgy 20: 129-202) demonstrated the potential invention of thiourea in the acid sulfate leaching of sulfide concentrates, such as those containing chalcopyrite, to selectively recover gold and silver. Thiourea is an organosulfide compound with a thiocarbonyl functional group. However, thiourea did not appear to have an effect on copper recovery from copper sulfides.

[006] The leaching of metals in the presence of halogens has also been widely investigated in recent decades. The use of chloride at elevated temperature can result in high copper recoveries (Winand, Hydrometallurgy, 27: 285-316) from chalcopyrite. Room temperature leaching of chlorides has also been shown to be effective, making them suitable for heap leaching (WO2015059551). Bromide leaching has been investigated mainly for gold (Li et al. Proceedings of the 3rd Pan American Materials Congress, 2017: 653-660). However, several technologies also demonstrate their beneficial effect in extracting copper from sulphide ores (US5989311, US9290827). Iodide leaching has also been shown to be effective under several conditions (US5989311, US8163063, US8287623, and US8865119).

SUMMARY

[007] This disclosure relates, at least in part, to the unexpected discovery that various reagents comprising a thiocarbonyl functional group (e.g., thiourea) can be used to facilitate the leaching of metal from various metal sulfides ( e.g., chalcopyrite copper) with acidic leach solutions, e.g., an acid sulfate leach solution or a halide leach solution. When added in small amounts, such reagents can increase the rate of metal leaching over that observed in their absence.

[008] This description relates to a method for recovering at least one metal from at least one metal sulfide in an ore, the method comprising: contacting the ore with an acid sulfate solution containing ferric sulfate and a reagent with a functional group thiocarbonyl to produce a charged solution containing metal ions; and recovering the at least one metal from the charged solution, wherein the at least one metal includes: copper, wherein at least one metal sulfide includes chalcopyrite, covellite, bornite, enargite, a copper sulfide of the formula CuxSy wherein the ratio x:y is between 1 and 2, or a combination thereof; cadmium, wherein the at least one metal sulfide is greenockite; nickel, in which at least one metal sulfide is pentlandite, violarite or a combination thereof; or a combination thereof.

[009] This description relates to a method for recovering at least one metal from at least one metal sulfide in a concentrate, the method comprising: contacting the concentrate with an acid sulfate solution containing a reagent with a thiocarbonyl functional group to produce a charged solution containing metal ions; and recovering the at least one metal from the charged solution, wherein the at least one metal includes: copper, wherein at least one metal sulfide includes chalcopyrite, covellite, bornite, enargite, a copper sulfide of the formula CuxSy wherein the ratio x:y is between 1 and 2, or a combination thereof; cadmium, wherein the at least one metal sulfide is greenockite; nickel, in which at least one metal sulfide is pentlandite, violarite or a combination thereof; or a combination thereof.

[0010] This description relates to a method for recovering at least one metal from at least one metal sulfide in a material, the method comprising: contacting the material with an acid sulfate solution containing a reagent having a thiocarbonyl functional group to produce a charged solution containing metal ions; and recovering the at least one metal from the charged solution, wherein the at least one metal includes: copper, wherein at least one metal sulfide includes chalcopyrite, covellite, bornite, enargite, a copper sulfide of the formula CuxSy wherein the ratio x:y is between 1 and 2, or a combination thereof; cadmium, wherein the at least one metal sulfide is greenockite; nickel, in which at least one metal sulfide is pentlandite, violarite or a combination thereof; or a combination thereof.

[0011] The concentrate, ore or other material may be provided as coarse particles. Coarse particles may be agglomerated particles.

[0012] In the methods described above, the concentration of the reagent in the acid sulfate solution can be in the range of about 0.2 mM to about 100 mM, from about 0.2 mM to about 20 mM, about 0 .2 mM to about 10 mM, about 0.2 mM to about 5 mM, about 0.2 mM to about 4 mM, about 0.2 mM to about 3 mM, about 0.2 mM mM to about 2 mM, about 0.2 mM to about 1.5 mM, about 0.2 mM to about 1.0 mM, or about 0.2 mM to about 0.5 mM.

[0013] Where the metal is a copper sulfide of the formula CuxSy where the x:y ratio is between 1 and 2, the copper sulfide may include chalcocite, djurleite, digenite or a combination thereof.

[0014] In the methods described above, the reagent can be thiourea (Tu), ethylene thiourea (Etu), thioacetamide (TA), sodium dimethyldithiocarbamate (SDDC), ethylene trithiocarbonate (ETC), thiosemicarbazide (TSCA), or a combination thereof.

[0015] This description further relates to a method for recovering at least one metal from at least one metal sulfide in an ore, the method comprising: contacting the ore with an acid sulfate solution comprising ferric sulfate and formadin disulfide (FDS ) to produce a charged solution containing metal ions; and recovering the metal from the charged solution, wherein the at least one metal includes: copper, wherein the at least one metal sulfide includes chalcopyrite, covellite, bornite, enargite, a copper sulfide of the formula CuxSy wherein the ratio x:y is between 1 and 2, or a combination thereof; cadmium, wherein the at least one metal sulfide is greenockite; nickel, in which at least one metal sulfide is pentlandite, violarite or a combination thereof; or a combination thereof.

[0016] This description further relates to a method for recovering at least one metal from at least one metal sulfide in a concentrate, the method comprising: contacting the concentrate with an acid sulfate solution comprising ferric sulfate and formadin disulfide (FDS ) to produce a charged solution containing the metal ions; and recovering the metal from the charged solution, wherein the at least one metal includes: copper, wherein the at least one metal sulfide includes chalcopyrite, covellite, bornite, enargite, a copper sulfide of the formula CuxSy wherein the ratio x:y is between 1 and 2, or a combination thereof; cadmium, wherein the at least one metal sulfide is greenockite; nickel, in which at least one metal sulfide is pentlandite, violarite or a combination thereof; or a combination thereof.

[0017] This description further relates to a method for recovering at least one metal from at least one metal sulfide in a material, the method comprising: contacting the material with an acid sulfate solution comprising ferric sulfate and formadin disulfide (FDS ) to produce a charged solution containing the metal ions; and recovering the metal from the charged solution, wherein the at least one metal includes: copper, wherein the at least one metal sulfide includes chalcopyrite, covellite, bornite, enargite, a copper sulfide of the formula CuxSy wherein the ratio x:y is between 1 and 2, or a combination thereof; cadmium, wherein the at least one metal sulfide is greenockite; nickel, in which at least one metal sulfide is pentlandite, violarite or a combination thereof; or a combination thereof.

[0018] The concentrate, ore or other material can be provided as coarse particles. Coarse particles may be agglomerated particles.

[0019] The concentration of SDS in the acid sulfate solution may be in the range of about 0.1 mM to about 50 mM, about 0.1 mM to about 15 mM, about 0.1 mM to about 10 mM, about 0.2 mM to about 5 mM, about 0.1 mM to about 2.5 mM, about 0.1 mM to about 2 mM, about 0.1 mM to about 1.5 mM, about 0.1 mM to about 1.0 mM, about 0.1 mM to about 0.5 mM, or about 0.1 mM to about 0.25 mM. Where the metal is a copper sulfide of the formula CuxSy where the x:y ratio is between 1 and 2, the copper sulfide may include chalcocite, djurleite, digenite or a combination thereof.

[0020] The concentration of SDS in the acid sulfate solution may be sufficient to provide sufficient thiourea to increase the metal ion extraction rate relative to an acid sulfate solution that does not contain the reagent to produce the loaded leach solution containing the metal ions.

[0021] In the methods described above, in which the ore can be provided as coarse particles, which can be agglomerated particles. Ferric ions can be used to oxidize metal sulfide. In the methods described above, ferric ions can be generated at least in part by bacteria.

[0022] The methods may involve percolation leaching. Percolation leaching can be a heap leaching. Percolation leaching can be a vat leaching. The leaching may be a tank leaching.

[0023] Metal recovery from the charged leach solution may include solvent extraction and electrolytic extraction.

[0024] In the methods described above, the acid sulfate solution may comprise halide ions. The halide ions comprise chloride ions, bromide ions, iodide ions, or a combination thereof. The chloride concentration in the acid sulfate solution can be about 20 g/L or less, about 50 g/L or less, about 80 g/L or less, about 20 g/L or less, in a range from about 20 g/L to about 120 g/L, in a range of about 20 g/L to about 80 g/L, or in a range of about 20 g/L to about 50 g/L L. The iodide concentration in the acid sulfate solution can be about 300 ppm or less, about 100 ppm or less, or in the range of about 100 ppm to about 300 ppm. The bromide concentration in the acid sulfate solution can be about 10 g/L or less, about 30 g/L or less, or in the range of about 10 g/L to about 30 g/L.

[0025] This description relates to the use of a reagent having a thiocarbonyl functional group for the extraction of at least one base metal from at least one base metal sulfide in a material. The reagent can be, but is not necessarily limited to, thiourea (Tu), ethylene thiourea (ETu), thioacetamide (TA), sodium dimethyldithiocarbamate (SDDC), ethylene trithiocarbonate (ETC), thiosemicarbazide (TSCA), or combinations thereof. same. The reagent concentration can be in the range of about 0.2 mM to about 100 mM, or in the range of about 0.2 mM to about 30 mM.

[0026] The description further relates to the use offormamidine disulfide (FDS) for extracting at least one base metal from at least one base metal sulfide in a material.

[0027] The SDS can be at a concentration in the range of about 0.1 mM to about 50 mM, or in the range of about 0.1 mM to about 15 mM.

[0028] In the uses described above, the at least one base metal may include copper, cadmium, nickel or a combination thereof. The at least one base metal may comprise: copper, wherein the at least one base metal sulfide is chalcopyrite, covellite, bornite, enargite, a copper sulfide of the formula CuxSy wherein the x:y ratio is between 1 and 2, or a combination thereof; cadmium, wherein the at least one parent metal sulfide is greenockite; nickel, wherein at least one base metal sulfide is pentlandite, violarite or a combination thereof; or a combination thereof.

[0029] The material may be an ore or a concentrate.

[0030] Such use can be made in the presence of halide ions. Halide ions can include chloride ions, bromide ions, iodide ions, or a combination thereof. The chloride concentration in the acid sulfate solution can be about 20 g/L or less, about 50 g/L or less, about 80 g/L or less, about 20 g/L or less, in a range from about 20 g/L to about 120 g/L, in a range of about 20 g/L to about 80 g/L, or in a range of about 20 g/L to about 50 g/L L. The iodide concentration in the acid sulfate solution can be about 300 ppm or less, about 100 ppm or less, or in the range of about 100 ppm to about 300 ppm. The bromide concentration in the acid sulfate solution can be about 10 g/L or less, about 30 g/L or less, or in the range of about 10 g/L to about 30 g/L.

[0031] This description further relates to a method for recovering a reagent with a thiocarbonyl functional group from an aqueous charged leach solution (PLS), wherein the aqueous PLS comprises the reagent and base metal ions, wherein a portion of the reagent is complexed with base metal ions, the method comprising: mixing the PLS with an organic solvent containing a base metal ion extractor to form a mixture; extracting the base metal ions from the PLS in the organic solvent; and separating the mixture into a base metal ion depleted raffinate comprising the reagent and a base metal ion enriched organic phase comprising the organic solvent and base metal ions. Extracting the base metal ions from the PLS into the organic solvent may comprise decomplexing the reagent from the base metal ions to increase the amount of free reagent in the raffinate compared to the PLS. The reagent can be thiourea (Tu), ethylene thiourea (ETu), thioacetamide (TA), sodium dimethyldithiocarbamate (SDDC), ethylene trithiocarbonate (ETC), thiosemicarbazide (TSCA), or a combination thereof. The raffinate may further comprise Formamidine Disulfide (SDF), in which case the method may further comprise contacting the raffinate with a reducing agent to reduce SDS to Tu. Contacting the raffinate with a reducing agent to reduce SDF to Tu may comprise reducing SDS to obtain a Tu:SDF ratio in the range of from about 0.5:1 to about 9:1. The reducing agent can be H2S, SO2, or NaSH.

[0032] This description further relates to a method for recovering SDS from an aqueous charged leach solution (PLS), wherein the aqueous SLP comprises the reagent and base metal ions, the method comprising: mixing the SLP with a solvent organic containing a base metal ion extractor to form a mixture; extracting the base metal ions from the PLS in the organic solvent; and separating the mixture into a base metal ion depleted raffinate comprising SDS and a base metal ion enriched organic phase comprising the organic solvent and base metal ions.

[0033] The base metal ions may include cadmium, nickel, copper or a combination thereof.

[0034] The organic solvent may be an aliphatic solvent, an aromatic solvent or a combination thereof. The organic solvent can be kerosene, aromatic alkyls, cycloparaffins or a combination thereof.

[0035] The base metal ion extractor may be an aldoxime, a ketoxime or a combination thereof. The base metal ion extractor may further comprise an ester modifier, an alkylphenol modifier, or a combination thereof.

[0036] PLS can additionally comprise Tu complexed with base metal ions, and extracting the base metal ions from PLS comprises decomplexing Tu from base metal ions to increase the amount of free Tu in the raffinate compared to PLS .

[0037] This description further relates to a method for recovering at least one base metal from at least one base metal sulfide in a material containing the at least one base metal sulfide, the method comprising: contacting the material with a leachate, wherein the leachate comprises an acid sulfate solution containing ferric sulfate and a reagent having a thiocarbonyl functional group, for extracting base metal ions from the at least one base metal sulfide to produce a charged leach solution (PLS ); mixing the PLS with an organic solvent containing a base metal ion extractor to form a mixture; extracting base metal ions from the PLS into the organic solvent; and separating the mixture into a base metal ion depleted raffinate comprising the reagent and a base metal ion enriched organic phase comprising the organic solvent and base metal ions.

[0038] Extracting the base metal ions from PLS into the organic solvent comprises decomplexing the reagent from base metal ions to increase the amount of free reagent in the raffinate compared to PLS. The reagent can be, but is not necessarily limited to, thiourea (Tu), ethylene thiourea (ETu), thioacetamide (TA), sodium dimethyldithiocarbamate (SDDC), ethylene trithiocarbonate (ETC), thiosemicarbazide (TSCA), or combinations thereof. same. Where the reagent comprises Tu, the raffinate may additionally compriseformamidine disulfide (FDS), where the method further comprises contacting the raffinate with a reducing agent to reduce FDS to Tu. Contacting the raffinate with a reducing agent to reduce SDF to Tu may comprise reducing SDS to obtain a Tu:SDF ratio in the range of from about 0.5:1 to about 9:1. The reducing agent can be H2S, SO2, or NaSH.

[0039] This description further relates to a method for recovering at least one base metal from at least one base metal sulfide in a material containing the at least one base metal sulfide, the method comprising: contacting the material with a leachate, wherein the leachate comprises an acid sulfate solution containing ferric sulfate and formadin disulfide (FDS), for extracting base metal ions from the at least one base metal sulfide to produce a charged leach solution (PLS) ; mixing the PLS with an organic solvent containing a base metal ion extractor to form a mixture; extracting base metal ions from the PLS into the organic solvent; and separating the mixture into a base metal ion depleted raffinate comprising the reagent and a base metal ion enriched organic phase comprising the organic solvent and base metal ions. The PLS may further comprise thiourea (Tu) complexed with base metal ions, wherein the method further comprises extracting the base metal ions from the PLS comprises decomplexing the Tu from base metal ions to increase the amount of free Tu in the raffinate compared to PLS. The method may further comprise contacting the raffinate with a reducing agent to reduce SDS to Tu. Contacting the raffinate with a reducing agent to reduce SDF to Tu may comprise reducing SDS to obtain a Tu:SDF ratio in the range of from about 0.5:1 to about 9:1. The reducing agent can be H2S, SO2, or NaSH.

[0040] The organic solvent may be an aliphatic solvent, an aromatic solvent or a combination thereof. The organic solvent can include kerosene, aromatic alkyls, cycloparaffins or a combination thereof. Base metal ions can include cadmium, nickel or copper. The base metal ion extractor can be an aldoxime, a ketoxime or a combination thereof.

[0041] The base metal ions may include cadmium, nickel, copper or a combination thereof.

[0042] The base metal ion extractor may be an aldoxime, a ketoxime or a combination thereof. The base metal ion extractor may further comprise an ester modifier, an alkylphenol modifier, or a combination thereof.

[0043] The leachate and/or the PLS may comprise halide ions. Halide ions can include chloride ions, bromide ions, iodide ions, or a combination thereof. The chloride concentration in the leachate or PLS can be about 20 g/L or less, about 50 g/L or less, about 80 g/L or less, about 20 g/L or less, in a range from about 20 g/L to about 120 g/L, in a range of about 20 g/L to about 80 g/L, or in a range of about 20 g/L to about 50 g/L L. The iodide concentration in the leachate or PLS can be about 300 ppm or less, about 100 ppm or less, or in the range of about 100 ppm to about 300 ppm. The concentration of bromide in the leachate or PLS can be about 10 g/L or less, about 30 g/L or less, or in a range from about 10 g/L to about 30 g/L.

[0044] The methods may further comprise recirculating a portion of the raffinate comprising the reagent with a thiocarbonyl functional group to the leachate. The leachate comprising the portion of the raffinate that is recirculated from the solvent extraction can be supplemented with fresh reagent having a thiocarbonyl functional group to obtain the desired concentration of reagent having a thiocarbonyl functional group in the leachate.

[0045] This disclosure further relates to a method for recovering a reagent comprising a thiocarbonyl functional group sequestered in leach materials comprising at least one base metal sulfide, the method comprising rinsing the leach materials with a wash solution comprising base metal ions to produce a loaded wash solution (PWS) comprising the reagent. The method may further comprise: mixing the PWS with an organic solvent containing a base metal ion extractor to form a mixture; extracting the base metal ions from the PWS in the organic solvent; and separating the mixture into a base metal ion depleted solution comprising the reagent and a base metal ion enriched solution comprising the organic solvent and base metal ions. Extracting the base metal ions from the PWS in the organic solvent comprises decomplexing the base metal ion reagent to increase the amount of free reagent in the spent base metal ion solution compared to the PWS. The organic solvent can include an aliphatic solvent, an aromatic solvent or a combination thereof. The organic solvent may comprise kerosene, aromatic alkyls, cycloparaffins or a combination thereof. The reagent may include, but is not necessarily limited to, thiourea (Tu), ethylene thiourea (ETu), thioacetamide (TA), sodium dimethyldithiocarbamate (SDDC), ethylene trithiocarbonate (ETC), thiosemicarbazide (TSCA), or a combination of the same. Where the reagent comprises Tu, the base metal ion depleted solution further comprises SDS, where the method may further comprise contacting the base metal ion depleted solution with a reducing agent to reduce SDS to Tu. Contacting the base metal ion depleted solution with a reducing agent to reduce SDF to Tu comprises reducing SDS to obtain a Tu:SDF ratio in the range of from about 0.5:1 to about 9:1. The reducing agent can be H2S, SO2, or NaSH.

[0046] The organic solvent may be an aliphatic solvent, an aromatic solvent or a combination thereof. The organic solvent can include kerosene, aromatic alkyls, cycloparaffins or a combination thereof. Base metal ions can include cadmium, nickel or copper. The base metal ion extractor can be an aldoxime, a ketoxime or a combination thereof.

[0047] The base metal ions may include cadmium, nickel, copper or a combination thereof.

[0048] The base metal ion extractor may be an aldoxime, a ketoxime or a combination thereof. The base metal ion extractor may further comprise an ester modifier, an alkylphenol modifier, or a combination thereof.

[0049] The base metal ion concentration in the wash solution may be at least 100 ppm, at least 400 ppm, or at least 1000 ppm.

[0050] The method may additionally include, prior to rinsing the leach materials with the washing solution, rinsing the leach materials with an acidic solution. The acidic solution can have a pH of around 1.8.

[0051] The description further relates to a method for recovering at least one base metal from a material containing at least one base metal sulfide, the method comprising: recovering a reagent comprising a thiocarbonyl functional group sequestered in leach materials comprising at least one base metal sulfide according to a method as described above; mixing the recovered agent with an acid sulfate solution containing ferric sulfate to form a leachate; contacting the material with the leachate to extract base metal ions from the at least one base metal sulfide to produce a charged leach solution (PLS) comprising base metal ions. The acid sulfate solution, prior to mixing with the recovered agent, may comprise a pre-existing reagent comprising a thiocarbonyl functional group, pre-existing SDS, or a combination thereof. The preexisting reagent is thiourea (Tu), thioacetamide (TA), sodium dimethyldithiocarbamate (SDDC), ethylene trithiocarbonate (ETC), thiosemicarbazide (TSCA), or a combination thereof. The method may further comprise: mixing the PLS with an organic solvent containing a base metal ion extractor to form a mixture; extracting base metal ions from the PLS into the organic solvent; and separating the mixture into a base metal ion depleted raffinate comprising the reagent and a base metal ion enriched solution comprising the organic solvent and base metal ions. Extracting the base metal ions from PLS into the organic solvent comprises decomplexing the reagent from base metal ions to increase the amount of free reagent in the raffinate compared to PLS. Where the reagent is Tu, the raffinate may further comprise SDS, where the method may further comprise contacting the raffinate with a reducing agent to reduce SDS to Tu. Contacting the raffinate with a reducing agent to reduce SDF to Tu may comprise reducing SDS to obtain a Tu:SDF ratio in the range of from about 0.5:1 to about 9:1. The reducing agent is H2S, SO2, or NaSH.

[0052] The organic solvent may be an aliphatic solvent, an aromatic solvent or a combination thereof. The organic solvent can include kerosene, aromatic alkyls, cycloparaffins or a combination thereof. Base metal ions can include cadmium, nickel or copper. The base metal ion extractor can be an aldoxime, a ketoxime or a combination thereof.

[0053] The base metal ions may include cadmium, nickel, copper or a combination thereof.

[0054] The base metal ion extractor may be an aldoxime, a ketoxime or a combination thereof. The base metal ion extractor may further comprise an ester modifier, an alkylphenol modifier, or a combination thereof.

[0055] The leachate and/or the PLS may comprise halide ions. Halide ions can include chloride ions, bromide ions, iodide ions, or a combination thereof. The chloride concentration in the leachate or PLS can be about 20 g/L or less, about 50 g/L or less, about 80 g/L or less, about 20 g/L or less, in a range from about 20 g/L to about 120 g/L, in a range of about 20 g/L to about 80 g/L, or in a range of about 20 g/L to about 50 g/L L. The iodide concentration in the leachate or PLS can be about 300 ppm or less, about 100 ppm or less, or in the range of about 100 ppm to about 300 ppm. The concentration of bromide in the leachate or PLS can be about 10 g/L or less, about 30 g/L or less, or in a range from about 10 g/L to about 30 g/L.

[0056] Other aspects and features of the present invention will become apparent to those skilled in the art upon review of the following description of specific embodiments of the invention in conjunction with the accompanying figures.

BRIEF DESCRIPTION OF THE DRAWINGS

[0057] In the drawings illustrating embodiments of the invention, Figure 1 is a flowchart of the recovery of a leaching process according to embodiments of the invention; Figure 2 is a flowchart of the recovery of a leaching process in accordance with embodiments of the invention which involves a reduction step prior to recirculation of the raffinate to the leachate; Figure 3 is a graph showing the effect of thiourea concentration on mixed potential and dissolution (idisolve) current density of the CuFeS2 electrode; Figure 4 is a bar graph showing the electrochemical dissolution rates of a CuFeS2 electrode in sulfuric acid solution at pH 2 and 25°C with varying initial concentrations of thiourea, formadin disulfide (FDS), and Fe(III ); Figure 5 is a schematic diagram for the leaching column used in connection with the leaching experiments belonging to Figures 4, 5 and 6; Figure 6 is a graph showing the effect of thiourea concentration on copper leaching from Ore A in column leaching experiments; Figure 7 is a graph showing the effect of thiourea concentration on copper leaching from Ore B in column leaching experiments; Figure 8 is a graph showing the effect of thiourea concentration on copper leaching from Ore C in column leaching experiments; Figure 9 is a graph showing the effect of thiourea concentration on the rate of copper leaching from Ore C in column leaching experiments; Figure 10 is a graph showing the effect of thiourea concentration on ORP over time; Figure 11 is a graph showing the effect of thiourea concentration on copper dissolution for coarse A Ore in roller bottle experiments; Figure 12 is a graph showing the effect of thiourea concentration on copper dissolution for Coarse B Ore in roller bottle experiments; Figure 13 is a graph showing the effect of Tu addition on various Cu(I) containing minerals. Diamonds refer to bornite, triangles refer to covellite, inverted triangles refer to chalcocite, and squares refer to chalcopyrite. Open symbols refer to control treatments without Tu, while solid symbols refer to mineral-treated solutions with an initial Tu concentration of 2 mM; Figure 14 is a graph showing the effect of Tu on cadmium extraction from greenockite; Figure 15 is a graph showing the effect of Tu on copper extraction from enargite; Figure 16 is a graph showing the effect of Tu on violarite nickel extraction; Figure 17 is a graph showing the percentage of Cu ions remaining in the solution after various amounts of Tu addition; Figure 18 is a graph showing Cu extraction from chalcopyrite under various Tu dosages; Figure 19 is a graph showing the relationship between Tu dosage and Cu extraction after 172 hours; Figure 20 is a graph showing copper leaching from chalcopyrite in stirred reactor assays using reagents comprising thiocarbonyl functional groups. Circles refer to Tu, triangles refer to TA, inverted triangles refer to SDDC, diamonds refer to ETC, stars refer to TSCA, and squares refer to controls; Figure 21 is a graph showing copper leaching from covellite in stirred reactor assays using reagents comprising thiocarbonyl functional groups. Circles refer to Tu, triangles refer to TA, diamonds refer to SDDC, and squares refer to controls; Figure 22 is a graph showing the leaching of copper from bornite in stirred reactor tests using reagents comprising thiocarbonyl functional groups. Triangles refer to Tu, circles refer to TA, and squares refer to controls; Figure 23 is a graph showing copper leaching from enargite in stirred reactor assays using reagents comprising thiocarbonyl functional groups. Circles refer to Tu, triangles refer to TA, inverted triangles refer to ETC, and squares refer to controls; Figure 24 is a graph showing copper leaching from chalcopyrite in stirred reactor assays using reagents comprising thiocarbonyl, urea, and carbon disulfide functional groups. Circles refer to urea, triangles refer to controls, inverted triangles refer to TA, diamonds refer to Tu, stars refer to ETC, and squares refer to carbon disulfide ; Figure 25a is a graph comparing copper leaching from chalcopyrite (circles) or bornite (triangles) using leaching solutions with an initial concentration of 2mM Tu (solid symbols) or an initial concentration of 1mM SDF (open symbols) ; Figure 25b is a graph comparing copper leaching from covellite (circles) or chalcocite (triangles) using leach solutions with an initial concentration of 2mM Tu (solid symbols) or an initial concentration of 1mM SDF (open symbols) ; Figure 26 is a graph monitoring bacterial activity and SDS content with ORP and HPLC; and Figure 27 is a graph showing bioleaching of CuFeS2 using only Fe3+ (day 0 to 50) and using Fe3+ + Tu (day 90 to 150) in closed-loop experiments; Figure 28 are graphs showing copper leaching from chalcopyrite in the presence of Tu with varying chloride concentrations; Figure 29 are graphs showing copper leaching from chalcopyrite in the presence of (a) Tu and (b) ETu with varying chloride concentrations; Figure 30 are graphs showing copper leaching from chalcopyrite in the presence of (a) Tu and (b) ETu with varying concentrations of bromide; Figure 31 are graphs showing the leaching of copper from chalcopyrite with Tu or ETu in the presence of (a) 100 ppm iodine and (b) 300 ppm iodine in a sealed reactor; Figure 32 are graphs showing iodine concentration in a sealed reactor over time in the presence or absence of Tu and ETu at (a) 100 ppm iodine and (b) 300 ppm iodine; Figure 33 is a graph showing iodine concentration in an open air reactor over time in the presence or absence of Tu; Figure 34 are graphs showing the leaching of copper from chalcopyrite with Tu or ETu in the presence of (a) 100 ppm iodine and (b) 300 ppm iodine in an unsealed reactor (ie, open air); Figure 35 are graphs showing iodine concentration in an unsealed (i.e., open air) reactor over time in the presence or absence of Tu and ETu at (a) 100 ppm iodine and (b) 300 ppm of iodine; Figure 36 is a bar diagram showing free Tu equivalents in a simulated PLS and in the resulting simulated raffinate after solvent extraction; Figure 37 is a bar diagram showing free ETu in simulated PLS and the resulting simulated raffinate after solvent extraction; Figure 38 is a graph of total thiourea concentration in effluent versus time for three ores during irrigation with a solution having an equivalent Tu concentration of 2 mM; Figure 39 is a graph of total thiourea concentration in effluent versus time for three ore samples in Figure 38 during acid water washing; Figure 40 is a bar diagram showing the amount of Tu equivalent remaining in columns of three ore samples after various treatments.

DETAILED DESCRIPTION

[0058] This disclosure relates to methods for recovering base metals from base metal sulfide minerals, and relates in particular to the unexpected discovery that various reagents with a thiocarbonyl functional group, for example thiourea ("Tu" , also known as thiocarbamide), can be used to facilitate base metal leaching of base metal sulfides in various minerals with acid sulfate leaching solutions, even in the presence of halide species. Such reagents can increase the rate of metal sulfide leaching.

[0059] Other aspects of this disclosure relate to the recovery of reagents with a thiocarbonyl functional group from the charged leach solution (“PLS”) for recirculation to the leach solution (ie, the leachate). Such recirculation can provide the advantage of reducing the amount of new reagent that must be added to the leachate over time.

[0060] The skilled person will understand that there is an equilibrium between Tu and Formamidine Disulfide (FDS) in solution. The equilibrium between FDS and Tu in solution can be described by the following equation:

[0061] Tu provides a stronger effect in enhancing the leaching of base metals from materials containing base metal sulfides. For example, copper leaches more readily from sulfide ores/concentrates in the presence of TU than SDS or the TU-Cu complex. Therefore, the leaching process will be intensified by recirculating a solution with higher TU free for leaching. Consequently, more particular aspects of this disclosure relate to the addition of a reducing agent to the raffinate comprising Tu (Tu) and Formamidine Disulfide (FDS) to bias the equilibrium in favor of Tu prior to recirculation to the leach solution.

[0062] This disclosure also relates to methods for recovering catalysts from spent leach materials. More particularly, it relates to the recovery of reagents having a thiocarbonyl functional group from spent leach materials containing base metal sulfides from which the base metal has been leached.

[0063] “Base metal”, as used herein, refers to non-ferrous metals, excluding precious metals. These can include copper, lead, nickel and cadmium. These may additionally include zinc, aluminum, tin, tungsten, molybdenum, tantalum, cobalt, bismuth, cadmium, titanium, zirconium, antimony, manganese, beryllium, chromium, germanium, vanadium, gallium, hafnium, indium, niobium, rhenium and thallium.

[0064] Such methods can be particularly useful in recovering metal from low-grade ores that do not contain the parent metal sulfide mineral in high proportions. The method involves contacting the base metal sulfide mineral with an acid sulfate solution containing the reagent having a thiocarbonyl functional group.

[0065] One skilled in the art further understands that just because a reagent having a thiocarbonyl functional group may be useful in extracting a base metal from a metal sulfide, or mineral containing that metal sulfide, does not mean that that reagent will be useful in extracting of the same metal as other metal sulphides comprising the metal.

Minerals Chalcopyrite (CuFeS2)

[0066] Chalcopyrite leaching is carried out in the ferric acid sulfate solution according to the following reaction formula:

Covellite (CuS)

[0067] The leaching of covellite in the ferric sulfate solution proceeds according to the following reaction formula:

Chalcocite (Cu2S)

[0068] The leaching of chalcocite in the ferric solution proceeds according to the following formula:

[0069] The skilled person understands that “chalcocite” ores often contain a mixture of minerals with the formula CuxSy, where the ratio x:y is between 1 and 2. Additional minerals within this formula include digenite and djurleite.

Bornite (Cu5FeS4)

[0070] Bornite is an important copper mineral that usually coexists with chalcopyrite. The bornite leaching process in ferric solution is described in two stages:

Enargite (Cu3AsS4)

[0071] Unlike the other copper minerals mentioned above (chalcopyrite, covellite, chalcocite and bornite), the copper in enargite is primarily Cu(II) rather than Cu(I). The difference in copper oxidation state will also influence its leaching kinetics under catalyzed conditions. A previous study showed that enargite leaching at atmospheric pressure is extremely slow. The dissolution of enargite in ferric sulfate media can take several routes. Two of them are described below:

Greenockite (CdS)

[0072] The metal and cadmium compounds are mainly used for alloys, coatings, batteries and plastic stabilizers. There are no mines specifically for cadmium extraction. Cadmium sulfide is commonly associated with zinc sulfides and is recovered as a by-product of zinc leaching from roasted sulfide concentrates.

Violarite (FeNi2S4)

[0073] Violarite is a nickel(III) sulfide mineral that is usually associated with primary pentlandite nickel sulfate ores.

Reagents

[0074] One skilled in the art will understand that any compound with a thiocarbonyl functional group can potentially be used in accordance with the technology described herein. The skilled person will also understand that reagents with a thiocarbonyl functional group include, but are not limited to, Tu, ethylene thiourea (ETu), thioacetamide (TA), sodium dimethyldithiocarbamate (SDDC), ethylene trithiocarbonate (ETC), and thiosemicarbazide ( TSCA).

[0075] A non-exhaustive list of additional compounds with a thiocarbonyl functional group is: isothiourea; N-N' substituted thioureas, of which ETu (also known as 2-Thioxoimidazolidine or N,N-Ethylenethiourea) is an example; 2,5-dithiobiurea; dithiobiuret; thiosemicarbazide purum, thiosemicarbazide; methyl chlorothiolformate; dithio-oxamide; thioacetamide; 2-methyl-3-thiosemicarbazide; 4-methyl-3-thiosemicarbazide; purum vinylene trithiocarbonate; vinylene trithiocarbonate; 2-cyanothioacetamide; ethylene trithiocarbonate; potassium ethyl xanthogenate; dimethylthiocarbamoyl chloride; dimethyldithiocarbamate; S,S'-dimethyl dithiocarbonate; dimethyl trithiocarbonate; N,N-dimethylthioformamide; 4,4-dimethyl-3-thiosemicarbazide; 4-ethyl-3-thiosemicarbazide; O-isopropylxantic acid; ethyl thiooxamate; ethyl dithioacetate; pyrazine-2-thiocarboxamide; diethylthiocarbamoyl chloride; diethyldithiocarbamate; tetramethylthiuram monosulfide; tetramethylthiuram disulfide; pentafluorophenyl chlorothionoformate; 4-fluorophenyl chlorothionoformate; O-phenyl chlorothionoformate; O-phenyl chlorothionoformate; phenyl chlorothionoformate; 3,4-difluorothiobenzamide; 2-bromothiobenzamide; 3-bromothiobenzamide; 4-bromothiobenzamide; 4-chlorothiobenzamide; 4-fluorothiobenzamide; thiobenzoic acid; thiobenzamide; 4-phenylthiosemicarbazide; O-(p-tolyl) chlorothionoformate; 4-bromo-2-methylthiobenzamide; 3-methoxythiobenzamide; 4-methoxythiobenzamide; 4-methylbenzenethioamide; thioacetanilide; thiosemicarbazone salicylaldehyde; indole-3-thiocarboxamide; S-(thiobenzoyl)thioglycolic acid; 3-(acetoxy)thiobenzamide; 4-(acetoxy)thiobenzamide; N'-[(e)-(4-chlorophenyl)methylidene] methyl hydrazonothiocarbamate; 3-ethoxythiobenzamide; 4-ethylbenzene-1-thiocarboxamide; tert-butyl 3-[(methylsulfonyl)oxy]-1-azethanecarboxylate; diethyldithiocarbamic acid; 2-(phenylcarbonothioylthio)propanoic acid; 2-hydroxybenzaldehyde N-ethylthiosemicarbazone; (1R,4R)-1,7,7-trimethylbicyclo[2.2.1]heptane-2-thione; tetraethylthiuram disulfide; tetraethylthiuram disulfide; 4'-hydroxybiphenyl-4-thiocarboxamide; 4-biphenylthioamide; dithizone; 4'-methylbiphenyl-4-thiocarboxamide; tetraisopropylthiuram disulfide; anthracene-9-thiocarboxamide; phenanthrene-9-thiocarboxamide; sodium dibenzyldithiocarbamate; and 4,4'-bis(dimethylamino)thiobenzophenone. Such agents are readily available, for example, from Sigma Aldrich.

[0076] Tu, ETu, TA, SDDC, ETC, and TSCA each have a thiocarbonyl functional group with a sulfur that 1) has a partial negative charge, 2) has a negative surface electrostatic potential, and 3) has a π* antibonding orbital empty as its lowest unoccupied molecular orbital (LUMO). Therefore, one skilled in the art can reasonably expect that other reagents, including the additional reagents listed above, that share such criteria and are sufficiently soluble in water, would be useful in performing the methods described herein (provided they do not complex with the metal or iron oxidant to form precipitates). It will be within the skill of the skilled person to identify potentially useful reagents and test them to determine effectiveness with any particular ore, if any.

[0077] For example, Tu has a thiocarbonyl functional group with sulfur having a partial charge of -0.371 as calculated using Gaussian 09 software, a negative electrostatic potential around sulfur, and antibonding orbital π* as its LUMO. Thus, Tu satisfies the three criteria and demonstrated catalytic effect.

[0078] TA has a structure similar to Tu, but with a side chain of CH3 instead of NH2. It has a thiocarbonyl functional group with sulfur with a partial charge of -0.305, as calculated using Gaussian 09 software, which is slightly smaller than that of Tu, a negative electrostatic potential around sulfur, and a π* antibonding orbital as its LUMO. Consequently, TA also satisfies all three criteria and has demonstrated catalytic effect.

[0079] ETC differs from Tu and TA in that it does not contain any thioamide groups. It has a thiocarbonyl functional group with the two sulfur atoms o-attached to the carbon as the side chain. Sulfur in the thiocarbonyl group has a partial charge of -0.122, as calculated using Gaussian 09 software, which is much lower than Tu, a negative electrostatic potential around sulfur, and a π* antibonding orbital as its LUMO. Consequently, ETC also satisfies all three criteria and has demonstrated catalytic effect.

[0080] In comparison, urea has a carbonyl functional group with a C=O bond instead of C=S. The oxygen in the C=O bond has a partial charge of -0.634 as calculated using Gaussian 09 software, and a negative electrostatic potential around it, which is very similar to the sulfur atom in Tu. However, your LUMO does not contain π* antibonding. Consequently, urea is not expected to have a catalytic effect on metal leaching.

[0081] Carbon disulfide (CS2) contains two thiocarbonyl functional groups. Although the sulfur atoms of each functional group contain antibonding π* orbitals as their LUMO, they have a partial positive charge of +0.012 as calculated using Gaussian 09 software. Therefore, CS2 is not predicted to have a catalytic effect.

[0082] Of course, the reagent must also be soluble in water. ETC, for example, is only sparingly soluble in water, which may explain why it appears to be less effective than Tu in leaching copper from chalcopyrite.

[0083] Preferably, the reagent will not complex/precipitate with Fe2+/Fe3+ ions. TSCA, for example, is able to form a red colored complex with Fe3+ in solution, which may explain why it is less effective than Tu in leaching copper from chalcopyrite.

[0084] The reagent must also not complex/precipitate with target metal ions such as Cu+, Cu2+, Cd2+, or Ni2+. Dithiooxamide forms an insoluble complex with copper ions and therefore cannot be used for leaching copper sulfide minerals, while TA complexes with Cd2+ ions to form an insoluble complex and therefore cannot be used. to leach cadmium sulfide minerals such as greenockite.

[0085] Again, the skilled person will appreciate that not all compounds comprising a thiocarbonyl functional group will be useful for increasing the rate of metal extraction from a metal sulfide. Furthermore, the skilled person will appreciate that a reagent which works to increase the rate of metal extraction from one metal sulfide may not be useful for increasing the rate of extraction of one metal from a different metal sulfide. Again, it will be within the skill of the skilled person to identify potentially useful reagents and test them to determine effectiveness with any particular ore, concentrate or other material, if any.

Formamidine Disulfide (FDS)

[0086] Formamidine disulfide (FDS) is generated by the oxidation of Tu. In the presence of an oxidant such as ferric sulfate, Tu will partially oxidize Formamidine Disulfide (FDS) according to the following half-cell reaction:

[0087] The SDF contains no thiocarbonyl functional group, but rather a sulfur-sulfur sigma bond. There is an equilibrium between SDS and Tu in a ferric sulfate solution, so a leaching solution prepared with SDS instead of Tu will provide the Tu needed for catalysis of metal sulfide leaching. That is, one SDF molecule will dissociate into two Tu molecules upon dissolution in the ferric sulfate leach solution. Consequently, a leach solution using Tu as the reagent with the thiocarbonyl functional group can be effectively prepared using Tu or SDS.

[0088] The skilled person will understand that because of this balance, the concentration of Tu (and SDF) can fluctuate over time. Accordingly, the "concentration" or "Tu equivalent" as used herein to refer to the concentration of Tu in the leach solution refers to the amount of Tu present in the solution as if all the SDF in the solution were dissociated into Tu (i.e. that is, ignoring the interconversion between the two forms). Similarly, "concentration", as used herein to refer to the concentration of SDS in the leach solution, refers to the amount of SDS present in the solution as if all of the Tu in the solution were converted to SDS (i.e., ignoring the interconversion between the two forms).

[0089] The “initial concentration” is used herein to refer to the initial concentration of the reagent at the time the leach solution is applied to the ore sample. However, one skilled in the art will understand that the reagent concentration can decrease over time (eg, through precipitation or decay) as the solution percolates through the column or stack. Consequently, the skilled person will recognize that the processes described herein should work to increase the rate of metal extraction from the metal sulphide, provided the concentration of the reactant is within a suitable range during some portion of the percolation through the ore. Accordingly, "contact" material (eg, ore or concentrate, or any other material comprising a base metal sulfide), as used herein, refers to material contact at any point in the leaching process. For greater certainty, “contacting” is not limited to the initial action by which the leachate and/or reagent is applied to the material to be leached, but includes contact between the leachate and/or the reagent at any point during the leaching process.

[0090] In the presence of SDS and ferric sulfate (or other suitable oxidant), the anodic dissolution of a copper sulfide mineral such as chalcopyrite can proceed according to the following two reactions, with oxidation of chalcopyrite by SDS or ferric, respectively :

[0091] After the chalcopyrite is oxidized and the copper is leached from the concentrate, it is desirable to recover the copper from the charged leach solution.

[0092] The methods described here involve two basic steps, namely, leaching and recovery of metals, for example, solvent extraction (SX) and electrolytic extraction (EW), collectively SX-EW. The leaching process can be carried out as a percolation leaching (as a heap leaching), a leaching in or a leaching in a tank as known in the field.

[0093] For purposes of this description, the words “containing” and “comprising” are used in a non-limiting sense to mean that items following the word are included, but items not specifically mentioned are not excluded. A reference to an element by the indefinite article "a" does not exclude the possibility that more than one of the elements is present, unless the context clearly requires that there be one and only one of those elements.

[0094] A "percolation leaching", as used herein, refers to the selective removal of a mineral by causing a suitable solvent to seep into and through a mass or pile of material containing the desired soluble mineral, for example , a column leaching or a heap leaching.

[0095] A “column leaching”, as used herein, refers to leaching through the use of a long, narrow column in which the ore sample and solution are in contact to measure the effects of typical variables encountered in real heap leaching.

[0096] A “pile leaching”, as used herein, is a process by which metals are extracted from the ore in which they are found, ie without beneficiation. A heap leach is often chosen for its efficiency and cost-effectiveness. After being removed from the ground, the ore is normally sent through a crusher to break the ore into smaller particles (although heap ores can be “raw” where the ore is leached into a “popped-like” state without further grinding. ). Pile ores can be the primary, secondary or tertiary crushing product. Traditionally, the crushed particles are then “stacked”, or “heaped up” into a large pile.

[0097] A persistent cause of failure in heap leaching operations is the presence of excess fines in the materials placed in the block. Excess fines result in a low permeability material and therefore the leachate infiltration rate is too slow, or contact with the ore solution is insufficient, for economical block operations. Consequently, the efficiency of a heap leach can be increased by agglomeration after grinding. "Agglomeration", as used herein, refers to a technique that joins together fine materials or particles to create a larger product. Agglomeration can be achieved by different methods known in the art. Typically, heap leach agglomeration is performed in a drum agglomerator with sulfuric acid and no binder, or on conveyor belts with acid sprayed onto the ore at drop points.

[0098] The pile is irrigated with a solution that depends on the type of ore being mined. The acid for the leaching will preferably be generated by bacteria using processes known in the art. Alternatively, additional acid can be added as needed.

[0099] The irrigated solution is allowed to percolate through the ore and drain to the bottom of the pile. The ore pile sits on top of an impermeable layer, such as a plastic sheet, which collects the charged leach solution as it drains and directs it to a collection pond. Once the solution is collected, it is pumped to a recovery plant to extract the copper by solvent extraction and electrolytic extraction (SX-EW).

[00100] Applying the methods described herein to a heap leach, ore containing an appropriate sulfide mineral is selectively leached in the presence of the acid sulfate and the reagent having a thiocarbonyl functional group. The concentration of the reagent with a thiocarbonyl functional group in the leach solution can be around 30 mM or perhaps even higher. The skilled person will understand that it is only necessary that the reagent concentration be within a range sufficient to increase the metal sulfide leaching rate.

[00101] Furthermore, although reagent concentrations of about 100 mM or less are low enough to facilitate metal leaching from a particular metal sulfide, concentrations of 100 mM may not be economically feasible at the present time. Consequently, it may be preferable to use lower reagent concentrations that are economically and operationally feasible, e.g., about 90 mM or less, about 80 mM or less, about 70 mM or less, about 60 mM or less, about 50 mM or less, about 40 mM or less, about 30 mM or less, about 20 mM or less, about 10 mM or less, about 5 mM or less, about 4 mM or less, about 3 mM or less, about 2 mM or less, about 1.5 mM or less, about 1 mM or less, about 0.9 mM or less, about 0.8 mM or less, about 0.8 mM or less less, about 0.7 mM or less, about 0.6 mM or less, about 0.5 mM or less, about 0.4 mM or less, 0.3 mM or less, or about 0, 2 mM.

[00102] Consequently, the concentration of the reagent in the acid sulfate solution can be in the range of about 0.2 mM to about 0.3 mM, about 0.2 mM to about 0.4 mM, about 0 .2 mM to about 0.5 mM, about 0.2 mM to about 0.6 mM, about 0.2 mM to about 0.7 mM, about 0.2 mM to about 0.0, 8 mM, about 0.2 mM to about 0.9 mM, about 0.2 mM to about 1.0 mM, about 0.2 to about 1.5 mM, about 0.2 to about 2.0 mM, about 0.2 to about 2.5 mM, about 0.2 to about 3 mM, about 0.2 to about 4 mM, about 0.2 to about 5 mM, about 0.2 to about 10 mM, about 0.2 to about 20 mM, about 0.2 to about 30 mM, about 0.2 to about 40 mM, about 0 .2 to about 50 mM, about 0.2 to about 60 mM, about 0.2 to about 70 mM, about 0.2 to about 80 mM, about 0.2 to about 90 mM mM, or about 0.2 to 100 mM.

[00103] The leaching process can be performed at temperatures between 0°C (that is, the freezing point of water) and 80°C. However, the process would normally be carried out at room temperature and atmospheric pressure.

[00104] In some situations, it may be necessary or preferable to perform the leaching with a leachate comprising a halide. A halide can include chloride, bromide or iodide. For example, it may be necessary to perform leaching with brackish water, seawater or brine. Consequently, the leaching process described herein can be carried out with a leaching solution comprising chloride at a concentration of up to 120 g/L. The chloride concentration in the acid sulfate solution can range from about 1 g/L to about 10 g/L, about 1 g/L to about 20 g/L, about 1 g/L to about from 30 g/L about 1 g/L to about 40 g/L, about 1 g/L to about 50 g/L, about 1 g/L to about 60 g/L, about 1 g/L to about 700 g/L, about 1 g/L to about 80 g/L, about 1 g/L to about 120 g/L, about 1 g/L to about 90 g /L, about 1 g/L to about 100 g/L, about 1 g/L to about 110 g/L, or about 1 g/L to about 120 g/L. In specific embodiments, the concentration of chloride in the acid sulfate solution is in the range of about 20 g/L to about 120 g/L, 20 g/L to about 80 g/L, or 20 g/L to about of 50 g/L.

[00105] Alternatively, the leaching process described herein can be carried out with a leaching solution comprising bromide at a concentration of up to 30 g/L. The concentration of bromide in the acid sulfate solution can range from about 1 g/L to about 10 g/L, about 1 g/L to about 20 g/L, or about 1 g/L to about 30g/L. In specific embodiments, the concentration of chloride in the acid sulfate solution is in the range of about 10 g/L to about 30 g/L.

[00106] Alternatively, the leaching process described herein can be carried out with a leaching solution comprising iodide at a concentration of up to 300 ppm. The chloride concentration in the acid sulfate solution can range from about 1 g/L to about 10 ppm, about 1 ppm to about 20 ppm, about 1 ppm to about 30 ppm, about 1 ppm about 40 ppm, about 1 ppm to about 50 ppm, about 1 ppm to about 60 ppm, about 1 ppm to about 70 ppm, about 1 ppm to about 80 ppm, about 1 ppm about 90 ppm, about 1 ppm to about 100 ppm, about 1 ppm to about 110 ppm, about 1 ppm to about 120 ppm, about 1 ppm to about 130 ppm, about 1 ppm about 140 ppm, about 1 ppm to about 150 ppm, about 1 ppm to about 160 ppm, about 1 ppm to about 170 ppm, about 1 ppm to about 180 ppm, about 1 ppm about 190 ppm, about 1 ppm to about 200 ppm, about 1 ppm to about 210 ppm, about 1 ppm to about 220 ppm, about 1 ppm to about 230 ppm, about 1 ppm about 240 ppm, about 1 ppm to about 250 ppm, about 1 ppm to about 260 ppm, about 1 ppm to about 270 ppm, about 1 ppm to about 280 ppm, about 1 ppm to about 290 ppm, or about 1 ppm to about 300 ppm. In specific embodiments, the chloride concentration in the acid sulfate solution is in the range of about 100 ppm to about 300 ppm.

Solvent extraction

[00107] After the leaching process, copper can be extracted from the leaching solution. After a solid-liquid separation, i.e., draining the charged leach solution containing copper from the heap, the charged solution is preferably subjected to conventional solvent extraction and electrolytic extraction to produce pure copper cathodes according to the following general reaction :

[00108] Reagents with a thiocarbonyl functional group in the charged leach solution should not present any problems in the electrolytic extraction operation and, in fact, may even be useful as a leveling agent. The Tu-containing raffinate can then be recirculated to the stockpile for further leaching. The recirculated leach solution can also be supplemented with Tu to arrive at the desired initial Tu concentration for the leach.

[00109] The PLS recovered from heap leaching will contain iron and copper ions. It is known that reagents comprising functional thiocarbonyl can form various stable complexes with copper ions ( Doona and Stanbury, Inorg Chem 35:3210-3216 ; Mironov and Tsvelodub, J Solution Chem 25:315-325 ; Bowmaker et al., Inorg Chem 48:350-368). Extractors commonly used for solvent extraction of copper (SX), such as hydroxymes and aldoximes, are strong complexing agents for copper ions. Solvent extractors can alter the balance between copper ions and thiocarbonyl ligands, releasing thiocarbonyl ligands from copper complexes. As free thiocarbonyl ligands enter the raffinate solution, they can be returned to the pile and continue to catalyze leaching.

[00110] Accordingly, the PLS recovered from leaching through solid-liquid separation is then mixed with an organic solvent containing a base metal ion extractor to form a mixture. The skilled person will be able to select an appropriate solvent depending on the metal ion to be extracted. The organic solvent can be an aliphatic solvent, an aromatic solvent or a combination thereof. The organic solvent can include kerosene, aromatic alkyls, cycloparaffins or a combination thereof.

[00111] The knowledgeable person will also be able to select an appropriate extractor. The base metal ion extractor can be an aldoxime, a ketoxime or a combination thereof. The base metal ion extractor can additionally include an ester modifier, an alkylphenol modifier, or a combination thereof.

[00112] During solvent extraction, the base metal cations are decomplexed from the reagent, releasing the reagent and allowing the base metal cations to be extracted from the PLS into the organic solvent. The free reagent remains in the aqueous phase. Separation of the organic solvent from the aqueous phase results in a base metal ion depleted raffinate comprising the free reagent and a base metal ion enriched organic phase comprising the organic solvent and base metal ions.

[00113] The base metal ion enriched solution can then be processed to recover the base metal. Raffinate, on the other hand, can be recirculated for use in the leachate.

[00114] Retention of the free reagent in the aqueous phase during solvent extraction to produce the raffinate comprising the free reagent can be achieved with halides, for example, chloride, bromide or iodide, present in PLS in concentrations discussed above.

[00115] As discussed above, the skilled person will understand that there is a balance between Tu and SDF, such that the ratio of SDF and TU-Cu complexes to Tu in PLS is greater than in the leachate. Since Tu has a stronger effect on enhancing the leaching of base metals from sulphide ores/concentrates than SDS or TU-Cu complex, increase the proportion of free Tu in the raffinate prior to recirculation for leaching, for example, by decomplexing Tu from the base metal ions in the PLS or adding a reducing agent to bias the balance in favor of Tu, you can enhance the leaching process.

[00116] With reference to Figure 5, a method for recovering a base metal from a base metal sulfide is shown at 500. The method begins by contacting the material comprising at least one base metal sulfide, e.g. ore or concentrated, with a bleach. The leachate comprises an acidic sulfate solution and a reagent having a thiocarbonyl functional group as described above for extracting base metal ions from at least one base metal sulfide to produce a charged leach solution (PLS) comprising reagent and base metal. A portion of the reagent is complexed with base metal ions. Leaching can take place in a reactor (i.e., a reaction vessel), or in a heap that does not involve a reactor.

[00117] With reference to Figure 6, in particular embodiments in which the reagent is Tu, the raffinate is mixed with a reducing agent before returning the raffinate to leaching in order to influence the balance between SDF and Tu from SDS to Tu. The skilled person will be able to select an appropriate reducing agent. For example, the reducing agent can be H2S, NaSH, or Zinc. The reducing agent can be added to obtain a Tu:SDF ratio in the range of about 0.5:1 to about 9:1.

Examples

[00118] To facilitate the extraction of metal ions from the minerals listed above, reagents with a thiocarbonyl functional group were added to ferric acid sulfate solutions as catalysts. In the experiments described here, it was found that reagents containing thiocarbonyl functional groups have a positive catalytic effect on the extraction of minerals. Among all reagents, Tu consistently provides the highest catalytic performance. Consequently, Tu was the most studied reagent of those identified. However, results from experiments with other reagents with thiocarbonyl functional groups are provided to compare their catalytic effects. SDF, which does not contain a thiocarbonyl functional group but has comparable catalytic effect as Tu, was studied as a special case due to its equilibrium with Tu. Leach reactions were carried out at atmospheric pressure in a variety of ore compositions, reagent concentrations, ferric concentrations and under various other conditions as described below.

Example 1 Extraction of copper from chalcopyrite using thiourea Example 1.1

[00119] The effect of Tu on the electrochemical behavior of a chalcopyrite electrode was studied in a conventional 3-electrode glass-lined cell. A CuFeS2 electrode was used as a working electrode, a saturated calomel electrode (SCE) was used as a reference, and a graphite bar was used as a counter electrode. The CuFeS2 electrode was polished using 600 and 1200 grit carbide paper. All experiments were performed at 25°C using a temperature-controlled water bath. The electrolyte composition was 500 mM H2SO4, 20 mM Fe2SO4 and 0 to 100 mM Tu. Before starting any measurements, the solutions were bubbled with N2 for 30 minutes to reduce the dissolved O2 concentration. The open circuit potential (OCP) was recorded until changes of no more than 0.1 mV/min were observed. After observing a constant OCP value, electrochemical impedance spectroscopy (EIS) was performed on the OCP using a sinusoidal perturbation 5 mV a.c. from 10 kHz to 10 mHz. Linear polarization resistance (LPR) tests were also performed using a sweep rate of 0.05 mV/s at ± 15 mV OCP.

[00120] Linear potential scans were conducted at electrode potential ± 15 mV from the measured OCP at each Tu concentration. All scans showed a linear behavior within the potential range of the analyzed electrode. An increase in the slope of the experimental graphs was observed with increasing Tu concentration. The slope of these curves was used to estimate the polarization resistance value (Rct) at each concentration. These values were then used to estimate the dissolution current density values using equation 1:

[00121] Figure 3 shows the effect of Tu on the dissolution current density and the mixed potential of the CuFeS2 electrode, and indicates that a maximum dissolution current density was reached when the Tu concentration is 30 mM. Increasing the Tu concentration to 100 mM resulted in a decrease in the current density and mixed potential of the CuFeS2 electrode. Furthermore, after immersion of the CuFeS2 electrode in the 100 mM Tu solution, a copper-like film was observed on the electrode surface, which film could only be removed by polishing the electrode with carbide paper.

Example 1.2

[00122] Figure 4 is a bar graph showing the effect of the initial concentration of Tu or SDS on the electrochemical dissolution of a chalcopyrite electrode in the sulfuric acid solution at pH 2 and 25°C. A 10 mM Tu concentration in the leach solution resulted in a six-fold increase in dissolution rate compared to no Tu, and a 5 mM SDF concentration resulted in a six-fold increase over 10 mM Tu . A concentration of 10 mM Tu in leach solution also containing 40 mM Fe(III) resulted in a thirty-fold increase in dissolution rate compared to 40 mM Fe(III) alone.

Example 1.3

[00123] A column leaching of different acid-cured copper ores was conducted with Tu added to the leaching solution. A schematic description of the column configuration is shown in Figure 5. The column diameter was 8.84 cm, the column height was 21.6 cm, and the column stack height was 15.9 cm. The irrigation rate was 0.77 mL/min or 8 L/m2/h. The charged leach solution emitted from these columns was sampled for copper every 2 or 3 days using Atomic Absorption Spectroscopy (AAS).

[00124] The specific mineralogical composition of these ores is provided in Table 1. The Cu contents of Ore A, Ore B and Ore C were 0.52%, 1.03% and 1.22% w/w, respectively. Prior to leaching, the ore was “acid cured” to neutralize the acid-consuming material present in the ore. That is, the ore was mixed with a concentrated sulfuric acid solution composed of 80% concentrated sulfuric acid and 20% deionized water and left to stand for 72 hours. For a treatment using Ore C, Tu was added to the sulfuric acid curing solutions.

[00125] The initial composition of the leach solutions included 2.2 g/L of Fe (that is, 40 mM, provided as ferric sulfate) and pH 2 for the control experiment, with or without 0.76 g/L of Thou (i.e. 10 mM). The initial mineral charge in each column was 1.6 to 1.8 kg of ore. The surface velocity of the solution through the ore column was 7.4 L m-2 h-1. The pH was adjusted using dilute sulfuric acid. These two columns were maintained in an open-loop or open-loop configuration (i.e., no recycled solution) for the entire leaching period.

[00126] The results of the leaching tests on Ore A, Ore B and Ore C are shown in Figures 6, 7 and 8, respectively. The presence of Tu in the leachate clearly has a positive effect on the leaching of copper from chalcopyrite. On average, the leaching rate in the presence of Tu was increased by a factor of 1.5 to 2.4 compared to control tests where the leaching solutions did not contain Tu. From the latest time points depicted in Figures 6 to 8, the copper extractions for columns containing Ore A, Ore B and Ore C leached with a solution containing sulfuric acid and ferric sulfate alone, without the addition of Tu, were 21.2 % (after 198 days), 12.4% (after 50 days) and 40.6% (after 322 days), respectively. With 10 mM Tu added, these extractions were 37.9%, 32.0% and 72.3%, respectively.

[00127] Referring to Figure 8, 2 mM Tu was added to the leaching solution originally not containing Tu from day 322 onwards, after which the leaching rate increased markedly. From day 332 to day 448, copper leached from this column increased from 40% to 58%, and rapid leaching was maintained throughout this period.

[00128] The averages for the last 7 days reported in Figure 9 indicate that the leaching rate of Ore C cured with acid leach in the presence of 10 mM Tu is 3.3 times greater than for Ore C cured with acid leach in the absence of thiourea and 4.0 times higher than acid-cured and Tu-cured Ore C leached in the absence of Tu.

[00129] Figure 10 shows the effect of Tu on solution potential. All potentials are reported against an Ag/AgCl (saturated) reference electrode. The solution potential of the leach solutions containing Tu was generally between 75 and 100 mV lower than the solution potential of the leach solution that did not include Tu. Lower solution potentials are consistent with Tu working to prevent chalcopyrite passivation.

Example 1.4 Spin bottle leaching

[00130] “Spinner bottle” leaching experiments in the presence of various concentrations of Tu were performed for coarse Ore A and Ore B. Tests were performed using coarsely ground ore (100% ^inch pass).

[00131] Before leaching, the ore was cured using a procedure similar to that performed on the ore used in the column leaching experiments. The ore was mixed with a concentrated sulfuric acid solution composed of 80% concentrated sulfuric acid and 20% deionized water and allowed to settle for 72 hours to neutralize acid-consuming material present in the ore. For various experiments, different concentrations of Tu were added to the ore using the sulfuric acid curing solutions.

[00132] The bottles used for the experiments were 20 cm long and 12.5 cm in diameter. Each bottle was filled with 180g of cured ore and 420g of leach solution, filling to about one-third of the volume of the bottle.

[00133] The leach solution from each vial was sampled at 2, 4, 6, and 8 hours, and then every 24 hours thereafter. Samples were analyzed using atomic absorption spectroscopy (AAS) for their copper content.

[00134] The conditions for the rotating bottle experiments are listed in Table 2. Experiments #1 to #6 were performed using only the original addition of Tu to the bottles. For experiments #7 to #11, Tu was added every 24 hours to restore the Tu concentration.

[00135] A positive effect of Tu on copper leaching was observed. For the coarse ore experiments, a plateau was not observed until after 80 to 120 hours. Tu was periodically added to coarse ore experiments, producing positive results in copper dissolution.

[00136] The effect of different concentrations of Tu in the leach solution in the coarse ore leaching (experiments No. 1 to 11 as described in Table 2) is shown in Figures 11 and 10.

[00137] For ore B, Tu was periodically added every 24 hours to restore the thiourea concentration in the system and thus better emulate the conditions in the column leaching experiments. As can be seen from Figure 9, 8 mM and 10 mM Tu yielded higher copper dissolution results than the other Tu concentrations tested for ore A. A plateau in dissolution is not observed until approximately 120 hours, the which varied with Tu concentration as shown in Figure 11. Table 1.

[00138] As can be seen from Figure 12, 5 mM Tu yielded higher copper dissolution results than the other Tu concentrations tested for ore B. As with ore A, a plateau in dissolution is not observed up to approximately 80 to 120 hours, which varied with Tu concentration as shown in Figure 12. Periodic addition of Tu resulted in increased copper dissolutions and produced a delay in the dissolution plateau.

[00139] Interestingly, solutions containing 100 mM Tu did not appear to be much more effective in extracting copper than those containing no Tu, and even worse at times. This is consistent with the results of Deschênes and Ghali, who reported that solutions containing ~200 mM Tu (ie 15 g/L) did not improve copper extraction from chalcopyrite. Tu is less stable at high concentrations and breaks down. Consequently, it is possible that when initial Tu concentrations are slightly greater than 30 mM, sufficient elemental sulfur may be produced by decomposition of Tu to form a film on the chalcopyrite mineral and thereby aid in its passivation. It is also possible that, at high Tu dosages, some copper precipitates from the solution (eg see Figure 17) are responsible for some of the poor extraction results.

Example 2 Extraction of Chalcopyrite, Covellite, Chalcocite, Bornite, Enargite, Pentlandite, Violarite, and Greenockite using Thiourea

[00140] The catalytic effect of Tu was further demonstrated in stirred reactor tests. All reactors contained 1.9 L of ferric sulfate solution at pH 1.8 and total iron concentration of 40 mM. 1 g of mineral samples was used in each reactor test. These experimental conditions were designed to maintain an unlimited supply of oxidant.

[00141] To demonstrate the catalytic effect of chalcopyrite, a 100% pure synthetic chalcopyrite was used instead of chalcopyrite concentrate which contains various impurities. Chalcopyrite was synthesized through a hydrothermal approach. CuCl, FeCl3 and Tu were first mixed at a 1:1 molar ratio and dissolved in 150 mL of DI water. The solution was transferred to a Teflon-lined reaction vessel and heated to 240°C for 24 hours. At the end of the reaction, the precipitated powder was washed with acidic water (pH = 1) and dried in air at room temperature. XRD analysis showed that the synthetic chalcopyrite was free of any impurities compared to the chalcopyrite mineral concentrate. This synthetic chalcopyrite was used in all tests performed in stirred reactors as described herein. Table 2. List of rotating bottle leaching experiments involving Ore A and Ore B.

[00142] The covellite mineral used in the experiment described here was also synthesized through a hydrothermal approach. CuCl and Tu were mixed at a 1:1 molar ratio and dissolved in 150 mL of DI water. The solution was transferred to a Teflon-lined reaction vessel and heated to 220°C for 24 hours. The synthesized CuS was washed with acid and dried in air. XRD analysis showed it to be 100% pure with no interference from other species.

[00143] The sample of chalcocite mineral used in the experiments described here was 100% pure natural mineral.

[00144] The mineral bornite used in the experiments described here was obtained from Butte, Montana with a copper content of 58.9% based on ICP-AES. XRD analysis showed that the mineral contains 76.8% bornite, 8.1% chalcopyrite, 6.3% pyrite, 5.8% tennantite and 3.0% enargite. The calculated copper content from XRD was 55.6%, which is relatively consistent with the chemical assay.

[00145] The mineral enargite used in the experiments described here was in the form of an enargite concentrate, which contained approximately 70% enargite (34% copper) according to XRD analysis.

[00146] The mineral greenockite used in this experiment was synthesized through a hydrothermal approach. CdCl2 and Tu were mixed in a 1:1 molar ratio and dissolved in 100 mL of DI water. The solution was transferred to a Teflon-lined reaction vessel and heated to 150°C for 24 hours. The synthesized CdS was washed with acid and dried in air. XRD analysis showed it to be 100% pure with no interference from other species.

[00147] The violarite used in the experiments described here was the natural violarite mineral containing 15.8% Ni according to ICP-AES. XRD analysis showed that the mineral was approximately 42% violinite and 13.1% NiSO4.6H2O.

[00148] The sulfur in thiocarbonyl groups contains a lone pair of electrons and a filled π orbital that can be used for donor-acceptor bonding with a transition metal, along with an antibonding π* orbital that could potentially accept the further donation of electrons from full d orbitals in the transition metal. Consequently, without wishing to be bound by theory, it is suspected that the interaction between the surface ion and the thiocarbonyl functional group, especially the reverse metal donation to the ligand, is responsible for the catalytic effect. Furthermore, it is suspected that the catalytic effect should be more pronounced for transition metals with a higher number of d electrons, with the catalytic effect being more pronounced for minerals with d10 electron configuration.

[00149] Figure 13 shows that Tu catalyzes the leaching of common copper sulfide minerals, including chalcopyrite, covellite, chalcocite, and bornite, which contain Cu(I). After 96 hours of leaching, chalcopyrite extraction reaches 64.1% with 2 mM Tu compared to 21.1% without Tu; covellite extraction reaches 74.4% with 2 mM Tu compared to 7.2% without Tu; chalcocite extraction reaches 85.6% with 2 mM Tu compared to 65.1% without Tu; bornite extraction reaches 91.4% with 2 mM Tu compared to 56.7% without Tu.

[00150] Like Cu(I), Cd(II) also contains the electronic configuration d10. Figure 14 shows that CdS mineral leaching is significantly increased with the addition of Tu. With Tu, the cadmium extraction reached 100% at 48 hours, while the extraction in the uncatalyzed reaction stabilized at 47% after 96 hours.

[00151] The copper ion in the mineral enargite has fewer d electrons than other primary and secondary sulfides, and therefore the catalytic effect can be expected to be slower than that of Cu(I) minerals. However, the results shown in Figure 15 clearly demonstrate that leaching with a leaching solution comprising an initial concentration of 2 mM Tu increases the rate of copper leaching from enargite compared to a control without Tu, which did not show significant extraction. after 96 hours of leaching.

[00152] Minerals containing transition metal ions with electronic configuration d7, such as Ni(III), can also undergo catalyzed leaching with the addition of Tu. Similar to Cu(II), as Ni(III) is the highest stable oxidation state with 7 d electrons, the catalytic effect is not expected to be as drastic as for d10 minerals. Referring to Figure 16, leaching with a leaching solution comprising an initial concentration of 2 mM Tu increases the nickel leaching rate of violarite compared to a control without Tu.

[00153] The results of the leaching experiments referred to in Example 2 are summarized in Table 3, in which the extraction percentages under uncatalyzed and catalyzed conditions (with an initial Tu concentration of 2 mM) are compared. Table 3. Comparisons of reactor leaching for various minerals under uncatalyzed and 2 mM Tu catalyzed conditions

Example 3 Reagent dosing

[00154] The ideal dosage of reagent can increase the efficiency of leaching. First, at certain concentrations, the reagent can form an insoluble complex with the metal ion of interest and precipitate. For example, Tu can form an insoluble complex with Cu(I) ions with a molar ratio of 3:1. A precipitation test was performed to examine the concentration range in which Cu-Tu complex precipitation can occur. 20 mL of Cu solution was divided into several identical portions followed by the addition of various doses of Tu (ie, 0 to 60 mM). The solution was stirred for 24 hours and the Cu remaining in the solution phase was analyzed by AAS. The results are shown in Figure 17, plotted as the percentage of Cu remaining.

[00155] Second, heap leaching of metal sulfides is based on a bioleaching mechanism, an excessive amount of reagent can be harmful to bioleaching microbes. For example, bacteria commonly used for bioleaching, such as Acidithiobacillus ferrooxidans and Acidithiobacillus thiooxidans, grow very slowly in a solution containing 10 mM Tu and cannot survive 100 mM Tu.

[00156] Third, with regard to Tu specifically, ferric reacts with Tu and converts it to SDF (see Hydrometallurgy 28, 381-397 (1992)). Although the reaction is reversible under certain conditions, a high concentration of SDS tends to irreversibly decompose to cyanamide and elemental sulfur (see J Chromatogr 368, 444-449).

[00157] Therefore, the excess addition of Tu in the leachate can cause the loss of Fe3+ and Tu due to oxidation and decomposition. Irreversible decomposition of SDS was observed when adding 4 mM Tu in a 40 mM ferric sulfate solution at pH 1.8.

[00158] To further investigate the effect of Tu dosage on copper extraction, stirred reactor tests were performed using 1 g of synthetic chalcopyrite in 1.9 L of 40 mM ferric sulfate solution at pH 1.8 with various initial Tu concentrations. Treatments were administered for 172 hours to approximate maximum extraction. The results are presented in Figure 18, and show that, for 1 g of chalcopyrite, the dosage of higher Tu content results in faster leaching kinetics between the tested Tu concentrations.

[00159] For Tu dosages of 5 mM and below, the initial solution of 40 mM of ferric sulfate can be considered as a sufficient supply of oxidant. However, for higher dosages, such as 10 mM and 20 mM Tu, extra ferric (in a 1:1 ratio with Tu) had to be added to the solution to allow oxidation of Tu to SDS. For 10 mM Tu, an extra 10 mM Fe3+ was added at time zero. For 20 mM Tu, an extra 20 mM Fe3+ was added at 72 hours, which led to the continuation of the extraction as shown in Figure 18.

[00160] The dosage of Tu versus Cu extraction at 172 hours is plotted in Figure 19. An initial dosage of Tu up to 5 mM appears to have the most pronounced effect on Cu dissolution.

[00161] As indicated above, in previous flask shaking tests with acidic solutions (pH 1.8) containing various concentrations of Fe3+ and Cu2+ ions, slight precipitation occurred after the addition of 4 mM Tu due to SDF decomposition. Consequently, concentrations of Tu concentration below 4 mM can prevent such precipitation. A series of flask shake tests were performed on solutions containing initial concentrations of 2 mM Tu and various concentrations in a matrix containing Fe3+ (0 to 100 mM) and Cu2+ (0 to 50 mM) in order to identify concentration ranges of [Fe3+] and [Cu2+] that do not result in precipitation of Cu complexes. The results showed that no precipitation and no loss of Cu from the solution phase resulted from the use of 2 mM Tu in this wide range of Fe and Cu matrix concentrations.

Example 4 Alternative reagents

[00162] The catalytic effect of several other reagents with a thiocarbonyl functional group has been examined in the leaching of synthetic chalcopyrite, covellite, bornite and enargite. The experiments were carried out in stirred reactors containing 40 mM ferric sulfate solution at pH 1.8. 1 g of chalcopyrite or covellite was added to the reactors along with an initial 2 mM concentration of various thiocarbonyl reagents including Tu, TA, SDDC, ETC and TSCA. Cu extraction curves for chalcopyrite, covellite, bornite and enargite using all or a subset of the above reagents are shown in Figures 20, 21, 22 and 23.

[00163] From Figures 20 to 23, it is clear that each of these additional reagents that have a thiocarbonyl functional group shows a beneficial effect on the leaching of ferric sulfate from each of chalcopyrite, covellite, bornite and enargite.

[00164] Figure 24 summarizes the results of other stirred reactor tests in chalcopyrite that additionally investigate urea and carbon disulfide. These results confirm that, as expected, neither urea nor carbon disulfide are effective reagents.

Example 5 SDS

[00165] The catalytic effect of leaching solutions prepared with SDS on the leaching of chalcopyrite, bornite, covellite and chalcocite was determined in stirred reactor tests. All reactors contained 1.9 L of ferric sulfate solution at pH 1.8 and total iron concentration of 40 mM. 1 g of mineral samples was used in each reactor test. An initial SDF concentration of 1 mM or an initial Tu concentration of 2 mM Tu was used.

[00166] The results of stirred reactor tests shown in Figures 25a and 25b demonstrate that SDF has comparable efficiency to Tu in leaching each of chalcopyrite, bornite, covellite and chalcocite after 96 hours.

Example 6 Graded closed-loop bioleaching with Tu

[00167] A closed-loop bioleaching with Tu was conducted. 7 kg of ore containing approximately 0.25% Cu content, mainly in the form of CuFeS2, was leached at a flow rate of 1 L/day at an aeration rate of approximately 300 mL/min.

[00168] The ore was pretreated with sulfuric acid to leach oxides (eg chalcanthite and basic copper salts) using sulfuric acid. After the acid leaching period, the residual solutions were collected and replaced by a ferrous sulfate solution with nutrients (40mM of FeSO4, 0.4 g/L of magnesium sulfate heptahydrate and 0.04 g/L of di -potassium hydrogen phosphate, with pH adjusted to 1.6 to 1.8). The ferrous and nutrient solution was blasted through the column to establish a good habitat for bacterial growth. Inoculation of bacteria showed an increase in ORP from 274 mV to 550 mV within 48 hours. The solution used in this step and in future steps was kept circulating through the column, forming a self-sustaining closed-loop system.

[00169] At this stage, the remaining copper source is mainly CuFeS2. After the survival of the bacteria on the column, Tu was progressively added to the leach solution. As discussed above, Tu is converted to SDS at a 2:1 molar ratio in the presence of 40 mM Fe3+. Operating potential (ORP) was used as an indicator of bacterial activity and HPLC was used to monitor SDS content. From day 0 to day 50, the leach solution included 40 mM Fe3+ with inoculated bacteria (no Tu added). From day 90 to day 98, a total of 1.878 g of Tu was progressively added, over which HPLC analysis of the effluent showed that the SDF was being maintained at approximately 1.5 mM and no more Tu was added.

[00170] As shown in Figure 26, the ORP of the effluent was always equal to or greater than the influent, indicating that the bacteria were actively oxidizing Fe2+ to Fe3+. SDS contents were analyzed by HPLC, showing approximately 1.5 mM SDS (equivalent to 3 mM added Tu) existing in the solution phase with no precipitation observed. Therefore, it appears that 1.5 mM SDF (3 mM Tu equivalent) can be used in the solution without ferric precipitation.

[00171] The results of the closed-loop leaching test are shown in Figure 27. From day 0 to day 50, the bacteria were able to maintain high activity and oxidize Fe2+ to Fe3+. However, with constant flow (1 L/day), the leaching rate was only 1.97 mg Cu/day during the first 50 days. Addition of Tu from day 90 increased the Cu extraction rate to 6.54 mg/day, which remained constant after day 98. This indicates that the reagent did not undergo decomposition and remained effective in the closed-loop system.

Example 7 Extraction of chalcopyrite in the presence of chloride using a reagent with a thiocarbonyl functional group. Example 7.1

[00172] The effect of chloride on the ability of Tu to facilitate the leaching of a copper sulfide was tested in stirred reactors. Each reactor contained 1g of 100% pure synthetic chalcopyrite in 2L of ferric sulfate solution at pH 1.7, with a total ferric concentration of 40 mM. The experimental reactors included Tu at an initial concentration of 2 mM and a chloride concentration of 20 g/L, 50 g/L, 80 g/L or 120 g/L. Reagents comprising no Tu, no chloride, and no Tu or chloride were included as controls. An additional reactor containing 2 mM Tu and 80 g/L chloride with 200 ppm Cu was also included. These experimental conditions were designed to maintain an unlimited supply of oxidant.

[00173] As shown in Figure 28, the presence of Tu has a positive effect on the extraction of copper from chalcopyrite in the presence of chloride at a concentration of 120 g/L. Although the amount of copper extracted decreased with increasing chloride concentration, copper extraction was nevertheless higher in the presence of Tu compared to the absence of Tu. For example, copper extraction was higher in solutions comprising Tu and 120 g/L of chloride than solutions without Tu and only 20 g/L.

Example 7.2

[00174] The effect of chloride on the ability of Tu or ETu to facilitate the leaching of a copper sulfide was tested in stirred reactors. Each reactor contained 1g of chalcopyrite concentrate which has 21.6% copper per liter of ferric sulfate solution at pH 1.7, with a total ferric concentration of 40 mM. The experimental reactors included Tu or ETu at an initial concentration of 0 or 2 mM and a chloride concentration of 0 g/L, 20 g/L, 80 g/L or 200 g/L. The composition of the solution is listed in Table 4.

[00175] Table 4. Composition of the solution for testing the compatibility of reagents with a thiocarbonyl functional group with chloride

[00176] As shown in Figures 29a and 29b, the presence of Tu or ETu has a positive effect on the extraction of copper from chalcopyrite in the presence of chloride at a concentration of 200 g/L. Although the amount of copper extracted decreased with increasing chloride concentration, copper extraction was nevertheless higher in the presence of Tu compared to the absence of Tu. For example, copper extraction was higher in solutions comprising Tu and 120 g/L of chloride than solutions without Tu and only 20 g/L of chloride.

Example 8 Extraction of chalcopyrite in the presence of bromide using reagents with a thiocarbonyl functional group.

[00177] The effect of bromide on the ability of a reagent with a thiocarbonyl functional group to facilitate the leaching of a copper sulfide was tested in stirred reactors for 180 h. Each reactor contained 1g of chalcopyrite concentrate which has 21.6% copper per liter of ferric sulfate solution at pH 1.7, with a total ferric concentration of 40 mM. The experimental reactors included Tu or ETu at an initial concentration of 2 mM and a bromide concentration of 10 g/L or 30 g/L (supplied as potassium bromide). Reagents comprising neither Tu nor ETu were included as controls. The reactors were stirred at room temperature. Solution compositions are listed in Table 5. Table 5. Solution Composition for Testing Compatibility of Reagents with a Thiocarbonyl Functional Group with Bromide

[00178] As shown in Figures 30a and 30b, Tu and ETu had a positive effect on copper extraction from chalcopyrite in the presence of bromide at an initial concentration as high as 30 g/L.

Example 9 Extraction of chalcopyrite in the presence of iodide using reagents with a thiocarbonyl functional group.

[00179] The ability of a reagent with a thiocarbonyl functional group to facilitate the leaching of a copper sulfide in the presence of iodide was tested in stirred reactors for 180 h. Each reactor contained 1g of chalcopyrite concentrate which has 21.6% copper per liter of ferric sulfate solution at pH 1.7, with a total ferric concentration of 40 mM. The experimental reactors included Tu or ETu at an initial concentration of 2 mM and an iodide concentration of 100 ppm or 300 ppm (supplied as potassium iodide). Reagents comprising neither Tu nor ETu were included as controls. The reactors were stirred at room temperature in a sealed condition. Solution compositions are listed in Table 3. Table 6. Solution Composition for Sealed Reactor Compatibility Tests of Reagents with a Thiocarbonyl Functional Group with Iodide

[00180] As shown in Figures 31a and 31b, the addition of thiocarbonyl compounds (TU and ETU as examples here) to an iodide medium results in slightly slower kinetics than leaching in pure iodide in sealed reactor tests. Previous studies suggest that complexation can occur between metal, iodide and thiocarbonyl species ( Bowmaker et al., Inorganic Chemistry, 48:350-368 ). Therefore, the slower leaching kinetics is possibly due to iodide entering these complexes and therefore not available for catalysis.

[00181] Given the balance between iodine, iodide and triiodide and the fact that the ferric ion can oxidize iodide to iodine by the following reaction total iodine (in this case, iodine + iodine) can only be accurately detected by in-situ oxidation prior to ICP-AES detection. Consequently, only conventional ICP-AES was performed and the results were normalized.

[00182] With reference to Figure 32, the analysis of the solutions in the sealed reactors indicates that most of the iodide remains in solution. In a practical, outdoor environment, however, iodide is expected to be oxidized to iodine by ferric acid, with iodine being lost from the leachate due to its volatility. Consequently, iodide retention under simulated outdoor conditions was tested in the presence or absence of a reagent with a thiocarbonyl functional group. Two parallel open surface evaporation tests were performed to demonstrate this phenomenon. Both pots were placed in the shade with the surface of the solution directly exposed to the air. The solution was kept stagnant (without stirring). Residual iodide was measured over a period of 72 hours. Solution compositions are listed in Table 7. Table 7. Solution composition for thiocarbonyl compound effect in iodine (open surface, stagnant solution)

[00183] With reference to Figure 33, the results indicate that when iodide enters a solution of ferric acid sulfate, it rapidly transforms into I2 and evaporates from the aqueous phase. In the presence of a reagent with a thiocarbonyl functional group, i.e. Tu, the total iodide concentration remained stable throughout the test period.

[00184] Consequently, the ability of a reagent with a thiocarbonyl functional group to facilitate the leaching of a copper sulfide in the presence of iodide was tested again in stirred reactors under open air conditions over 83 h. Each 2L reactor contained 1g of chalcopyrite concentrate which has 21.6% copper in 1L of ferric sulfate solution at pH 1.7, with a total ferric concentration of 40 mM. The experimental reactors included Tu or ETu at an initial concentration of 2 mM and an iodide concentration of 100 ppm or 300 ppm (supplied as potassium iodide). Reagents comprising neither Tu nor ETu were included as controls. The reactors were stirred at room temperature in a sealed condition. Solution compositions are listed in Table 8. Table 8. Solution composition for TU-I compatibility testing in unsealed reactors

[00185] As shown in Figures 34a and 34b, Tu and ETu had a positive effect on copper extraction from chalcopyrite in the presence of iodide at an initial concentration as high as 300 ppm. Although the amount of copper extracted increased with increasing iodide concentration, copper extraction was nevertheless higher in the presence of Tu and ETu compared to their absence.

[00186] The iodide concentration was also monitored during the course of the leaching. The results shown in Figures 35a and 35b reveal that, under open air conditions, iodide was rapidly lost from the aqueous phase. The amount of iodide in the solution decreased over time in each treatment, probably due to the volatility of the iodine generated by ferric iodide oxidation. However, the decrease in iodide over time was significantly less for solutions containing Tu or ETu. Consequently, reagents with a thiocarbonyl functional group can be useful in maintaining the stability of iodide in solution.

[00187] In general, reagents with a thiocarbonyl functional group are compatible with leach systems that have a halide component. They facilitate copper extraction in chloride and bromide leaching environments. In the iodide system, although these reagents may not facilitate extraction under sealed conditions, under actual operating conditions, such as heap leaching, these reagents may increase the stability of the iodide species in solution.

Example 10 Recovery of a thiocarbonyl functional group from a SLP.

[00188] It is desirable to recover the reagent from the PLS for recirculation for leaching. However, it was initially unclear whether it would be possible to effectively recover the reagents from PLS. Reagents with a thiocarbonyl functional group are organics that can dissolve in the organic solvent used for solvent extraction. These could potentially have the undesirable effect of removing all the catalyst from the aqueous phase, thus increasing the cost by eliminating the possibility of recycling the catalyst for leaching. This could also compromise or even destroy the effectiveness of solvent extraction.

[00189] Reagents with a thiocarbonyl functional group are complexing agents for copper. This can prevent reagents from being efficiently extracted from copper complexes in solvent extraction.

[00190] Reagents with a thiocarbonyl functional group are also surfactants. They could interact with solvent extraction organics, causing a two-phase interlayer (also known as “crude”), which could compromise solvent extraction and recovery performance.

[00191] Therefore, tests were performed to determine whether reagents with a thiocarbonyl functional group could be recovered from the PLS for recirculation to the leachate.

Example 10.1

[00192] A PLS of a chalcopyrite ore column leached with a ferric acid sulfate solution containing Tu was mixed with an organic solvent containing a copper extractor for a specified period of time. The organic solvent was a mineral oil distillate comprising aliphatic hydrocarbons including naphthenic, paraffinic and isoparaffinic components (ExxsolTM D80). The copper extractor was a weak ester modified aldoxime (Acorga® M5910). The content of copper extractor in the organic solvent was 6% v/v. The ratio of PLS to organic solvent during mixing was 5:1 v/v. PLS contained 2.5 mM free Tu equivalent.

[00193] After mixing, the organic solvent and aqueous phases were separated and samples of the aqueous phase were analyzed for reagent content. Feed PLS contained 2.5 mM free Tu equivalent.

[00194] Table 9 shows the equivalent of free Tu in the raffinate obtained after contacting the PLS with the organic solvent comprising the copper extractor for 2, 4 and 10 minutes. The table also shows the amounts of Tu and SDF in the PLS and the amounts of copper that remained in the aqueous phase (i.e. the raffinate).

[00195] The results obtained indicate that: the catalytic reagent (in the form of TU and SDF) is recovered from PLS to free copper raffinate; and increasing the mixing time of the organic solvent and PLS increases the ratio of Tu to SDS in the raffinate compared to PLS. Table 9.

Example 10.2

[00196] Synthetic solutions with different concentrations of ferric, cupric, chloride, bromide, iodide and Tu were prepared in acid sulfate medium (pH = 1.7) to simulate charged leach solutions. Treatments involving halogen species were included to simulate the PLS obtained from different halogen leaching systems. Solution compositions are listed in Table 10. Table 10. Synthetic PLS solution composition

[00197] The TU equivalent was then determined before and after solvent extraction of the synthetic PLS solutions with Acorga M5910 to form a synthetic raffinate. Elemental analysis was performed using ICP-AES. Thiocarbonyl compounds were analyzed using HPLC. Zinc dust was added to synthetic PLS and synthetic raffinate prior to analysis as a reducing agent to convert all SDF species back to TU in order to facilitate accurate determination of the recovered Tu equivalent.

[00198] Figure 36 is a bar diagram showing free Tu equivalents in a simulated PLS and in the resulting simulated raffinate after solvent extraction. The recovery percentage is calculated based on the inlet concentration. More Tu was recovered from the synthetic raffinate than from the synthetic PLS, indicating that the Tu/SDF species were complexed Tu/SDF released by copper upon removal of copper ions from solution by SX.

Example 10.3

[00199] Synthetic solutions with different concentrations of ferric, cupric, chloride, bromide, iodide and ETu were prepared in acid sulfate medium (pH = 1.7) to simulate charged leach solutions. Treatments involving halogen species were included to simulate the PLS obtained from different halogen leaching systems. Solution compositions are listed in Table 11. Table 11. Synthetic ETu solution composition

[00200] The ETu equivalent was then determined before and after solvent extraction of the synthetic PLS solutions with Acorga M5910 to form a synthetic raffinate. Elemental analysis was performed using ICP-AES. Thiocarbonyl compounds were analyzed using HPLC.

[00201] Figure 37 is a bar diagram showing free ETu in a simulated PLS and in the resulting simulated raffinate after solvent extraction. The recovery percentage is calculated based on the inlet concentration. More ETu was recovered from synthetic raffinate than synthetic PLS, indicating that the ETu species were complexed ETu released by copper upon removal of copper ions from solution by SX.

Example 11. Recovery for reagents comprising a thiocarbonyl functional group from spent leach materials

[00202] With reference to Figures 38 and 39, the inventors have currently observed that part of the Tu supplied to the material to be leached is sequestered within the materials during the initial stages of leaching. Columns of three different copper ore samples were irrigated with solutions containing Tu at a concentration of 2 mM (152 ppm). The effluent solutions were monitored for Tu equivalent concentration. When this concentration reached 2 mM, irrigation was discontinued.

[00203] Figure 38 shows graphs of the total Tu concentrations (ie, is Tu equivalent) in the effluent solutions. After approximately 28 hours of irrigation, the effluent concentrations were equal to the influent concentrations. Figure 39 shows graphs of effluent concentrations during the first of two acid water rinse stages (pH 1.8) for each ore sample. After 24 hours, effluent Tu concentrations drop to nearly zero in each case. However, as shown in Table 12, a significant amount of Tu remained sequestered in the columns even after two of these acid washes. Table 12

[00204] Figure 40 is a bar diagram that provides the data presented in Table 12 in graphical form.

[00205] Without wanting to be limited by theory, sequestration can occur through the mechanisms of adsorption to the solid surfaces of the ore and/or by diffusion in the pore spaces of the solids of the ore. It would be desirable to recover this Tu from the spent leach material to minimize catalyst costs.

[00206] Accordingly, the inventors tested the ability of a dilute solution comprising base metal ions to recover Tu from leach materials. More particularly, and referring to Table 12 and Figure 38, rinsing the columns with dilute copper sulfate solutions (eg 100 ppm, 500 ppm or 1000 ppm Cu) proved to be effective in recovering Tu from the columns. Assuming that interstitial and porous Tu is recovered during the acid rinse stages, dilute copper solutions appear to be effective in recovering Tu adsorbed on ore surfaces. This is especially important given the highly variable performance of the acid rinse alone with different ores. Furthermore, although increasing the copper concentration in the rinse solution increased the total amount of Tu recovered, even the lowest concentration of 100 ppm provided significant results.

[00207] Indeed, the skilled person will understand that solutions comprising base metal ions other than copper ions can be useful for recovering, from spent leach materials, catalyst reagents other than Tu that comprise a thiocarbonyl functional group. “Spent” or “spent”, as used herein to refer to leaching materials, may refer to materials, including ore or concentrate, that contain or contained at least one base metal sulfide that is sensitive to leaching with acid sulfate solutions comprising reagents with a thiocarbonyl functional group, and which have undergone a certain amount of leaching.

[00208] Thus, the skilled person will understand that this description relates to a general method for recovering a reagent comprising a thiocarbonyl functional group that is sequestered in leaching materials from which at least one base metal sulfide has been leached. The method comprises rinsing the leach materials with a wash solution comprising base metal ions to produce a loaded wash solution (PWS) comprising the reagent.

[00209] The skilled person will understand that the methods will work within a wide range of base metal ion concentration. In various embodiments, the base metal ion concentration in the wash solution is at least 100 ppm, at least 500 ppm, or at least 1000 ppm.

[00210] Before rinsing the leaching materials with the washing solution, the leaching materials can be rinsed with an acidic solution. The acidic solution can have a pH of around 1.8.

[00211] In various embodiments, the base metal ions include copper ions. In various embodiments, the copper ions include cupric ions.

[00212] The PWS comprising base metal ions and recovered reagent can then be added to a leachate comprising an acid sulfate solution for use in recovering at least one base metal ion from materials comprising at least one metal sulfide of basis as discussed below and exemplified more fully in PCT Patent Application No. PCT/CA2016/050444, filed April 15, 2016 and which is incorporated herein by reference.

[00213] Alternatively, the PWS can be subjected to solvent extraction steps as discussed below to remove the base metal ions before the base metal ion depleted solution is added to a leachate comprising an acid sulfate solution for use in recovering at least one base metal ion from materials comprising at least one base metal sulfide as discussed below. As Tu has a stronger effect in enhancing the base metal leaching of materials containing base metal sulfides, subsequent leaches will be enhanced by recirculating a depleted base metal ion solution with more free Tu. Accordingly, more particular aspects of this disclosure relate to the addition of a reducing agent to the base metal ion depleted solution comprising Tu and SDF to bias the equilibrium in favor of Tu prior to addition to a leachate.

[00214] The skilled person will understand that the recovered reagent can be used to supplement reagents with a thiocarbonyl functional group that are preexisting in the leachate (i.e., were previously added to the leachate). Alternatively, additional reagents with thiocarbonyl or SDS functional groups can be added to the leachate after the recovered reagent has been added.

[00215] The combination of acid and cupric washes will allow maximum recovery, perhaps complete recovery, of Tu from copper ore heaps, thus improving the economics of Tu-catalysed heap leaching.

[00216] Although specific embodiments of the invention have been described and illustrated, such embodiments should be considered only as illustrative of the invention and not as limiting, as interpreted in accordance with the appended claims.

Claims (33)

1. Use of a reagent with a thiocarbonyl functional group for extracting at least one base metal ion from a material comprising at least one base metal sulfide in an acidic solution comprising halide ions and at least one an oxidizing agent. 2. Use according to claim 1, characterized in that the reagent comprises thiourea (Tu). 3. Use according to claim 1 or 2, characterized by the fact that the reagent concentration in the acid solution is in the range of 0.002 mM to 100 mM. 4. Use according to any one of claims 1 to 3, characterized in that the at least one oxidizing agent comprises a source of ferric ions. 5. Use according to any one of claims 1 to 4, characterized in that the at least one base metal includes copper, where the at least one base metal sulfide includes at least one copper sulfide. 6. Use according to any one of claims 1 to 5, characterized in that the at least one base metal includes cadmium, where the at least one base metal sulfide includes at least one cadmium sulfide. 7. Use according to any one of claims 1 to 6, characterized in that the at least one base metal includes nickel, where the at least one base metal sulfide includes at least one nickel sulfide. 8. Use according to any one of claims 1 to 7, characterized in that the halide ions comprise chloride ions. 9. Use according to any one of claims 1 to 8, characterized in that the halide ions comprise bromide ions, iodide ions or a combination thereof. 10. Use according to any one of claims 1 to 8, characterized by a concentration of halide ions in the acid solution of 1 g/L or more. 11. Method of recovering at least one base metal ion from a material comprising at least one base metal sulfide, comprising: contacting the material with an acidic solution comprising halide ions, at least one oxidizing agent and a reagent having a thiocarbonyl functional group for producing a charged solution comprising base metal ions; and recovering the at least one base metal ion from the loaded solution. 12. Method according to claim 11, characterized in that the reagent comprises thiourea (Tu). 13. Method according to claim 11, characterized in that the reagent comprises thioacetamide (TA). 14. Method according to claim 11, characterized in that the reagent comprises sodium dimethyldithiocarbamate (SDDC). 15. Method according to claim 11, characterized in that the reagent comprises ethylene trithiocarbonate (ETC). 16. Method according to claim 11, characterized in that the reagent comprises thiosemicarbazide (TSCA). 17. Method according to claim 11, characterized in that the reagent comprises ethylenethiourea (ETu). 18. Method according to claim 11, characterized in that the reagent comprises N-N' substituted thioureas, 2,5-dithiobiurea; dithiobiuret; thiosemicarbazide purum, thiosemicarbazide; thioacetamide; 2-methyl-3-thiosemicarbazide; 4-methyl-3-thiosemicarbazide; purum vinylene trithiocarbonate; vinylene trithiocarbonate; 2-cyanothioacetamide; ethylene trithiocarbonate; potassium ethyl xanthogenate; dimethylthiocarbamoyl chloride; dimethyldithiocarbamate; dimethyl trithiocarbonate; N,N-dimethylthioformamide; 4,4-dimethyl-3-thiosemicarbazide; 4-ethyl-3-thiosemicarbazide; O-isopropylxantic acid; ethyl thiooxamate; ethyl dithioacetate; pyrazine-2-thiocarboxamide; diethylthiocarbamoyl chloride; diethyldithiocarbamate; tetramethylthiuram monosulfide; tetramethylthiuram disulfide; pentafluorophenyl chlorothionoformate; 4-fluorophenyl chlorothionoformate; O-phenyl chlorothionoformate; phenyl chlorothionoformate; 3,4-difluorothiobenzamide; 2-bromothiobenzamide; 3-bromothiobenzamide; 4-bromothiobenzamide; 4-chlorothiobenzamide; 4-fluorothiobenzamide; thiobenzoic acid; thiobenzamide; 4-phenylthiosemicarbazide; O-(p-tolyl) chlorothionoformate; 4-bromo-2-methylthiobenzamide; 3-methoxythiobenzamide; 4-methoxythiobenzamide; 4-methylbenzenethioamide; thioacetanilide; thiosemicarbazone salicylaldehyde; indole-3-thiocarboxamide; S-(thiobenzoyl)thioglycolic acid; 3-(acetoxy)thiobenzamide; 4-(acetoxy)thiobenzamide; N'-[(e)-(4-chlorophenyl)methylidene] methyl hydrazonothiocarbamate; 3-ethoxythiobenzamide; 4-ethylbenzene-1-thiocarboxamide; tert-butyl 3-[(methylsulfonyl)oxy]-1-azethanecarboxylate; diethyldithiocarbamic acid; 2-(phenylcarbonothioylthio)propanoic acid; 2-hydroxybenzaldehyde N-ethylthiosemicarbazone; (1R,4R)-1,7,7-trimethylbicyclo[2.2.1]heptane-2-thione; tetraethylthiuram disulfide; 4'-hydroxybiphenyl-4-thiocarboxamide; 4-biphenylthioamide; dithizone; 4'-methylbiphenyl-4-thiocarboxamide; tetraisopropylthiuram disulfide; anthracene-9-thiocarboxamide; phenanthrene-9-thiocarboxamide; sodium dibenzyldithiocarbamate; 4,4'-bis(dimethylamino)thiobenzophenone, or a combination thereof. 19. Method according to any one of claims 11 to 18, characterized in that the initial concentration of the reagent in the acid solution is in the range of 0.002 mM to 100 mM. 20. Method according to any one of claims 11 to 18, characterized in that the initial concentration of the reagent in the acid solution is in the range of 0.002 mM to 60 mM. 21. Method according to any one of claims 11 to 18, characterized in that the initial concentration of the reagent in the acidic solution is in the range of 0.002 mM to 30 mM. 22. Method according to any one of claims 11 to 18, characterized in that the initial concentration of the reagent in the acid solution is in the range of 0.002 mM to 2.0 mM. 23. Method according to any one of claims 11 to 22, characterized in that the at least one oxidizing agent comprises a source of ferric ions. A method according to any one of claims 11 to 23, characterized in that the at least one base metal includes copper, wherein the at least one base metal sulfide includes at least one copper sulfide. A method according to any one of claims 11 to 24, characterized in that the at least one base metal includes cadmium, wherein the at least one base metal sulfide includes at least one cadmium sulfide. 26. Method according to any one of claims 11 to 25, characterized in that the at least one base metal includes nickel, wherein the at least one base metal sulfide includes at least one nickel sulfide. 27. Method according to any one of claims 11 to 26, characterized in that it maintains the operational potential of the acid solution above 500 mV vs Ag / AgCl. 28. Method according to any one of claims 11 to 27, characterized in that the acid solution comprises a ferrous sulfate solution. 29. Method according to any one of claims 11 to 27, characterized in that the halide ions comprise chloride ions. 30. Method according to any one of claims 11 to 29, characterized in that the halide ions comprise bromide ions. 31. Method according to any one of claims 11 to 30, characterized in that the halide ions comprise iodide ions. 32. Method according to any one of claims 11 to 31, characterized by a concentration of halide ions in the acid solution of 1 g/L or more. 33. Method according to any one of claims 11 to 32, characterized in that the reagent does not complexify / precipitate with the at least one base metal ion.