PROCESS FOR LEACHING METAL SULFIDES WITH REAGENTS CONTAINING THIOCARBONYL FUNCTIONAL GROUPS

Abstract

This application relates to methods for recovering metals from metal sulfides involving contacting the metal sulfide with an acidic sulfate solution containing ferric sulfate and a reagent having a thiocarbonyl functional group, wherein the concentration of the reagent in the acidic sulfate solution is sufficient to increase the extraction rate of metal ions relative to an acidic sulfate solution not containing the reagent, in order to produce an impregnated solution containing the metal ions.

Description

PROCESS FOR LEACHING METAL SULFIDES WITH REAGENTS CONTAINING THIOCARBONYL FUNCTIONAL GROUPS BACKGROUND This application claims priority over U.S. patent application No. 62 / 149.015, filed on April 17, 2015, the contents of which are incorporated herein by reference. 1. Field of dissemination This disclosure relates to methods for leaching metals from metal sulfide-containing ores. More specifically, it relates to a hydrometallurgical process for extracting metals from metal sulfide-containing ores using reagents containing a thiocarbonyl functional group. 2. Description of previous art Aqueous mineral processing offers 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 ores, is the low extraction rates observed. It is therefore desirable to develop a process that can achieve high metal extraction rates within timeframes of interest to industry. For example, chalcopyrite is a semiconductor and therefore corrodes electrochemically in oxidizing solutions. In ferric sulfate media, the general leaching reaction is as follows: CuFeS2(s) + 2 Fe2(SO4)3(a) CuSO4(a) + 5 FeSO4(a) + 2S° (s) This reaction can be represented as a combination of anodic and cathodic half-cell reactions: Anodic half-cell reaction: CuFeS2—> Cu2+ + Fe2+ + 2 + 4 et CQonn / nznz / Ε / γΐΛΐ Cathodic half-cell reaction: Fe3+ + 4 e“ —► 4 Fe2+ A fundamental problem with chalcopyrite oxidation is that chalcopyrite mineral surfaces become resistant to electrochemical decomposition at solution potentials above a certain level (generally considered to be between approximately 550 mV and 600 mV with respect to the Ag / AgCl electrode). A widely held view is that this results from the formation of some kind of passivating film on the mineral surface, consisting largely of an altered, partially iron-free form of chalcopyrite. It is therefore desirable to provide leaching processes in which this passivation is reduced or prevented. Some work has been done in extractive hydrometallurgy to recover precious metals such as gold and silver from copper concentrates or chalcopyrite residues after copper extraction. Deschenes and Ghali (Hydrometallurgy 20:129-202) demonstrated the potential application of thiourea in an acid sulfate leach of sulfide concentrates, such as those containing chalcopyrite, to selectively recover gold and silver. Thiourea is an organosulfur compound that has a thiocarbonyl functional group. However, thiourea did not appear to have any effect on copper recovery from copper sulfides. SUMMARY OF THE INVENTION This disclosure relates, at least in part, to the unexpected discovery that several reagents comprising a thiocarbonyl functional group (e.g., thiourea) can be used to facilitate the leaching of metal from various metal sulfides (e.g., copper from chalcopyrite) with acidic sulfate leaching solutions. When added in small amounts, these reagents can increase the rate of metal leaching compared to that observed in their absence. This disclosure relates to a method for recovering at least one metal from an ore containing at least one metal sulfide. The method involves contacting the ore with an acidic sulfate solution containing ferric sulfate and a reagent having a thiocarbonyl functional group to extract metal ions from at least one metal sulfide, wherein the concentration of the reagent in the acidic sulfate solution is sufficient to increase the rate of metal ion extraction relative to an acidic sulfate solution not containing the reagent, such as to produce an impregnated solution containing metal ions. The method further involves recovering at least one metal from the impregnated solution.The at least one metal includes: copper, where the at least one metal sulfide includes chalcopyrite, covellite, bornite, enargite, a copper sulfide of formula CuxSy where the ratio of x and y is between 1 and 2, or one of their combinations; cadmium, where the at least one metal sulfide is greenockite; nickel, where the at least one metal sulfide is pentlandite, violarite, or one of their combinations; or one of their combinations. The concentration of the reagent may be between approximately 0.2 mM and approximately 30 mM. This disclosure also relates to a method for recovering at least one metal from an ore containing at least one metal sulfide. The method involves contacting the metal sulfide with an acidic sulfate solution containing a reagent with a thiocarbonyl functional group, wherein the initial concentration of the reagent in the acidic sulfate solution is between about 0.2 mM and about 30 mM or less, in order to produce an impregnated solution containing metal ions. The method further involves recovering copper from the impregnated solution.The at least one metal includes: copper, where the at least one metal sulfide includes chalcopyrite, covellite, bornite, enargite, a copper sulfide of formula CuxSy where the ratio between x and y is between 1 and 2, or one of their combinations; cadmium, where the at least one metal sulfide is greenockite; nickel, where the at least one metal sulfide is pentlandite, violarite, or one of their combinations; or one of their combinations. In the methods described above, the concentration of the reagent in the acidic sulfate solution can be found to be between about 0.2 mM and about 20 mM, between about 0.2 mM and about 10 mM, between about 0.2 mM and about 5 mM, between about 0.2 mM and about 4 mM, between about 0.2 mM and about 3 mM, between about 0.2 mM and about 2 mM, between about 0.2 mM and about 1.5 mM, between about 0.2 mM and about 1.0 mM, or between about 0.2 mM and about 0.5 mM. When the metal is a copper sulfide of formula CuxSy where the ratio between x and y is between 1 and 2, the copper sulfide may include chalcocite, djurleite, digenite or one of their combinations. In the methods described above, the reagent can be thiourea (Tu), thioacetamide (TA), sodium dimethyl-dithiocarbamate (SDDC), ethylene trioxycarbonate (ETC), thiosemicarbazide (TSCA) or one of their combinations. This disclosure also relates to a method for recovering a metal from an ore containing at least one metal sulfide. The method involves contacting an ore with an acidic sulfate solution comprising ferric sulfate and formamidine disulfide (FDS) to produce an impregnated solution containing metal ions. The method further involves recovering the metal from the impregnated solution. The at least one metal includes: copper, where the at least one metal sulfide includes chalcopyrite, covellite, bornite, enargite, a copper sulfide of formula CuxSy where the ratio of x and y is between 1 and 2, or one of their combinations; cadmium, where the at least one metal sulfide is greenockite; nickel, where the at least one metal sulfide is pentlandite, violarite, or one of their combinations; or one of their combinations.The concentration of FDS in the acidic sulfate solution can range from approximately 0.1 mM to approximately 15 mM, from approximately 0.1 mM to approximately 10 mM, from approximately 0.2 mM to approximately 5 mM, from approximately 0.1 mM to approximately 2.5 mM, from approximately 0.1 mM to approximately 2 mM, from approximately 0.1 mM to approximately 1.5 mM, from approximately 0.1 mM to approximately 1.0 mM, from approximately 0.1 mM to approximately 0.5 mM, or from approximately 0.1 mM to approximately 0.25 mM. When the metal is a copper sulfide of the formula CuxSy, where the ratio of x and y is between 1 and 2, the copper sulfide may include chalcocite, djurleite, digenite, or one of their combinations.The concentration of FDS in the acidic sulfate solution may be sufficient to provide enough thiourea to increase the rate of metal ion extraction compared to an acidic sulfate solution that does not contain the reagent, in order to produce the impregnated solution containing metal ions. In the methods described above, the ore can be supplied as coarse particles, which may be agglomerated. Ferric ions can be used to oxidize the metal sulfide. In the methods described above, the ferric ions can be generated, at least in part, by bacteria. The methods may involve percolation leaching. Percolation leaching can be heap leaching. Percolation leaching can be vat leaching. Leaching can be tank leaching. Metal recovery from an impregnated solution may include solvent extraction and electrowinning. ) CQonn / nznz / E / YiAi Other aspects and features of the present invention will become evident to those skilled in the art upon reviewing the following description of specific embodiments of the invention together with the accompanying figures. BRIEF DESCRIPTION OF THE DRAWINGS In the drawings illustrating the ways of carrying out the invention: - Figure 1 is a diagram showing the effect of thiourea concentration on the mixed potential and solution current density (idissoi) of the CuFeSi electrode; - Figure 2 is a bar chart showing the electrochemical dissolution rates of a CuFeSi electrode in sulfuric acid solution with a pH of 2 at a temperature of 25 °C with varying initial concentrations of thiourea, formamidine disulfide (FDS) and Fe(lll); - Figure 3 is a schematic diagram for the leaching column used with respect to the leaching experiments related to Figures 4, 5 and 6; - Figure 4 is a graph showing the effect of thiourea concentration on copper leaching from Ore A in column leaching experiments; - Figure 5 is a graph showing the effect of thiourea concentration on copper leaching from Ore B in column leaching experiments; - Figure 6 is a graph showing the effect of thiourea concentration on copper leaching from Ore C in column leaching experiments; - Figure 7 is a graph showing the effect of thiourea concentration on the copper leaching rate from Ore C in column leaching experiments; - Figure 8 is a graph showing the effect of thiourea concentration on the oxidation-reduction potential (ORP) over time; - Figure 9 is a graph showing the effect of thiourea concentration on copper dissolution for a coarse sample of Ore A in bottle roller tests; - Figure 10 is a graph showing the effect of thiourea concentration on copper dissolution for a coarse sample of Ore B in bottle roller tests; Figure 11 is a graph showing the effect of Tu addition on various Cu(I)-containing minerals. The rhombuses refer to bomite, the triangles to covellite, the inverted triangles to chalcocite, and the squares to chalcopyrite. The hollow symbols refer to treatments without thiourea, while the solid symbols refer to solutions treated with minerals having an initial thiourea concentration of 2 mM. - Figure 12 is a graph showing the effect of thiourea on the extraction of cadmium from greenockite; - Figure 13 is a graph showing the effect of thiourea on the extraction of copper from enargite; - Figure 14 is a graph showing the effect of thiourea on the extraction of nickel from violarite; - Figure 15 is a graph showing the percentage of Cu ions that remain in solution after the addition of various amounts of thiourea; - Figure 16 is a graph showing the extraction of Cu from chalcopyrite under various thiourea dosages; - Figure 17 is a graph showing the relationship between thiourea dosage and Cu extraction after 172 hours; Figure 18 is a graph showing the leaching of copper from chalcopyrite in stirred reactor tests using reagents comprising thiocarbonyl functional groups. Circles refer to thiourea, triangles to TA, inverted triangles to SDDC, rhombuses to ETC, stars to TSCA, and squares to controls. Figure 19 is a graph showing the leaching of copper from covellite in stirred reactor tests using reagents comprising thiocarbonyl functional groups. Circles refer to thiourea, triangles to TA, rhombuses to SDDC, and squares to controls. Figure 20 is a graph showing the leaching of copper from bornite in stirred reactor tests using reagents comprising thiocarbonyl functional groups. The triangles refer to thiourea, the circles refer to TA, and the squares refer to the controls. Figure 21 is a graph showing the leaching of copper from enargite in stirred reactor tests using reagents comprising thiocarbonyl functional groups. Circles refer to thiourea, triangles refer to TA, inverted triangles refer to ETC, and squares refer to controls. Figure 22 is a graph showing the leaching of copper from chalcopyrite in stirred reactor tests using reagents comprising thiocarbonyl, urea, and carbon disulfide functional groups. Circles refer to thiourea, triangles to controls, inverted triangles to TA, rhombuses to thiourea, stars to ETC, and squares to carbon disulfide. - Figure 23a is a graph comparing the leaching of copper from chalcopyrite (circles) or bomite (triangles) using leaching solutions with an initial concentration of 2 mM Tu (full symbols) or an initial concentration of 1 m FDS (hollow symbols); - Figure 23b is a graph comparing the leaching of copper from covellite (circles) or chalcocite (triangles) using leaching solutions with an initial concentration of 2 mM thiourea (full symbols) or an initial concentration of 1 mM FDS (hollow symbols); - Figure 24 is a graph that tracks bacterial activity and SDS content using ORP and HPLC; and Figure 25 is a graph showing the bioleaching of CuFeSi using only Fe3+ (between days 0 and 50) and using Fe3++ Tu (between days 90 and 150) in closed-loop experiments. DETAILED DESCRIPTION This disclosure relates to methods for recovering metal from a metal sulfide ore, and in particular to the unexpected discovery that several reagents containing a thiocarbonyl functional group, such as thiourea (also known as thiocarbamide), can be used to facilitate the leaching of metal from a metal sulfide ore with acidic sulfate leaching solutions. These reagents can increase the rate of metal sulfide leaching. These methods can be particularly useful for recovering metal from low-grade ores that do not contain the metal sulfide mineral in large proportions. The method involves contacting the copper sulfide mineral with an acidic sulfate solution containing a reagent with a thiocarbonyl functional group. Minerals Chalcopyrite (CuFeSi) Chalcopyrite leaching is achieved in an acidic solution of ferric sulfate according to the following reaction formula: i CQonn / nznz / E / YiAi CuFeS2+ 4 Fe3+Cu2++ 5 Fe2++ 2 S° Covellite (CuS) The leaching of covellite in ferric sulfate solution is carried out according to the following reaction formula: CuS + 2 Fe3+Cu2++ 2 Fe2++ S° Chalcocite (CU2S) The leaching of chalcocite in ferric solution is carried out according to the following formula: Cu2S + 2 Fe3+Cu2++ 2 Fe2++ CuS Experts in the field understand that these chalcocite ores often contain a mixture of minerals with the formula CuxSy, where the ratio of x and y is between 1 and 2. Additional minerals included within this formula are digenite and djurleite. Bornite (CusFeS4) Bornite is a significant copper mineral that generally coexists with chalcopyrite. The leaching process of bornite in ferric solution is described in two stages: Cu5FeS4+ 4 Fe3+Cu3FeS4+ 2 Cu2++ 4 Fe2+ Cu3FeS4+ 8 Fe3+3 Cu2++ 9 Fe2++ 4 S° Enargite (Cu3AsS4) Unlike other copper minerals mentioned previously (chalcopyrite, covellite, chalcocite, and bornite), the copper in enargite is primarily Cu(II) rather than Cu(I). This difference in the copper's oxidation state also influences the leaching kinetics under catalyzed conditions. A previous study demonstrated that enargite leaching at atmospheric pressure is extremely slow. The dissolution of enargite in a ferric sulfate medium can occur via several pathways. Two of these are described below: CQonn / nznz / E / γΐΛΐ Cu3AsS4+ 20 H2O + 35 Fe3+ Cu2++ AsO43“ + 4 SO42' + 40 H++ 35 Fe2+ Cu3AsS4+ 4 H2O + 11 Fe3+ Cu2++ AsO43' + 4 S° + 8 H++ 11 Fe2+ Greenockite (CdS) Cadmium metal and compounds are primarily used in alloys, coatings, batteries, and plastic stabilizers. There are no mines that specifically extract cadmium. Cadmium sulfide is generally associated with zinc sulfides and is recovered as a byproduct of zinc leaching from roasted sulfide concentrates. Violarite (FeNi2S4) Violarite is a nickel(III) sulfide mineral that is generally associated with primary pentlandite nickel sulfide ores. Reagents Those knowledgeable in the field will also understand that reagents with a thiocarbonyl functional group include, but are not limited to, thiourea (Tu), thioacetamide (TA), sodium dimethyldithiocarbamate (SDDC), ethylene trithiocarbonate (ETC), and thiosemicarbazamide (TSCA). The following is a non-exclusive list of additional compounds that have a thiocarbonyl functional group: isothiourea; NN'-substituted thioureas; 2,5-dithiobiurea; dithiobiuret; thiosemicarbazide; methyl elorothiolformate; dithiooxamide; thioacetamide; 2-methyl-3-thiosemicarbazide; 4-methyl-3-thiosemicarbazide; 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; Oisopropylxanthic acid; ethyl thioxamate; ethyl dithioacetate; pyrazine-2-thiocarboxamide; diethylthiocarbamoyl chloride; diethyldithiocarbamate; tetra-methylthiuram monosulfide; tetramethylthiuram disulfide; pentafluorophenyl chlorothionoformate; 4-fluorophenyl chlorothionoformate; O-phenyl chlorothionoformate;O-phenyl chlorothio-formate; phenyl chlorodithioformate; 3,4-difluorothiobenzamide; 2-bromothiobenzamide; 3-bromothiobenzamide; 4-bromothiobenzamide; chlorothiobenzamide; 4-fluorothiobenzamide; thiobenzoic acid; thiobenzamide; 4-phenylthiosemicarbazide; O-(p-tolyl) chlorothio-formate; 4-bromo-2-methylthiobenzamide; 3-methoxythiobenzamide; 4-methoxythiobenzamide; 4-methylbenzenethioamide; thioacetanilide; salicylaldehyde thiosemicarbazone; indole-3-thiocarboxamide; S-(thiobenzoyl)thioglycolic acid; 3-(acetoxy)thiobenzamide; 4-(acetoxy)thiobenzamide; Methyl N'-[(e)-(4-chlorophenyl)methyliden]hydrazonothiocarbamate; 3-ethoxythiobenzamide; 4-ethylbenzen-1-thiocarboxamide; tert-butyl3-[(methylsulfonyl)-oxy]-1-azethanecarboxylate; diethyldithiocarbamic acid; 2(phenylcarbonothioylthio)propanoic acid; 2-hydroxybenzaldehyde N-ethylthiosemi-carbazone; (1R,4R)1,7,7-trimethyl-bicyclo[2.2.1]heptane-2-thione; tetraethyl thiurama 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 on the market through Sigma Aldrich, for example. Each Tu, TA, SDDC, ETC, and TSCA possesses a thiocarbonyl functional group with a sulfur atom that 1) carries a partial negative charge, 2) has a negative electrostatic potential surface, and 3) has an empty antibonding π* orbital as its lowest energy unoccupied molecular orbital (LUMO). Consequently, those skilled in the art may reasonably expect that other reagents, including those listed above, that meet this criterion and are sufficiently soluble in water, may be useful in carrying out the methods disclosed herein (provided they do not form complexes with the metal or the iron oxidizing agent to form precipitates). Those skilled in the art may identify potentially useful reagents and evaluate them for effectiveness with any particular ore, if any. For example, thiourea has a thiocarbonyl functional group with sulfur carrying a partial charge of -0.371, a negative electrostatic potential around the sulfur, and an antibonding π* orbital as its LUMO. Therefore, thiourea satisfies all three criteria and has demonstrated a catalytic effect. Thioacetamide has a structure similar to that of thiourea, but with a CH3 side chain instead of NH2. It has a thiocarbonyl functional group with the sulfur carrying a partial charge of -0.305, which is slightly less than that of thiourea, a negative electrostatic potential. CQonn / nznz / E / γΐΛΐ around the sulfur and an antibonding π* orbital as its LUMO. Consequently, thioacetamide satisfies all three criteria and has demonstrated a catalytic effect. ETC differs from thiourea and thioacetamide in that it does not contain a thioamide group. It features a thiocarbonyl functional group with the two sulfur atoms σ-bonded to the carbon as a side chain. The sulfur in the thiocarbonyl group carries a partial charge of -0.122, which is much lower than that of thiourea, a negative electrostatic potential around the sulfur, and an antibonding π* orbital as its LUMO. Consequently, ETC satisfies all three criteria and has demonstrated a catalytic effect. In comparison, urea has a carbonyl functional group with a C=O bond instead of a C=S bond. The oxygen in the C=O bond carries a partial charge of -0.634 and a negative electrostatic potential around it, which is very similar to that of the sulfur atom in thiourea. However, its LUMO does not contain an antibonding π* orbital. Consequently, urea is not expected to exert a catalytic effect on metal leaching, a finding confirmed for chalcopyrite by the results of the stirred reactor test shown in Figure 22. Carbon disulfide (CS2) contains two thiocarbonyl functional groups. Although the sulfur atoms in each functional group contain antibonding π* orbitals as their LUMO, these carry a partial positive charge of +0.012. Therefore, carbon disulfide is not expected to exert a catalytic effect, which is confirmed with respect to chalcopyrite by the results of the stirred reactor test shown in Figure 23. The reagent must also be water-soluble. ETC, for example, is only moderately soluble in water, which may explain why it appears less effective than thiourea in leaching copper from chalcopyrite. Ideally, the reagent should not form complexes / precipitates with Fe2+ / Fe3+ ions. TSCA, for example, is capable of forming a red complex with Fe3+ in solution, which may explain why it is less effective than thiourea in leaching copper from chalcopyrite. The reagent must also not form complexes / precipitates with the 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 thioactamide forms an insoluble complex with Cd+ ions and therefore cannot be used for leaching cadmium sulfide minerals such as greenockite. Again, experts in the field will appreciate that not all compounds that CQonn / nznz / E / γΐΛΐ, which comprise a thiocarbonyl functional group, will be useful for increasing the extraction rate of metals from a metal sulfide. Furthermore, those skilled in the art will appreciate that a reagent that works to increase the extraction rate of one metal from a metal sulfide may not be useful for increasing the extraction rate of a metal from a different metal sulfide. Again, those skilled in the art will be able to identify potentially useful reagents and evaluate them to determine their effectiveness with any particular ore, if any. Formamidine disulfide (FDS) Formamidine disulfide (FDS) is generated by the oxidation of thiourea. In the presence of an oxidant such as ferric sulfate, thiourea will be partially oxidized to form formamidine disulfide (FDS) according to the following half-cell reaction: SC(NH2)2^[(NH2)2CS]22++ 2 e- E1 FDS does not contain a thiocarbonyl functional group but does contain a sulfur-sulfur sigma bond. An equilibrium exists between FDS and thiourea in a ferric sulfate solution such that a leaching solution prepared with FDS instead of thiourea will provide the thiourea necessary to catalyze the leaching of the metal sulfide. That is, one molecule of FDS will dissociate into two molecules of thiourea upon dissolving in the ferric sulfate leaching solution. Consequently, a leaching solution that uses thiourea as a reagent with a thiocarbonyl functional group can be efficiently prepared using either thiourea or FDS. Those skilled in the field will understand that, due to this equilibrium, the concentration of thiourea (and of FDS) can fluctuate over time. Consequently, the term concentration, as used herein, refers to the concentration of thiourea in the leaching solution related to the amount of thiourea present in the solution as if all the FDS in the solution were dissociated into thiourea (i.e., ignoring the interconversion between the two forms). Similarly, the term concentration, as used herein, refers to the concentration of FDS in the leaching solution related to the amount of FDS present in the solution as if all the thiourea in the solution had been converted into FDS (i.e., ignoring the interconversion between the two forms). t CQonn / nznz / E / γΐΛΐ The term "initial concentration" is used herein to refer to the initial concentration of the reagent at the time the leaching solution is applied to the ore sample. However, those skilled in the art will understand that the reagent concentration may decrease over time (e.g., through precipitation or decomposition) as the solution percolates through the column or heap. Consequently, those skilled in the art will appreciate that the processes disclosed herein should work to increase the rate of metal extraction from the metal sulfide provided the reagent concentration remains within a suitable range during some part of the percolation through the ore. In the presence of FDS and ferric sulfate (or another suitable oxidant), an anodic dissolution of a copper sulfide mineral such as chalcopyrite can be produced according to the following two reactions, with oxidation of the chalcopyrite by either FDS or ferric sulfate, respectively: CuFeS2(s) + 2 [(NH2)2CS]2SO4(aqueous) —> CuS (aqueous) + FeS (aqueous) + 2 S°(s) + 4 SC(NH2)2(aqueous) CuFeS2(s) + 2 Fe2(SO4)3(a) CuSO4(a) + 5 FeSO4(a) + 2 S°(s) After the chalcopyrite has oxidized and the copper has been leached from the concentrate, it is convenient to recover the copper from the impregnated leaching solution. The methods described herein involve two basic steps: leaching and metal recovery (e.g., by solvent extraction-electrodeposition or SX-EW). The leaching process can be carried out as percolation leaching (such as accumulation leaching), vat leaching, or tank leaching, as it is known in the industry. For the purposes of this disclosure, the expressions "contains" and "comprises" are used in a non-exclusive sense to indicate that what follows the word is included, but things not specifically mentioned are not excluded. Reference to an item using the indefinite articles "a" or "an" does not preclude the possibility of more than one item being present, unless the context clearly requires that there is one and only one of those items. t CQonn / nznz / E / YiAi The term percolation leaching, as used herein, refers to the selective extraction of a mineral by allowing a suitable solvent to filter through a mass or heap of material containing the desired soluble mineral, for example, column leaching or heap leaching. The term column leaching, as used herein, refers to leaching through the use of a long, narrow column in which the ore sample is brought into contact with the solution to measure the effects of typical variables found in actual heap leaching. The term heap leaching, as used here, refers to a process by which metals are extracted from the ore in which they are found, that is, without sorting or concentration. Heap leaching is often chosen for its efficiency and cost-effectiveness. Once removed from the ground, the ore is typically passed through a crusher to break it into smaller particles (although heap leach ore may be in the state as extracted from the mine, where the ore is leached at the size it was blasted to without further crushing). Heap ore may be the product of primary, secondary, or tertiary crushing. Traditionally, the crushed particles are then piled or stacked into large heaps. A persistent cause of heap leaching failure is the presence of excessive fines in the bed material. Excessive fines result in a material with low permeability, and therefore, the leachate filtration rate is too slow, or the ore-solution contact is insufficient, for economical heap leaching. Consequently, the efficiency of heap leaching can be increased by agglomeration after crushing. The term agglomeration, as used here, refers to a technique that binds the fines or particles together to create a larger product. Agglomeration can be achieved through various methods known in the industry.Typically, agglomeration in heap leaching is carried out in an agglomerating drum with sulfuric acid and without a binder, or on conveyor belts by spraying an acid onto the ore at drip points. The heap is irrigated with a solution that depends on the type of ore extracted. Preferably, the acid for leaching will be generated by bacteria using processes known in the industry. Alternatively, additional acid may be added as needed. t CQonn / nznz / E / γΐΛΐ The irrigated solution is allowed to percolate through the ore and drain to the bottom of the heap. The ore heap is placed on an impermeable layer, such as a plastic sheet, which collects the impregnated leaching solution as it drains and directs it to a collection pond. Once collected, the solution is pumped to a recovery plant to extract the copper using a solvent extraction and electrowinning (SX-EW) method. By applying the methods disclosed herein to heap leaching, the ore containing a suitable sulfide mineral is selectively leached in the presence of the acid sulfate and the reagent with a thiocarbonyl functional group. The concentration of the reagent with a thiocarbonyl functional group in the leaching solution may be around 30 mM or perhaps even higher. Those skilled in the art will understand that it is only necessary for the reagent concentration to be within a range sufficient to increase the leaching rate of the metal sulfide. Furthermore, although the results presented here indicate that reagent concentrations of around 30 mM or less are low enough to facilitate the leaching of the metal from a particular metal sulfide, concentrations of 30 mM may not be economically feasible at present. Consequently, it is preferable to use lower reagent concentrations that are more economically and operationally viable, for example, around 20 mM or less, around 10 mM or less, around 5 mM or less, around 4 mM or less, around 3 mM or less, around 2 mM or less, around 1.5 mM or less, around 1 mM or less, around 0.9 mM or less, around 0.8 mM or less, around 0.7 mM or less, around 0.6 mM or less, around 0.5 mM or less, around 0.4 mM or less, 0.3 mM or less, or around 0.2 mM. Consequently, the concentration of the reagent in the acidic sulfate solution can be found to be between approximately 0.2 mM and approximately 0.3 mM, between approximately 0.2 mM and approximately 0.4 mM, between approximately 0.2 mM and approximately 0.5 mM, between approximately 0.2 mM and approximately 0.6 mM, between approximately 0.2 mM and approximately 0.7 mM, between approximately 0.2 mM and approximately 0.8 mM, between approximately 0.2 mM and approximately 0.9 mM, between approximately 0.2 mM and approximately 1.0 mM, between approximately 0.2 and approximately 1.5 mM, between approximately 0.2 and approximately 2.0 mM, between approximately 0.2 and approximately 2.5 mM, between approximately 0.2 and approximately 3 mM, between approximately 0.2 and approximately of 4 mM, between about 0.2 and about t CQonn / nznz / E / YiAi of 5 mM, between about 0.2 and about 10 mM, between about 0.2 and about 20 mM, or between about 0.2 and about 30 mM. The leaching process can be carried out at temperatures between 0 °C (i.e., at the freezing point of water) and 80 °C. However, the process could normally be carried out at room temperature and atmospheric pressure. Following the leaching process, copper can be extracted from the leaching solution. After the separation of solids and liquids, i.e., the drainage of the copper-impregnated leaching solution from the ore heap, the impregnated solution is preferably subjected to conventional solvent extraction and electrowinning to produce pure copper cathodes according to the following general reaction: SX-EW: CuSO4(a) + H2O (1) Cu (s) + H2SO4(a) + 1 / 2 O2(g) Reagents with a thiocarbonyl functional group in the impregnated leaching solution should not present any problems in the electrowinning operation and, in fact, may even be useful as a leveling agent. The raffinate containing the thiourea can then be recirculated through the heap for further leaching. The recirculated leaching solution can also be supplemented with thiourea to reach the desired initial thiourea concentration for leaching. Examples To facilitate the extraction of metal ions from the minerals listed above, reagents containing a thiocarbonyl functional group were added to acidic ferric sulfate solutions as catalysts. In the experiments reported herein, reagents containing thiocarbonyl functional groups were found to have a positive catalytic effect on mineral extraction. Among all the reagents, thiourea consistently provided the highest catalytic yield. Consequently, thiourea was the most studied reagent among those identified. However, the results of experiments with other reagents containing thiocarbonyl functional groups are provided to compare their catalytic effects. Ferric sulfate solution (FDS), which does not contain a thiocarbonyl functional group but has a comparable catalytic effect to thiourea, was studied as a special case due to its equilibrium with thiourea. The leaching reactions were carried out at atmospheric pressure in a variety of ore compositions, reagent concentrations, ferric concentrations and under other diverse conditions, as described below. Example 1 Extraction of copper from chalcopyrite using thiourea Example 1.1 The effect of thiourea on the electrochemical behavior of a chalcopyrite electrode in a conventional three-electrode glass-jacketed cell was studied. A CuFeSa electrode was used as the operating electrode, a saturated calomel electrode (SCE) as the reference, and a graphite rod as the counter electrode. The CuFeSa electrode was polished using 600 and 1200 grit carbide paper. All experiments were conducted at 25 °C using a temperature-controlled water bath. The electrolyte composition was 500 mM H₂SO₄, 20 mM FeASO₄, and between 0 mM and 100 mM thiourea. Before starting any measurements, sodium bubbles were introduced into the solutions for 30 minutes to reduce the dissolved oxygen 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 using a 5 mV AC sinusoidal perturbation between 10 kHz and 10 mHz. Linear polarization resistance (LPR) tests were also performed using a sweep speed of 0.05 mV / s at ±15 mV OCP. Linear potential sweeps were performed at electrode potentials of +15 mV of the calculated OCP for each thiourea concentration. All sweeps showed linear behavior within the analyzed electrode potential range. An increase in the slope of the experimental graphs was observed with increasing thiourea concentration. The slope of these curves was used to calculate the polarization resistance (Rct) at each concentration. These values ​​were then used to calculate the solution current density values ​​used in Equation 1. RT j CQonn / nznz / E / γΐΛΐ Equation (1) Figure 1 shows the effect of thiourea on the dissolution current density and mixed potential of the CuFeS₂ electrode, indicating that a maximum dissolution current density was achieved at a thiourea concentration of 30 mM. Increasing the thiourea concentration to 100 mM resulted in a reduction in both the current density and mixed potential of the CuFeSi electrode. Furthermore, after immersing the CuFeSi electrode in the 100 mM thiourea solution, a copper-like film was observed on the electrode surface, which could only be removed by sanding the electrode with carbide paper. Example 1.2 Figure 2 is a bar graph showing the effect of initial FDS or thiourea concentration on the electrochemical dissolution of a chalcopyrite electrode in sulfuric acid solution at pH 2 and 25 °C. A 10 mM thiourea concentration in the leaching solution resulted in a 6-fold increase in the dissolution rate compared to the case without thiourea, and a 5 mM FDS concentration resulted in a 6-fold increase compared to 10 mM thiourea. A 10 mM thiourea concentration in the leaching solution that also contained 40 mM Fe(III) resulted in a 30-fold increase in the dissolution rate compared to 40 mM Fe(III) alone. Example 1.3 Column leaching of the different acid-cured copper ores was carried out with the addition of thiourea to the leaching solution. Figure 3 shows a schematic description of the column configuration. The column diameter was 8.84 cm, the column height was 21.6 cm, and the column heap height was 15.9 cm. The irrigation rate was 0.77 mL / min or 8 L / m² / h. Samples of the impregnated leaching solution from these columns were taken every 2 to 3 days to determine the copper content using atomic absorption spectroscopy (AAS). The specific mineralogical composition of these ores is shown in Table 1. The copper content of Ore A, Ore B, and Ore C was 0.52%, 1.03%, and 1.22% w / w, respectively. Before leaching, the ore was acid-cured to neutralize the acid-consuming material present in the ore. Specifically, the ore was mixed with a concentrated sulfuric acid solution composed of 80% concentrated sulfuric acid and 20% deionized water and allowed to stand for 72 hours. For the treatment of Ore C, thiourea was added to the sulfuric acid curing solutions. The initial composition of the leaching solutions included 2.2 g / L of Fe (i.e., 40 mM, provided as ferric sulfate), with a pH of 2 for the control experiment, with or without 0.76 g / L of thiourea (i.e., 10 mM). The initial ore loading on each column was 1.6 to 1.8 kg. The surface solution velocity through the ore column was 7.4 L / m² / h. The pH was adjusted using dilute sulfuric acid. These two columns were maintained in an open-circuit or open-cycle configuration (i.e., without solution recycling) throughout the leaching period. The results of the leaching tests for Ore A, Ore B, and Ore C are shown in Figures 4, 5, and 6, respectively. The presence of thiourea in the leaching agent clearly shows a positive effect on copper leaching from chalcopyrite. On average, the leaching rate in the presence of thiourea increased by a factor of 1.5 to 2.4 compared to the control tests where the leaching solutions did not contain thiourea. Regarding the last time points represented in Figures 4 to 6, the copper extractions for the columns containing Ore A, Ore B, and Ore C, which were leached with a solution containing only sulfuric acid and ferric sulfate, without the addition of thiourea, were 21.2% (after 198 days), 12.4% (after 50 days), and 40.6% (after 322 days), respectively. With the addition of 10 mM thiourea, these extractions were 37.9%, 32.0%, and 72.3%, respectively. With reference to Figure 6, 2 mM thiourea was added to the leaching solution, which initially did not contain thiourea, from day 322 onward, after which the leaching rate increased significantly. From day 322 to day 448, the copper leached from this column increased from 40% to 58%, and rapid leaching was maintained throughout this period. The averages for the last 7 days recorded in Figure 7 indicate that the leaching rate for acid-cured ore C leached in the presence of 10 mM thiourea is 3.3 times higher than that of acid-cured ore C leached in the absence of thiourea, and 4.0 times higher than that of acid-cured and thiourea-cured ore C leached in the absence of thiourea. Figure 8 shows the effect of thiourea on the solution potential. All potentials were recorded relative to the saturated Ag / AgCl reference electrode. The solution potential of the leaching solutions containing thiourea was generally between 75 mV and 100 mV lower than the solution potential of the leaching solution without thiourea. These lower solution potentials correspond to the activity of thiourea in preventing chalcopyrite passivation. Example 1.4 ) CQonn / nznz / Ε / γΐΛΐ Bottle roller leaching Bottle roller leaching tests were carried out in the presence of various thiourea concentrations for Coarse Ore A and Ore B. These tests were carried out using coarse crushed ore (100% larger than 1 / 2 inch). Prior to leaching, the ore was cured using a procedure similar to that employed in the column leaching experiments. The ore was mixed with a concentrated sulfuric acid solution consisting of 80% concentrated sulfuric acid and 20% deionized water and allowed to stand for 72 hours to neutralize the acid-consuming material present in the ore. In several experiments, varying concentrations of thiourea were added to the ore using sulfuric acid curing solutions. The bottles used for the experiments were 20 cm long by 12.5 cm in diameter. Each bottle was loaded with 180 g of cured ore and 420 g of leaching solution, filling it to about one third of its volume. Samples of the leaching solution were taken from each bottle at 2, 4, 6, and 8 hours, and every 24 hours thereafter. The samples were analyzed using atomic absorption spectroscopy (AAS) to determine their copper content. The conditions for the bottle roller experiments are listed in Table 2. Experiments 1 to 6 were performed using only the additional addition of thiourea to the bottles. For experiments 7 to 11, thiourea was added every 24 hours to restore the thiourea concentration. A positive effect of thiourea on copper leaching was observed. In the coarse ore experiments, no stagnation period was observed until between 80 and 120 hours had elapsed. Thiourea was added periodically to the coarse ore experiments, which yielded positive results in copper dissolution. The effect of different concentrations of thiourea in the leaching solution on the leaching of the coarse ore (experiments 1 to 11, as described in Table 2) is shown in Figures 9 and 10. For Ore B, thiourea was added periodically every 24 hours to restore the thiourea concentration in the system, thus better simulating the conditions in the column leaching experiments. As can be seen in Figure 9, 8 mM and 10 mM thiourea resulted in greater copper dissolution than the other thiourea concentrations tested for Ore A. No stagnation period in dissolution was observed until approximately 120 hours, which varied with the thiourea concentration, as shown in Figure 9. Table 1: t CQonn / nznz / E / γΐΛΐ Mineral Ideal formula Mena A Mena B Mena C Actinolite Ca2(Mg,Fe2+)5SÍ8C>22(OH)2 — 1.8 — Biotite K(MgFe2+)jAlSÍ30io(OH)2 — 4.2 — Calcite CaCOs — 19.3 — Chalcopyrite CuFeSi 1.4 3.5 2.6 Clinochlorine (Mg,Fe2+)5Al(SÍ3Al)Oio(OH)s — 15.0 — Diopside CaMgSÍ2Oó — 3.5 — Galena PbS — — 0.1 Gypsum CaSO42H2O — 1.2 — Hematite a-Fe2Os — 0.2 — K-feldspar KAlSi3Os 17.9 10.8 — Kaolinite A12SÍ2O5(OH)4 2.3 — 2.3 Magnetite FesO4 — 0.8 — Molybdenite M0S2 <0.1 — — Muscovite KA12A1Sí3Oio(OH)2 21.9 6.0 41.6 Plagioclase NaAlSÍ3O8-CaAlSÍ2Og 13.6 ,5.4 — Pyrite FeS2 2.3 — 8.0 Quartz SiO2 40.0 8.3 44.4 Rutile TiO2 0.5 — 0.9 Siderite Fe2+CO3 — 0.1 — Tot As can be seen from Figure 9, 5 mM thiourea resulted in higher copper dissolution than other thiourea concentrations tested for Ore B. With respect to Ore A, no dissolution stagnation period was observed until approximately 80 to 120 hours had elapsed, varying with the thiourea concentration as shown in Figure 9. Periodic addition of thiourea resulted in higher copper dissolution and delayed the dissolution stagnation period. Interestingly, solutions containing 100 mM thiourea did not appear to be significantly more effective at extracting copper than those without thiourea, and in fact, performed worse at some time points. This aligns with the results of Deschenes and Ghali, who reported that solutions containing 200 mM thiourea (i.e., 15 g / L) did not improve copper extraction from chalcopyrite. Thiourea is less stable at high concentrations and decomposes. Consequently, it is possible that when initial thiourea concentrations are somewhat higher than 30 mM, sufficient elemental sulfur may be produced by thiourea decomposition to form a film on the chalcopyrite ore and thus contribute to its passivation. It is also possible that at high thiourea concentrations, some copper precipitates from the solution (e.g., see Figure 15), accounting for some of the lower extraction results. Example 2 Extraction from chalcopyrite, covellite, chalcocite, bomite, enargite, pentlandite, violarite and greenockite using thiourea The catalytic effect of thiourea was demonstrated once again with stirred reactor tests. All reactors contained 1.9 L of ferric sulfate solution with a pH of 1.8 and a total iron concentration of 40 mM. One gram of ore sample was used in each reactor test. These experimental conditions were designed to maintain an unlimited supply of oxidant. In order to demonstrate the catalytic effect on chalcopyrite, 100% pure synthetic chalcopyrite was used instead of chalcopyrite concentrate containing various impurities. The chalcopyrite was synthesized via a hydrothermal approach. First, CuCl₂, FeCl₂, and thiourea were mixed in a 1:1:2 molar ratio and dissolved in 150 mL of deionized water. The solution was transferred to a Teflon-lined reaction vessel and heated to 240 °C for 24 hours. At the end of 7CQonn / nznz / E / Y The precipitated powder was washed with acidic water (pH 1) and air-dried at room temperature. X-ray diffraction (XDR) analysis showed that the synthetic chalcopyrite was free of all impurities compared to the chalcopyrite mineral concentrate. This synthetic chalcopyrite was used in all tests carried out in stirred reactors, as disclosed herein. The covellite mineral used in the experiment disclosed herein was also synthesized via a hydrothermal approach. CuCl and thiourea were mixed in a 1:1 molar ratio and dissolved in 150 mL of deionized water. The solution was transferred to a Teilon-coated reaction vessel and heated to 220 °C for 24 hours. The synthesized CuS was acid-washed and air-dried. X-ray diffraction analysis showed it to be 100% pure with no interference from any other species. The chalcocite mineral sample used in the experiments disclosed in the present was a 100% pure natural mineral. The bornite ore used in the experiments reported herein was obtained from Butte, Montana, with a copper content of 58.9% based on inductively coupled plasma atomic emission spectroscopy (ICP-AES). X-ray diffraction analysis showed that the ore contains 76.8% bornite, 8.1% chalcopyrite, 6.3% pyrite, 5.8% tenatite, and 3.0% enargite. The copper content calculated through X-ray diffraction analysis was 55.6%, which is relatively consistent with the chemical assay. The enargite ore used in the experiments disclosed herein was in the form of an enargite concentrate, containing about 70% enargite (34% copper) according to X-ray diffraction analysis. The greenockite mineral used in this experiment was synthesized via hydrothermal focusing. CdChL and thiourea were mixed in a 1:1 molar ratio and dissolved in 100 mL of deionized water. The solution was transferred to a Teflon-lined reaction vessel and heated to 150 °C for 24 hours. The synthesized CdS was acid-washed and air-dried. X-ray diffraction analysis showed it to be 100% pure with no interference from any other species. Table 2: List of bottle roller leaching experiments involving Mena Aya and Mena B 7CQonn / nznz / E / Y Experiment Brief description of experimental conditions No. 1 Coarse Ore A, 0 mM thiourea solution, 40 mM carbolic acid solution, acid cured, no thiourea filler No. 2 Coarse Ore A, 2 mM thiourea solution, 40 mM ferric acid solution, acid cured, no thiourea filler No. 3 Coarse Ore A, 4 mM thiourea solution, 40 mM ferric acid solution, acid cured, no thiourea filler No. 4 Coarse Ore A, 6 mM thiourea solution, 40 mM carbolic acid solution, acid cured, no thiourea filler No. 5 Coarse Ore A, 8 mM thiourea solution, 40 mM carbolic acid solution, acid cured, no thiourea filler No. 6 Coarse Ore A, 10 mM thiourea solution, solution Carbohydrate 40 mM, acid cured, without full filling with thiourea No. 7 Coarse ore B, 0 mM thiourea solution, 40 mM carbohydrate solution, acid cured N.No. 8 Coarse ore B, 1 mM thiourea solution, 40 mM carbolic acid solution, acid cured, periodic addition of thiourea until full fill with thiourea, solution with a concentration of 1 mM. No. 9 Coarse ore B, 5 mM thiourea solution, 40 mM carbolic acid solution, acid cured, periodic addition of thiourea until full fill with thiourea, solution with a concentration of 5 mM. No. 10 Coarse ore B, 10 mM thiourea solution, 40 mM ferric acid solution, acid cured, periodic addition of thiourea until full fill with thiourea, solution with a concentration of 10 mM. No. 11 Coarse ore B, 100 mM thiourea solution, 40 mM carbolic acid solution, acid cured, periodic addition of thiourea until full fill with thiourea, solution with a concentration of 100 mM. CQonn / nznz / E / YiAi The violarite used in the experiments disclosed herein was naturally occurring violarite mineral containing 15.8% Ni according to inductively coupled plasma atomic emission spectroscopy. X-ray diffraction analysis showed that the mineral contained approximately 42% violarite and 13.1% NiSO4·6H2O. 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 electron back-donation from the filled d orbitals of the transition metal. Consequently, without intending to be constrained by theory, it is suspected that the interaction between the surface ion and the thiocarbonyl functional group, especially the back-donation of electrons from the metal 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 most pronounced for minerals with a d10 electronic configuration. Figure 11 shows that thiourea catalyzes the leaching of common copper sulfide minerals, including chalcopyrite, covellite, chalcocite, and bornite, all containing Cu(I). After 96 hours of leaching, the extraction of chalcopyrite reaches 64.1% with 2 mM thiourea compared to 21.1% without thiourea; the extraction of covellite reaches 74.4% with 2 mM thiourea compared to 7.2% without thiourea; the extraction of chalcocite reaches 85.6% with 2 mM thiourea compared to 65.1% without thiourea; and the extraction of bornite reaches 91.4% with 2 mM thiourea compared to 56.7% without thiourea. Like Cu(l), Cd(ll) also has the d10 electronic configuration. Figure 12 shows that the leaching of the CdS ore is significantly enhanced by the addition of thiourea. With thiourea, cadmium extraction reaches 100% at 48 hours, whereas extraction in the uncatalyzed reaction reaches a standstill at 47% after 96 hours. The copper ion in enargite mineral has fewer d electrons than other primary and secondary sulfides, and thus, the catalytic effect might be expected to be slower than that of Cu(I) minerals. However, the results shown in Figure 13 clearly demonstrate that leaching with a leaching solution comprising an initial thiourea concentration of 2 mM increases the copper leaching rate from enargite compared to a control without thiourea, which showed no significant extraction after 96 hours of leaching. Minerals containing transition metal ions with a d7 electronic configuration, such as Ni(III), can also be subjected to leaching catalyzed by the addition of thiourea. Similar to Cu(II), since Ni(III) is the highest stable oxidation state with 7 d electrons, the catalytic effect is not expected to be significant for d10 minerals. With reference to Figure 14, leaching with a leaching solution comprising an initial thiourea concentration of 2 mM increases the leaching rate of nickel from violarite compared to a control without thiourea. The results of the leaching experiments mentioned in Example 2 are summarized in Table 3, where the extraction percentages under non-catalyzed and catalyzed conditions (with an initial thiourea concentration of 2 mM) are compared. i CQonn / nznz / E / YiAi Table 3: Comparisons of reactor leaching for various minerals under non-catalyzed and thiourea-catalyzed conditions 2 mM i CQonn / nznz / E / γΐΛΐ Mineral 96-hour extraction (without thiourea) 96-hour extraction (2 mM thiourea) Chalcopyrite, CuFeSj 21.1% 64.1% Covellite, CuS 6.8% 74.4% Chalcocite, Cu2S 65.1% 85.5% Bomite, CusFeS4 56.7% 91.4% Greenockite, CdS 46.5% 100.0% Enargite, Cu3ASS4 2.1% 10.0% Violarite, FeNi2S4 13.0% 22.2% Example 3 Reagent dosage Optimal reagent dosage can increase leaching efficiency. First, at certain concentrations, the reagent can form an insoluble complex with the metal ion in question and precipitate. For example, thiourea can form an insoluble complex with Cu(I) ions in a 3:1 molar ratio. A precipitation test was conducted to examine the concentration range in which precipitation of the CuTu complex can occur. 20 mL of Cu solution were divided into several identical portions, followed by the addition of various thiourea dosages (i.e., from 0 mM to 60 mM). The solution was stirred for 24 hours, and the remaining Cu in the solution phase was analyzed by atomic absorption spectroscopy. The results are shown in Figure 15, plotted as the percentage of remaining Cu. Secondly, heap leaching of sulfide ores is based on a bioleaching mechanism, and an excessive amount of reagent can be detrimental to the bioleaching microbes. For example, the bacteria most frequently used for bioleaching, such as Acidithiobacillus ferrooxidans and Acidithiobacillus thiooxidans, grow very slowly in a solution containing 10 mM thiourea and cannot survive in a solution with 100 mM thiourea. Third, with regard to thiourea specifically, ferric reacts with thiourea and converts it to FDS (refer to Hydrometallurgy 28, 381-397 (1992)). Although the reaction can be reversed under certain conditions, a high concentration of FDS tends to decompose irreversibly into cyanamide and elemental sulfur (refer to J Chromatogr 368, 444-449). Tu + 2 Fe3+θ FDS + 2 Fe2++ 2 H+ FDS —> Tu + cyanimide + S Therefore, an excess addition of thiourea to the leaching agent can lead to the loss of Fe3+ and thiourea due to oxidation and decomposition. Irreversible decomposition of FDS has been observed when 4 mM thiourea is added to a 40 mM ferric sulfate solution at pH 1.8. To further investigate the effect of thiourea dosage on copper extraction, stirred reactor tests were conducted using 1 g of synthetic chalcopyrite in 1.9 L of 40 mM ferric sulfate solution at pH 1.8 with various initial thiourea concentrations. Treatments were carried out for 172 hours to achieve maximum extraction. The results, shown in Figure 16, indicate that, for 1 g of chalcopyrite, higher thiourea dosages resulted in faster leaching kinetics among the tested thiourea concentrations. For thiourea dosages of 5 mM and lower, the initial 40 mM ferric sulfate solution can be considered a sufficient supply of oxidant. However, for higher dosages such as 10 mM and 20 mM thiourea, more ferric sulfate (in a 1:1 ratio with thiourea) had to be added to the solution to allow for the oxidation of thiourea in the FDS. For 10 mM thiourea, 10 mM Fe3+ was added at time zero. For 20 mM thiourea, 20 mM Fe3+ was added at 72 hours, leading to continued extraction, as shown in Figure 16. The comparison between thiourea dosage and Cu extraction at 172 hours is plotted in Figure 17. An initial thiourea dosage of up to 5 mM appears to have the most pronounced effect on Cu dissolution. As previously mentioned, in prior shake-flask tests with acidic solutions (pH 1.8) containing various concentrations of Fe3+ and Cu2+ ions, a slight precipitate formed upon the addition of 4 mM thiourea due to FDS decomposition. Consequently, thiourea concentrations below 4 mM may prevent this precipitate. A series of shake-flask tests were conducted in solutions containing initial concentrations of 2 mM thiourea and various concentrations of Fe3+ and Cu2+ in a matrix containing Fe3+ (0 mM to 100 mM) and Cu2+ (0 mM to 50 mM) to identify the concentration ranges of [Fe3+] and [Cu2+] that do not result in a Cu complex precipitate.The results showed that no precipitate or loss of Cu from the solution phase occurred using 2 mM thiourea in this wide range of Fe and Cu matrix concentrations. Example 4 Alternative reagents The catalytic effect of various reagents containing a thiocarbonyl functional group on the leaching of synthetic chalcopyrite, covellite, bornite, and enargite was examined. Experiments were conducted in stirred reactors containing a 40 mM ferric sulfate solution at pH 1.8. One g of chalcopyrite or covellite was added to the reactors along with an initial concentration of 2 mM of various thiocarbonyl reagents, including thiourea, thioactamide, SDDC, ETC, and TSCA. Figures 18, 19, 20, and 21 show the copper extraction curves for chalcopyrite, covellite, bornite, and enargite using all or a subset of the aforementioned reagents. From Figures 18 to 21 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 these: chalcopyrite, covellite, bomite and enargite. Figure 22 summarizes the results of the additional stirred reactor tests on chalcopyrite, which were used to investigate carbon disulfide and urea in more detail. These results confirm that, as expected, neither urea nor carbon disulfide are effective reagents. Example 5 Formamidine disulfide (FDS) The catalytic effect of leaching solutions prepared with FDS on the leaching of chalcopyrite, bomite, covellite, and chalcocite was determined in stirred reactor tests. All reactors contained 1.9 L of ferric sulfate solution with a pH of 1.8 and a total iron concentration of 40 mM. One gram of ore sample was used in each reactor test. An initial FDS concentration of 1 mM or an initial thiourea concentration of 2 mM was used. The results of the stirred reactor tests shown in Figures 23a and 23b demonstrate that FDS has comparable effectiveness after 96 hours with thiourea in the leaching of each of these: chalcopyrite, bornite, covellite, and chalcocite. t CQonn / nznz / E / γΐΛΐ Example 6 Closed-cycle gradual bioleaching with thiourea A closed-cycle bioleaching process was carried out using thiourea. 7 kg of ore containing approximately 0.25% Cu, mainly in the form of CuFeSi, were leached at a flow rate of 1 L / day with an aeration rate of approximately 300 ml / min. The ore was pretreated with sulfuric acid to leach the oxides (e.g., chalcanthite and basic copper salts). After the acid leaching period, the residual solutions were collected and replaced with a ferrous sulfate solution containing nutrients (40 mM FeSO4, 0.4 g / L magnesium sulfate heptahydrate, and 0.04 g / L potassium dihydrogen phosphate, with a pH adjusted between 1.6 and 1.8). The ferrous and nutrient solution was discharged through the column to establish a suitable habitat for bacterial growth. Bacterial inoculation showed an increase in oxidation-reduction potential between 274 mV and 550 mV within 48 hours. The solution used in this and subsequent steps was maintained in circulation through the column, forming a self-sustaining, closed-loop system. At this stage, the remaining copper source is primarily CuFeSi. After the bacteria survived in the column, thiourea was progressively added to the leaching solution. As previously analyzed, thiourea is converted to FDS at a molar ratio of 2:1 in the presence of 40 mM Fe3+. The operational potential (ORP) was used as an indicator of bacterial activity, and HPLC was employed to monitor the FDS content. From day 0 to day 50, the leaching solution included 40 mM Fe3+ with inoculated bacteria (without added thiourea). From day 90 to day 98, a total of 1,878 g of thiourea was progressively added, after which HPLC analysis of the effluent showed that the FDS was maintained at approximately 1.5 mM without the addition of thiourea. As shown in Figure 24, the ORP of the effluent was equal to or greater than that of the influent, indicating that the bacteria were actively oxidizing Fe2+ to Fe3+. The FDS content was analyzed by HPLC, which showed that 1.5 mM FDS (equivalent to the added 3 mM thiourea) was present in the solution phase without any precipitate being observed. Therefore, it appears that 1.5 mM FDS (equivalent to 3 mM thiourea) can be used in the solution without ferric precipitate. The results of the closed-loop leaching test are shown in Figure 25. From day 0 to day 50, the bacteria maintained high activity and oxidized Fe2+ to Fe3+. However, with a constant flow rate (1 liter per day), the leaching rate was only 1.97 mg of Cu per day for the first 50 days. The addition of thiourea starting on 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 decompose and remained effective in the closed-loop system. Although specific embodiments of the invention have been described and illustrated, these embodiments should be regarded as illustrative of the invention only and not as limiting the invention as interpreted in accordance with the accompanying claims.

Claims

CLAIMS 1. Use of a reagent having a thiocarbonyl functional group in a concentration in the range of approximately 0.2 mM to approximately 100 mM for extracting at least one base metal ion from a material containing at least one base metal sulfide in an acidic solution comprising at least one oxidizing agent.

2. Use according to claim 1, wherein the reagent is NN' substituted thioureas; 2,5-dithiobiurea; dithiobiuret; thiosemicarbazide; 2-methyl-3-thiosemicarbazide; 4-methyl-3-thiosemicarbazide; 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; isopropylxanthic acid; ethyl thiooxamate; ethyl dithioacetate; pyrazine-2-thiocarboxamide; diethylthiocarbamoyl chloride; diethyldithiocarbamate; tetra-methylthiuram monosulfide; tetramethylthiuram disulfide; pentafluorophenyl chlorothionoformate; 4-fluorophenyl chlorothionoformate; O-phenyl chlorothionoformate; phenyl chlorodithioformate; 3,4-difluorothiobenzamide; 2bromothiobenzamide; 3-bromothiobenzamide; 4-bromothiobenzamide;4-chlorothiobenzamide; 4-fluorothiobenzamide; thiobenzoic acid; thiobenzamide; 4-phenylthiosemicarbazide; O-(p-tolyl) chlorothioformate; 4-bromo-2-methylthiobenzamide; 3-methoxythiobenzamide; 4-methoxythiobenzamide; 4-methylbenzenethioamide; thioacetanilide; salicylaldehyde thiosemicarbazone; indole-3-thiocarboxamide; S-(thiobenzoyl)thioglycolic acid; 3-(acetoxy)thiobenzamide; 4-(acetoxy)thiobenzamide; methyl N'-[(e)-(4-chlorophenyl)methyliden]-hydrazonothiocarbamate; 3-ethoxythiobenzamide; 4-ethylbenzen-1-thiocarboxamide; tert-butyl-3-[(methylsulfonyl)-oxy]1-azethanecarboxylate; diethyldithiocarbamic acid; 2-(phenylcarbonothioylthio)propanoic acid; 2-hydroxybenzaldehyde netylthiosemicarbazone; (1R,4R)-1,7,7-trimethyl-bicyclo[2.2.1]heptan-2-thione; tetraethylthiuram disulfide; tetraethylthiuram disulfide; 4'-hydroxybiphenyl-4-thiocarboxamide; 4-biphenylthioamide; dithizone; 4'-methyl-biphenyl-4-thiocarboxamide; tetraisopropylthiuram disulfide; anthracene-9-thiocarboxamide; phenanthrene-9-thiocarboxamide;sodium dibenzyldithiocarbamate; and 4,4'-bis-(dimethylamino)thiobenzophenone; or any combination thereof.

3. Use according to claim 1 or 2, wherein the thiocarbonyl functional group of the reagent has a sulfur bearing a partial negative charge, has a negative electrostatic potential surface, and has an empty antibonding π* orbital as its lowest unoccupied molecular orbital.

4. The use of a reagent at a concentration in the range of approximately 0.2 mM to approximately 100 mM to extract at least one base metal ion from a material comprising at least one base metal sulfide in an acidic solution comprising at least one oxidizing agent, wherein the reagent is thiourea (Tu), thioacetamide (TA), sodium dimethyldithiocarbamate (SDDC), ethylene trithiocarbonate (ETC), thiosemicarbazide (TSCA), or a combination thereof.

5. The use of a reagent, wherein the reagent is thiourea (Tu), in a concentration in the range of approximately 0.2 mM to approximately 100 mM to extract at least one base metal ion from a material containing at least one base metal sulfide in an acidic solution comprising at least one oxidizing agent.

6. The use of a reagent, wherein the reagent is thioacetamide (TA), to extract at least one base metal ion from a material containing at least one base metal sulfide in an acidic solution comprising at least one oxidizing agent.

7. The use of a reagent, wherein the reagent is sodium dimethyldithiocarbamate (SDDC), for extracting at least one base metal ion from a material containing at least one base metal sulfide in an acidic solution comprising at least one oxidizing agent.

8. The use of a reagent, wherein the reagent is ethylene trithiocarbonate (ETC), to extract at least one base metal ion from a material containing at least one base metal sulfide in an acidic solution comprising at least one oxidizing agent.

9. The use of a reagent, wherein the reagent is thiosemicarbazide (TSCA), to extract at least one base metal ion from a material containing at least one base metal sulfide in an acidic solution comprising at least one oxidizing agent.

10. Use in accordance with any of claims 1 to 9, wherein the concentration of the reagent in the acid solution is sufficient to increase the extraction rate of at least one base metal ion relative to an acid solution not containing the reagent.

11. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 60 mM. t CQonn / nznz / E / YiAi 12. Use in accordance with any of claims 1 to 10, in a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 50 mM.

13. Use in accordance with any of claims 1 to 10, in a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 30 mM.

14. Use in accordance with any of claims 1 to 10, in a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 20 mM.

15. Use in accordance with any of claims 1 to 10, in a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 10 mM.

16. Use in accordance with any of claims 1 to 10, in a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 5 mM.

17. Use in accordance with any of claims 1 to 10, in a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 4 mM.

18. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 3 mM.

19. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 2.5 mM.

20. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 2 mM.

21. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 1.5 mM.

22. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 1.0 mM.

23. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 0.9 mM.

24. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 0.8 mM.

25. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 0.7 mM.

26. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 0.6 mM.

27. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 0.5 mM.

28. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 0.4 mM.

29. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 0.2 mM to approximately 0.3 mM.

30. Use in accordance with any of claims 1 to 10, at a concentration of the reagent in the acid solution of approximately 2 mM.

31. The use of a reagent, wherein the reagent is formamidine disulfide (FDS), to extract at least one base metal ion from a material containing at least one base metal sulfide in an acidic solution comprising at least one oxidizing agent.

32. Use in accordance with claim 31, at an SDS concentration in the acid solution in the range of approximately 0.1 mM to approximately 15 mM.

33. The use in accordance with claim 31, at a concentration of FDS t CQonn / nznz / E / γALA in the acidic solution in the range of approximately 0.1 mM to approximately 10 mM.

34. Use in accordance with claim 31, at an SDS concentration in the acid solution in the range of approximately 0.1 mM to approximately 5 mM.

35. Use in accordance with claim 31, at an SDS concentration in the acid solution in the range of approximately 0.1 mM to approximately 2.5 mM.

36. Use in accordance with claim 31, at an SDS concentration in the acid solution in the range of approximately 0.1 mM to approximately 2 mM.

37. Use in accordance with claim 31, at an SDS concentration in the acid solution in the range of approximately 0.1 mM to approximately 1.5 mM.

38. Use in accordance with claim 31, at an SDS concentration in the acid solution in the range of approximately 0.1 mM to approximately 1 mM.

39. Use in accordance with claim 31, at an SDS concentration in the acid solution in the range of approximately 0.1 mM to approximately 0.5 mM.

40. Use in accordance with claim 31, at an SDS concentration in the acid solution in the range of approximately 0.1 mM to approximately 0.25 mM.

41. Use in accordance with claim 31, at an SDS concentration in the acid solution in the range of approximately 0.1 mM to approximately 0.2 mM.

42. Use in accordance with claim 31, at an SDS concentration in the acidic solution of approximately 0.1 mM.

43. Use in accordance with any of claims 1 to 42, wherein the at least one oxidizing agent comprises a source of ferric ions.

44. Use in accordance with claim 43, wherein the source of ferric ions is Fe3+.

45. Use in accordance with claim 43 or 44, wherein the source of ferric ions comprises ferric sulfate.

46. ​​Use according to claim 43, 44 or 45, wherein the ferric ion source comprises ferric ions generated at least partly by bacteria.

47. Use in accordance with any one of claims 43 to 46, wherein the acid solution comprises a ferric sulfate solution.

48. Use in accordance with any of claims 43 to 46, wherein the acid solution comprises an ferric solution. t CQonn / nznz / E / γΐΛΐ 49. Use in accordance with any of claims 43 to 46, wherein the acid solution comprises ferric media.

50. Use in accordance with any of claims 1 to 49, in combination with a ferric sulfate solution for replacement of the acid solution.

51. Use according to claim 50, wherein the ferric sulfate solution is to provide a source of Fe2+ ions for oxidation to Fe3+ ions by bacteria.

52. The use in accordance with any of claims 1 to 51, wherein the at least one base metal comprises copper, wherein the at least one base metal sulfide comprises at least one copper sulfide.

53. Use according to claim 52, wherein the at least one copper sulfide is chalcopyrite, covellite, bomite, enargite, a copper sulfide of formula CuxSy, wherein the x:y ratio is between 1 and 2, or one of the combinations thereof.

54. Use in accordance with claim 52 or 53, wherein the at least one copper sulfide comprises chalcopyrite.

55. Use in accordance with claim 52, 53 or 54, wherein the at least one copper sulfide comprises covellite.

56. Use in accordance with any of claims 52 to 55, wherein the at least one copper sulfide comprises bornite.

57. Use in accordance with any of claims 52 to 56, wherein the at least one copper sulfide comprises enargite.

58. Use in accordance with any of claims 52 to 57, wherein the at least one copper sulfide comprises at least one copper sulfide of formula CuxSy, wherein the x:y ratio is between 1 and 2.

59. Use according to claim 58, wherein the at least one copper sulfide of formula CuxSy comprises chalcocite.

60. Use in accordance with claim 58 or 59, wherein the at least one copper sulfide of formula CuxSy comprises djurleite.

61. Use in accordance with claim 58, 59 or 60, wherein the at least one copper sulfide of formula CuxSy comprises digenite.

62. Use in accordance with any of claims 1 to 61, wherein the at least one base metal comprises cadmium, wherein the at least one base metal sulfide t CQonn / nznz / E / YiAi comprises a cadmium sulfide.

63. Use according to claim 62, wherein the cadmium sulfide comprises greenockite.

64. Use in accordance with any of claims 1 to 63, wherein the at least one base metal comprises nickel, wherein the at least one base metal sulfide includes at least one nickel sulfide.

65. Use in accordance with claim 64, wherein the at least one nickel sulfide comprises pentlandite.

66. Use in accordance with claim 64 or 65, wherein the at least one nickel sulfide comprises violarite.

67. Use in accordance with any of claims 1 to 66, wherein the material is an ore.

68. Use in accordance with any of claims 1 to 66, wherein the material is a concentrate of at least one base metal sulfide.

69. Use in accordance with any of claims 1 to 68, wherein the operating potential of the acid solution is maintained above 500 mV vs Ag / AgCl.

70. Use in accordance with any of claims 1 to 69, wherein the reagent does not form complexes / precipitate with at least one base metal ion.

71. Use in accordance with any of claims 1 to 70, at a temperature between 0 °C and 80 °C.

72. A method for recovering at least one base metal ion from a material containing at least one base metal sulfide, the method comprising: contacting the material with an acidic solution comprising at least one oxidizing agent and a reagent having a thiocarbonyl functional group to extract at least one base metal ion from the at least one base metal sulfide, wherein the concentration of the reagent is in the range of approximately 0.2 mM to approximately 100 mM, to produce an impregnated solution containing the at least one base metal ion; and recovering the at least one base metal ion from the impregnated solution.

73. The method according to claim 72, wherein the reagent is NN' substituted thioureas; 2,5-dithiobiurea; dithiobiuret; thiosemicarbazide; thiosemicarbazide; thioacetamide; 2-methyl-3-thiosemicarbazide; 4-methyl-3-thiosemicarbazide; 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-isopropylxanthic acid; ethyl thiooxamate; ethyl dithioacetate; pyrazine-2thiocarboxamide; diethylthiocarbamoyl chloride; diethyldithiocarbamate; tetramethylthiuram monosulfide; tetramethylthiuram disulfide; pentafluorophenyl chlorothionoformate; 4-fluorophenyl chlorothionoformate; O-phenyl chlorothionoformate; phenyl chlorodithioformate; 3,4-difluorothiobenzamide; 2-bromothiobenzamide; 3-bromothiobenzamide; 4-bromothiobenzamide;4-chlorothiobenzamide; 4-ylluo-rothiobenzamide; thiobenzoic acid; thiobenzamide; 4-phenylthiosemicarbazide; O-(p-tolyl)chlorothionienoformate; 4-bromo-2-methylthiobenzamide; 3-methoxythiobenzamide; 4-methoxythiobenzamide; 4-methylbenzenethioamide; thioacetanilide; salicylaldehyde thiosemicarbazone; indole-3-thiocarboxamide; S-(thiobenzoyl)thioglycolic acid; 3-(acetoxy)thiobenzamide; 4-(acetoxy)thiobenzamide; methyl N'-[(e)-(4-chlorophenyl)methyliden]-hydrazonothiocarbamate; 3-ethoxythiobenzamide; 4-ethylbenzen-1-thiocarboxamide; tert-butyl-3-[(methylsulfonyl)-oxy]1-azethanecarboxylate; diethyldithiocarbamic acid; 2-(phenylcarbonothio-ylthio)propanoic acid; 2-hydroxybenzaldehyde netylthiosemicarbazone; (1R,4R)-1,7,7-trimethyl-bicyclo[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 any combination thereof.; 74. The method according to claim 72 or 73, wherein the thiocarbonyl functional group of the reagent has a sulfur bearing a partial negative charge, has a negative electrostatic potential surface, and has an empty antibonding π* orbital as its lowest unoccupied molecular orbital.

75. A method for recovering at least one base metal ion from a material containing at least one base metal sulfide, the method comprising: contacting the material with an acidic solution comprising at least one oxidizing agent and a reagent having a thiocarbonyl functional group to extract at least one base metal ion from the at least one base metal sulfide, wherein the concentration of the reagent is in the range of approximately 0.2 mM to approximately 100 mM, to produce an impregnated solution containing the at least one base metal ion; and recovering the at least one base metal ion from the impregnated solution; wherein the reagent is thiourea (Tu), thioacetamide (TA), sodium dimethyldithiocarbamate (SDDC), ethylene trithiocarbonate (ETC), thiosemicarbazide (TSCA), or a combination thereof.

76. A method for recovering at least one base metal ion from a material containing at least one base metal sulfide, the method comprising: contacting the material with an acidic solution comprising at least one oxidizing agent and a reagent for extracting at least one base metal ion from the at least one base metal sulfide, wherein the reagent is thiourea (Tu), and wherein the concentration of the reagent is in the range of approximately 0.2 mM to approximately 100 mM, to produce an impregnated solution containing the at least one base metal ion; and recovering the at least one base metal ion from the impregnated solution.

77. A method for recovering at least one base metal ion from a material containing at least one base metal sulfide, the method comprising: contacting the material with an acidic solution comprising at least one oxidizing agent and a reagent having a thiocarbonyl functional group to extract at least one base metal ion from the at least one base metal sulfide, wherein the reagent is thioacetamide (TA), to produce an impregnated solution containing the at least one base metal ion; and recovering the at least one base metal ion from the impregnated solution.

78. A method for recovering at least one base metal ion from a material containing at least one base metal sulfide, the method comprising: contacting the material with an acidic solution comprising at least one oxidizing agent and a reagent having a thiocarbonyl functional group to extract at least one base metal ion from the at least one base metal sulfide, wherein the reagent is sodium dimethyldithiocarbamate (SDDC), to produce an impregnated solution containing the at least one base metal ion; and recovering the at least one base metal ion from the impregnated solution.

79. A method for recovering at least one base metal ion from a material containing at least one base metal sulfide, the method comprising: contacting the material with an acidic solution comprising at least one oxidizing agent and a reagent having a thiocarbonyl functional group, wherein the reagent is ethylene trithiocarbonate (ETC), to produce an impregnated solution containing the at least one base metal ion; and recovering the at least one base metal ion from the impregnated solution.

80. A method for recovering at least one base metal ion from a material containing at least one base metal sulfide, the method comprising: contacting the material with an acidic solution comprising at least one oxidizing agent and a reagent having a thiocarbonyl functional group to extract at least one base metal ion from the at least one base metal sulfide, wherein the reagent is thiosemicarbazide (TSCA), to produce an impregnated solution containing the at least one base metal ion; and recovering the at least one base metal ion from the impregnated solution.

81. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is sufficient to increase the extraction rate of at least one base metal ion compared to an acid solution that does not contain the reagent.

82. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 60 mM.

83. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 50 mM.

84. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 30 mM.

85. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 20 mM.

86. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 10 mM.

87. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 5 mM.

88. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 4 mM.

89. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 3 mM.

90. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 2 mM.

91. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 1.5 mM.

92. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 1.0 mM.

93. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 0.9 mM.

94. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 0.8 mM.

95. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 0.7 mM.

96. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 0.6 mM.

97. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 0.5 mM.

98. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 0.4 mM.

99. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 0.3 mM.

100. The method according to any of claims 72 to 80, wherein the concentration of the reagent in the acid solution is in the range of approximately 0.2 mM to approximately 2 mM.

101. A method for recovering at least one base metal ion from a material containing at least one base metal sulfide, the method comprising: contacting the material with an acidic solution comprising at least one oxidizing agent and a reagent for extracting at least one base metal ion from the at least one base metal sulfide, wherein the reagent is formamidine disulfide (FDS), to produce an impregnated solution containing the at least one base metal ion; and recovering at least one base metal ion from the impregnated solution.

102. The method according to claim 101, wherein the concentration of SDS in the acid solution is in the range of approximately 0.1 mM to approximately 15 mM.

103. The method according to claim 101, wherein the concentration of SDS in the acid solution is in the range of approximately 0.1 mM to approximately 10 mM.

104. The method according to claim 101, wherein the concentration of SDS in the acid solution is in the range of approximately 0.2 mM to approximately 5 mM.

105. The method according to claim 101, wherein the concentration of SDS in the acid solution is in the range of approximately 0.1 mM to approximately 2.5 mM.

106. The method according to claim 101, wherein the concentration of SDS in the acid solution is in the range of approximately 0.1 mM to approximately 2 mM.

107. The method according to claim 101, wherein the concentration of SDS in the acid solution is in the range of approximately 0.1 mM to approximately 1.5 mM.

108. The method according to claim 101, wherein the concentration of SDS in the acid solution is in the range of approximately 0.1 mM to approximately 1.0 mM.

109. The method according to claim 101, wherein the concentration of SDS in the acid solution is in the range of approximately 0.1 mM to approximately 0.5 mM.

110. The method according to claim 101, wherein the concentration of SDS in the acid solution is in the range of approximately 0.1 mM to approximately 0.25 mM.

111. The method according to any one of claims 101 to 110, wherein the concentration of FDS in the acid solution is sufficient to provide enough thiourea to increase the extraction rate of the base metal ion relative to an acid solution that does not contain FDS.

112. The method according to any of claims 72 to 111, wherein the at least one oxidizing agent comprises a source of ferric ions.

113. The method according to claim 112, wherein the source of ferric ions is Fe3+.

114. The method according to claim 112 or 113, wherein the source of ferric ions comprises ferric sulfate.

115. The method according to claim 112, 113 or 114, wherein the ferric ion source comprises ferric ions generated at least partly by bacteria.

116. The method according to any of claims 112 to 115, wherein the acid solution comprises a ferric sulfate solution.

117. The method according to any of claims 112 to 115, wherein the acid solution comprises a ferric solution.

118. The method according to any of claims 112 to 115, wherein the acid solution comprises ferric media.

119. The method according to any of claims 72 to 118, wherein the acid solution is replaced with a ferric sulfate solution.

120. The method according to claim 119, wherein the ferric sulfate solution provides a source of Fe2+ ions that are actively oxidized to Fe3+ ions by bacteria.

121. The method according to any of claims 72 to 120, wherein the at least one metal-based sulfide comprises at least one copper sulfide.

122. The method according to claim 121, wherein the at least one copper sulfide is chalcopyrite, covellite, bornite, enargite, a copper sulfide of formula CuxSy, wherein the x:y ratio is between 1 and 2, or one of the combinations thereof.

123. The method according to claim 121 or 122, wherein the at least one copper sulfide comprises chalcopyrite.

124. The method according to claim 121, 122 or 123, wherein the at least one copper sulfide comprises covellite.

125. The method according to any of claims 121 to 124, wherein the at least one copper sulfide comprises bomite.

126. The method according to any of claims 121 to 125, wherein the at least one copper sulfide comprises enargite.

127. The method according to any of claims 121 to 126, wherein the at least one copper sulfide comprises a copper sulfide of formula CuxSy, wherein the x:y ratio is between 1 and 2.

128. The method according to claim 127, wherein the at least one copper sulfide of formula CuxSy comprises chalcocite.

129. The method according to claim 127 or 128, wherein the at least one copper sulfide of formula CuxSy comprises djurleite.

130. The method according to claim 127, 128 or 129, wherein the at least one copper sulfide of formula CuxSy comprises digenite.

131. The method according to any of claims 72 to 130, wherein the at least one base metal sulfide comprises a cadmium sulfide.

132. The method according to claim 131, wherein the cadmium sulfide is greenockite.

133. The method according to any of claims 72 to 132, t CQonn / nznz / E / γΐΛΐ wherein the at least one base metal sulfide comprises at least one nickel sulfide.

134. The method according to claim 133, wherein the at least one nickel sulfide comprises pentlandite.

135. The method according to claim 133 or 134, wherein the at least one nickel sulfide comprises violarite.

136. The method according to any of claims 72 to 135, wherein the material is an ore.

137. The method according to any of claims 72 to 135, wherein the material is a concentrate of at least one base metal sulfide.

138. The method in accordance with any of claims 72 to 137, wherein the method is a leaching.

139. The method according to claim 138, wherein the leaching is a percolation leaching.

140. The method according to claim 138, wherein the leaching is a heap leaching.

141. The method according to claim 138, wherein the leaching is a vat leaching.

142. The method according to claim 138, wherein the leaching is a tank leaching.

143. The method according to claim 138, wherein the leaching is a column leaching.

144. The method according to any of claims 72 to 143, wherein the recovery of at least one base metal ion from the impregnated solution comprises solvent extraction and electrowinning.

145. The method according to any of claims 72 to 144, further comprising maintaining the operating potential of the acid solution above 500 mV vs Ag / AgCl.

146. The method according to any of claims 72 to 145, wherein the reagent does not form complexes / precipitate with at least one base metal ion.

147. The method according to any of claims 72 to 146, wherein the method is performed at a temperature between 0 °C and 80 °C.