Process for leaching valuable elements from metallurgical residues

Abstract

A process for leaching copper and lead from a metallurgical residue of smelted powder which has undergone a copper leaching process and comprises copper, iron, lead, silicon and optionally arsenic, antimony and bismuth, comprising: (i) leaching copper from the metallurgical residue with a first acid solution so as to obtain a first leach solution rich in copper and iron and optionally arsenic, antimony and bismuth and a first leached sludge having a reduced copper and iron content and optionally a reduced arsenic content and being rich in lead and silicon, (ii) leaching the first leached sludge with a first solution of carboxylate salt so as to obtain a second leached sludge lacking lead and a second leach solution rich in lead, (iii) precipitation, wherein a first base is added to the second leach solution rich in lead so as to obtain a first lead concentrate and a first precipitated solution lacking lead, (iv) alkaline leaching of the second leached sludge, wherein a second base is added so as to form an alkaline leaching solution so as to obtain a third leached sludge having a reduced silicon content and a third leach solution rich in silicon and optionally arsenic, (v) hydrochloric acid leaching of the third leached sludge, wherein an acid solution is used in a chloride environment so as to obtain a fourth leached sludge for final disposal and a fourth leach solution rich in copper, lead and iron and optionally arsenic, (vi) (vii) precipitating metals from the fourth leach solution rich in copper, lead and iron and optionally arsenic with a neutralizing slurry so as to produce a fifth solution rich in chlorides and a first precipitated solid rich in iron, copper and lead and optionally arsenic, and (viii) leaching the first precipitated solid rich in iron, copper and lead and optionally arsenic with a sulphuric acid solution so as to produce a sixth leach solution rich in copper, iron and optionally arsenic and a second lead concentrate.

Description

Field of the Invention

[0001] The present invention relates to a process for leaching valuable elements from metallurgical residues, with particular focus on the recovery of elements such as copper and lead, and the process may optionally contemplate the leaching of elements such as iron, arsenic, antimony, bismuth, silver and germanium.

[0002] In a more particular aspect, the present invention relates to a method for leaching valuable elements from metallurgical residues for producing a final residue that is a stable residue according to TCLP (Total Characteristic Leaching Method) and SPLP (Synthetic Precipitation Leaching Method) hazard analysis.

[0003] In another variant of the invention, a process is disclosed for obtaining a lead concentrate from metallurgical residues, which in an even more preferred aspect corresponds to lead carbonate.

[0004] In a more specific aspect, metallurgical residues are powders from a metal smelting process.

[0005] In an even more specific aspect, the metallurgical residue is a powder from a copper smelting process.

[0006] In an even more specific aspect, metallurgical residues or in particular smelter fines contemplate materials that have been subjected to a leaching process, such as sulfuric acid leaching.

[0007] For the purposes of this specification any metallurgical residue that has undergone a previous leaching process shall be considered to be sludge. Existing technology

[0008] Copper leaching

[0009] The copper in the sludge primarily consists of substances such as ferrite and / or spinel in the form of CuFeO, ZnFeO, and a corresponding portion of iron in the form of FeFeO. Leaching of these substances is dependent on temperature, acid concentration, and residence time, as described in a study by BS Boyanov et al., World Academy of Science, Engineering and Technology, Vol. 9, 2015, 1592-1598, which investigated the leaching of synthetic ferrites of zinc, copper, and cadmium, evaluating the aforementioned variables. The results of this study showed that ferrite dissolves well in HCl and HSO at elevated temperatures and high acid concentrations.

[0010] At high acid concentrations, an asymptotic behavior of copper leaching with respect to leaching temperature was observed, reaching copper leaching yields greater than 90% in the temperature range between 85 and 90°C once 60 minutes of reaction time had passed in sulfuric acid medium.

[0011] Lead leaching and precipitation

[0012] In 2011, global lead consumption exceeded 10 million tons, of which approximately 80% was destined for the manufacture of lead-acid batteries. These batteries contain a certain amount of lead in the form of Pb, PbO₂, and PbSO₄. The most traditional method for recovering lead is the pyrometallurgical route, characterized by the addition of reducing agents such as carbon powder, iron filings, and sodium oxalate. This process requires high energy requirements due to the use of ovens operating at temperatures exceeding 1000°C (He et al., Minerals 7, no. 6 (2017): 93).

[0013] On the other hand, hydrometallurgical approaches to lead recovery allow operation at reduced temperatures, reducing energy consumption and, in turn, eliminating the production of sulfur dioxide, a gas harmful to the environment. Hydrometallurgical approaches use desulfurizers such as sodium carbonate, ammonium carbonate, sodium bicarbonate, ammonium bicarbonate, sodium hydroxide, sodium citrate, acetic acid, sodium acetate, etc. The purpose of these processes is to exchange other anions with sulfate ions to form insoluble salts. Once recovered, lead salts such as lead citrate can be calcined to produce lead oxide (Zárate-Gutiérrez y Lapidus, Hydrometallurgy 144 (2014): 124-128).

[0014] Desulfurization with citrate

[0015] In the particular case of using citrate, a mixture of citric acid and sodium citrate facilitates the leaching of lead sulfate and the subsequent crystallization of lead citrate.

[0016] Leaching of lead in citrate solution

[0017] The solubility product constant of lead sulfate at 20℃ is 6.31·10 -7 , indicating that the solubility of PbSO4 is quite low. However, in the presence of concentrated citrate solution, lead forms a series of soluble complexes. 2+ In the solution, various citrate complex species are present in the solution in the pH range of 4.6 to 11.5. At a pH below 4.6, lead sulfate is predominant, while at a pH above 11.5, lead hydroxide is predominant.

[0018] He et al., Minerals 7, No. 6 (2017): 93, studied lead leaching from a paste having a lead sulfate to water weight ratio of 1:10 by adding 650 g / L sodium citrate at 35°C. These conditions allowed for greater than 99% conversion of lead sulfate to lead citrate after a 60-minute reaction time. Increasing the temperature up to 95°C allowed for efficiencies approaching 99% at a sodium citrate concentration of 300 g / L after a 60-minute reaction time. However, a decrease in lead citrate production was observed when citric acid was introduced into the mixture. The optimal pH for lead citrate production was in the range of 6 to 7. Increased lead leaching efficiency was also achieved from lead-acid batteries using citric acid and ammonium reagents at a pH of 5.5. In the pH range of 5.2 to 5.5, the presence of lead citrate trihydrate ([Pb3(C6H5O7)2]·[3H2O]) was reported as the main species. At higher pH values ​​in the range of 8 to 10, lead recovery as citrate is lower due to the formation of lead hydroxide. When the lead residue is rich in oxides such as PbO and PbO2, leaching with citric acid in a molar ratio of 1:1 and 4:1 to lead (II) oxide and lead (IV) oxide at 20°C for between 15 and 60 minutes achieves leaching efficiencies greater than 99% by weight, obtaining Pb(C6H6O7)·H2O as the main species (Sonmez and Kumar, Hydrometallurgy 95, No. 1-2 (2009), 82-86).

[0019] Slurry density is another important parameter for lead leaching with citrate solutions. Within a 10 to 50 g / L lead alum slurry range, leaching with a 1 M sodium citrate solution, pH 7, at 600 rpm and 25°C achieved high lead extraction levels of 90 to 94% at a slurry concentration of 10 g / L. Higher slurry concentrations resulted in lower lead extraction.

[0020] Therefore, the hydrometallurgical desulfurization process is affected by the diffusion of citrate ions in the lead paste within the reactor due to the increased density of the lead paste. In this context, it is crucial to design a reactor that maximizes mass transfer in the system.

[0021] Cambridge Enterprise Limited has developed a technology for recovering lead from lead waste using citric acid (WO2008056125A1). This technology essentially involves treating a lead residue containing lead (II) oxide, lead (IV) oxide, and lead sulfate with a citric acid solution, which can alternatively be combined with sodium citrate at a pH ranging from 1.4 to 6. Finally, hydrogen peroxide can be added as a reducing agent in an alkaline environment to promote the lead (IV) oxide leaching reaction, thereby producing lead citrate (Sonmez and Kumar, Hydrometallurgy 95, Nos. 1-2 (2009), 82-86).

[0022] The present invention differs from patent WO2008056125 A1, in which the pH required for leaching varies from 5.33 to 8.8, preferably using a pH equal to 7. In addition, the present invention proposes recycling the citrate solution obtained after the precipitation step with sodium carbonate, in order to leach again the output metallurgical residue from the sulfuric acid leaching step.

[0023] Lead citrate crystals

[0024] On the other hand, when crystallizing lead citrate, both temperature variations and changes in the molar ratio between citric acid and sodium citrate are effective alternatives. When the molar ratio of citric acid to sodium citrate is zero at temperatures greater than 75°C, a smaller amount of lead is observed in the filtrate, indicating a greater precipitation of lead citrate. Conversely, at initial temperatures less than 75°C, a larger amount of lead is observed in the filtrate, indicating a smaller fraction of lead crystallized. Indeed, elevated temperature differences greater than 75°C favor lead citrate crystallization (US Pat. No. 8,323,373).

[0025] Another strategy for improving the crystallization of lead citrate is to increase the citric acid to sodium citrate molar ratio. For example, at an initial crystallization temperature of 35°C and a citric acid to sodium citrate molar ratio of 0.92, only 0.42% of the lead was present in the filtrate, while greater than 99% of the remainder crystallized. Using a citric acid to sodium citrate molar ratio of 1.7:1 at 20°C, an acid to lead battery paste ratio of 1:5, and a reaction time of 8 hours allowed the conversion of 98% of the lead in the lead citrate. For the purposes of reducing costs and improving efficiency in the leaching step, the use of citric acid for dissolving the lead paste has been suggested. Thus, cooling the solution along with increasing the citric acid to sodium citrate ratio (acidification) has been observed to be effective in improving crystallization efficiency and achieving lead citrate recoveries greater than 99% ( US8323373 ).

[0026] Sodium carbonate precipitation

[0027] Most precipitation methods using sodium carbonate reported in the literature are based on lead battery processing.

[0028] Patent application AU2009350377A1 discloses a method for recovering lead contained in lead battery electrolytic paste by dissolving lead oxide in H2SO4 in the presence of acetate to obtain lead sulfate, which is soluble in this salt and subsequently used to precipitate lead carbonate / basic lead carbonate, lead oxide, or lead hydroxide, after adding a carbonate or hydroxide of the same cation. This method requires recycling the acetate-containing solution obtained from the lead precipitation step to dissolve new electrolytic paste. However, precipitation with sodium hydroxide is performed at 83°C, with no specific details on how precipitation with sodium carbonate is performed. This application differs from the present invention in that citrate is used for lead leaching, and since a sulfate removal step is not required, the present invention contemplates a purge that allows the sulfate concentration to be maintained below the saturation limit.

[0029] US Patent No. 8,568,670 discloses a method for producing lead carbonate from slag obtained from a bismuth refining process, wherein the slag is leached with sodium chloride to obtain lead chloride, which is filtered and neutralized and then added to a solution containing ammonium bicarbonate at 2 to 3 times the stoichiometric amount required for the reaction, and the lead carbonate is precipitated by adjusting the pH to between 8 and 11 for a period of 1 to 2 hours. The present invention differs from this application in that no chloride salt is used for lead leaching. Instead, sodium carbonate is used for precipitation at a pH between 7 and 8, and the sodium carbonate is added directly to the lead leach solution without requiring a prior lead chloride precipitation step.

[0030] Patent US Pat. No. 5,545,805 claims a method for fixing lead in materials containing elements that contribute to hardness, such as calcium and magnesium, wherein the lead-containing material is brought into contact with a carbonate salt, which is an alkali metal, in an amount sufficient for the metal that contributes to hardness to react with the carbonate salt, and polyacid oxyanions are added. The carbonate salt is essentially sodium carbonate or potassium carbonate, while the polyacid oxyanions are selected from phosphates, borates, selenates, arsenates, chromates or sulfates, and are used to precipitate lead as an oxysalt of the oxyanions. In particular, the method also describes that the oxysalt may be a carbonate, in which case lead carbonate will precipitate. The precipitation pH observed in the application example increased by approximately 12.3. The patent application differs from the present invention in that polyacid oxyanions are not required for lead leaching.

[0031] Patent application WO2005007904A1 discloses a method for desulfurizing a solid mixture resulting from the disintegration of a lead battery, containing lead oxide, oxysulfates, and sulfate residues, by contacting the mixture with ammonium carbonate, sodium carbonate, or potassium carbonate at a molar ratio exceeding the sulfate concentration by between 0.1 and 10%, and with a dissolved substance of yellow lead alum, citric acid and citrates being mentioned, with a carbonate to solvent molar ratio of between 1 and 2.75. In this method, the lead residue is desulfurized by the action of the dissolved substances mentioned above, and then lead is precipitated by the action of the carbonates, wherein the method is preferably carried out at a working temperature between 60 and 100°C. This application differs from the present invention in that the amount of lead carbonate added need not exceed the stoichiometric amount required for lead precipitation based on the sulfate content of the solution; the sulfate level is maintained below saturation for each removal considered, which does not affect the lead carbonate precipitation.

[0032] Alkaline leaching

[0033] Mufakhir et al., IOP Conf. Series: Materials Science and Engineering, 285 (2017) 012003, studied the leaching of silicon from slags from a process for obtaining nickel-iron alloys in the presence of sodium hydroxide. Alkaline leaching was performed when the slags were evaluated at NaOH concentrations between 6 and 14 mol / L, solids contents between 5 and 25% w / w, and temperatures between 25 and 110°C. The results showed a maximum yield of 31% for silicon leaching. In the subject matter of the present application, a method for maximizing copper and lead leaching is disclosed, comprising sulfuric acid and citric acid leaching steps, the purpose of which is to remove the Cu and Pb present in the sludge for subsequent alkaline leaching. Removing Cu, Fe, and Pb in an early step allows chemical modification of the sludge, leaving behind silicon species that are more vulnerable to leaching, as shown by the results obtained in the present application.

[0034] hydrochloric acid leaching

[0035] US Patent No. 7,329,396 describes a method for leaching valuable metals from oxidized materials, such as laterite nickel ore, comprising the steps of leaching the mineral with a leaching agent comprising a cationic salt (e.g., magnesium chloride) and HCl. Additional oxidizing agent or metal chloride may be added (as a result of the leaching operation). In one embodiment, the method comprises recovering valuable metals from the mineral, comprising the steps of: leaching the mineral with a leaching agent; separating a leachate rich in valuable mineral metals in a first solid-liquid separation; oxidizing and neutralizing the leachate rich in valuable metals thus obtained; and separating a magnesium chloride solution from the leachate thus obtained in a second solid-liquid separation. In another embodiment, the leaching agent solution is regenerated from the magnesium chloride solution. In an additional embodiment, the regeneration of the leachate comprises the step of producing magnesium oxide from the magnesium chloride solution.

[0036] The difference between the present invention and US Pat. No. 7,329,396 is that it states that a pH greater than 0.4 is preferred for precipitation of hematite. In the present invention, it is convenient to work at a low pH, preferably less than pH -0.25, in order to obtain iron ions in solution, which facilitates the use of the leach solution in other leaching processes, such as those for smelting powder leaching. Furthermore, precipitation of iron hydroxide is completely disadvantageous in the present invention, as the silver to iron concentration ratio is equivalent to 0.01 g Ag / g Fe, and as a result, the precipitation of iron hydroxide can hinder the presence of silver in solution.

[0037] Patent application CA2820631A1 relates to a method for efficiently processing several materials containing many different metals. These materials can be leached with HCl to obtain a leachate and a solid. These can then be separated from each other, and the first leachate metal can be separated. The second metal can then be separated from the leachate. The first and second metals can be substantially separated from each other and from the leachate. This can be achieved by controlling the leachate temperature, adjusting the pH, reacting even more of the leachate with HCl, etc. The metals that can be recovered as metal chlorides can ultimately be converted to the corresponding metal oxides, thereby allowing for the recovery of the HCl. The several metals can be selected from aluminum, iron, zinc, copper, gold, silver, molybdenum, cobalt, magnesium, lithium, manganese, nickel, palladium, platinum, thorium, phosphorus, uranium, titanium, rare earth elements, and rare metal elements.

[0038] The present invention differs from patent application CA2820631A1 in that it does not require temperatures greater than 90° C. for efficient metal leaching, unlike that application which required temperatures greater than 125° C. Furthermore, leaching of the aluminum-containing material was performed with HCl concentrations starting at 18% w / w, whereas the present invention requires HCl concentrations below 140 g / L (or below 11% w / w).

[0039] Description of the drawings

[0040] FIG1 shows a flow chart of the method disclosed in the present invention.

[0041] FIG. II shows an X-ray diffraction spectrum of the lead concentrate obtained in the application example.

[0042] FIG. III shows the Raman spectrum of the lead concentrate obtained in the application example.

[0043] Description of the Invention

[0044] In its broad aspects, the present invention describes a method for leaching copper and lead from a metallurgical residue which has undergone a copper leaching process and comprises smelter powder comprising copper, iron, lead and silicon and optionally arsenic, antimony and bismuth, and maximizing the recovery of copper and lead.

[0045] In another preferred embodiment, the present invention describes a method for leaching copper and lead from the metallurgical residue of smelting powder which has been subjected to a copper leaching process and comprises copper, iron, lead and silicon and optionally arsenic, antimony, silver, germanium leaving a final residue comprising mainly aluminosilicates and has been subjected to a risk analysis according to a TCLP analysis.

[0046] In another preferred embodiment, the present invention describes a method for obtaining a lead concentrate from the metallurgical residues of smelting powder which have been subjected to a copper leaching process and contain copper, iron, lead and silicon and optionally arsenic, antimony, bismuth, silver, and germanium.

[0047] The method of the present invention comprises the following steps:

[0048] Copper leaching step (i) of metallurgical residues (1), wherein a first acid leaching solution (2) is used in order to obtain a first leaching solution (3) rich in copper and iron and optionally arsenic, antimony and bismuth and a first leached sludge (4) having a reduced copper and iron content and optionally a reduced arsenic content and rich in lead and silicon,

[0049] a step (ii) of leaching a first leached sludge (4), wherein said first leached sludge (4) is treated with a first solution (5) of a carboxylate salt so as to obtain a second leached sludge (6) deficient in lead and a second leach solution (7) enriched in lead,

[0050] a precipitation step (iii) in which a first base (8) is added to the second leaching solution (7) enriched in lead so as to obtain a first lead concentrate (9) and a first precipitation solution (10) deficient in lead,

[0051] a step (iv) of alkaline leaching of the second leachable sludge (6), wherein a second base (11) is added so as to form an alkaline leach solution so as to obtain a third leachable sludge (12) having a reduced silicon content and a third leach solution (13) enriched in silicon and optionally arsenic,

[0052] a hydrochloric acid leaching step (v) of the third leached sludge (12) using an acid solution in a chloride environment (14) in order to obtain a fourth leached sludge (15) for final disposal and a fourth leaching solution (16) rich in copper, lead and iron and optionally arsenic,

[0053] a step (vi) of precipitating metals from a fourth leach solution (16) rich in copper, lead and iron and optionally arsenic using a neutralized slurry (17) to produce a fifth solution (18) rich in chloride and a first precipitated solid (19) rich in iron, copper and lead and optionally arsenic, and

[0054] step (vii) of leaching the first precipitated solid (19) rich in iron, copper and lead and optionally arsenic with a sulphuric acid solution (20) to produce a sixth leach solution (21) rich in copper, iron and optionally arsenic and a second lead concentrate (22).

[0055] In a preferred embodiment, the metallurgical residue to be treated is a powder obtained by a metal smelting process.

[0056] In a more preferred embodiment, the powder is a powder obtained by a copper smelting process.

[0057] In an even more preferred embodiment, the metallurgical residue has been subjected to a copper leaching process.

[0058] In an even more preferred embodiment, the metallurgical residue has been subjected to leaching with H2SO4.

[0059] In a preferred embodiment, the metallurgical residue to be processed contains the mineral substances lead vanadium, covellite, ferrocodite in the form of CuOFe2O3, galachite in the form of ZnOFe2O3, magnetite, iron(III) oxide, pyrite (pirita), scorodite, mucovita, kaolinite and lead(II) sulfate.

[0060] In an even more preferred embodiment, the copper contained in the metallurgical residue is present as copper sulfate, chalcocite, calcosina, and copper-iron spinel in the form of CuOFe2O3.

[0061] In an even more preferred embodiment, at least 50% of the copper contained in the metallurgical residue is in the form of copper-iron spinel in the form of CuOFe2O3.

[0062] In a preferred embodiment, the silicon contained in the metallurgical residue is present as muscovite and kaolinite.

[0063] In a further preferred embodiment, the lead contained in the metallurgical residue is present as lead(II) sulfate, galena or lead(II) oxide.

[0064] In an even more preferred embodiment, the lead is at least 95% as lead (II) sulfate.

[0065] In a preferred embodiment, the first H 2 SO 4 solution may comprise H 2 SO 4 and / or refinery wastewater.

[0066] In a preferred embodiment, step (i) is performed at a H2SO4 concentration between 150 and 300 g / L, more preferably at a H2SO4 concentration of 250 g / L.

[0067] In a preferred embodiment, step (i) is carried out at a temperature between 50 and 130°C, more preferably at a temperature of 85°C.

[0068] In a preferred embodiment, step (i) is performed for a period of between 3 and 12 hours, more preferably for a residence time of 6 hours.

[0069] In a preferred embodiment, step (i) is performed at a solids concentration of between 5 and 20% w / w, more preferably at a solids concentration of 15% w / w.

[0070] In a preferred embodiment, in step (ii) of leaching, the carboxylate salt is sodium citrate.

[0071] In a preferred embodiment, the sodium citrate solution in step (ii) has a molar concentration of sodium citrate between 0.5 and 1 M.

[0072] In a preferred embodiment, in step (ii), the first leached sludge is added to the sodium citrate solution in a mass ratio of 1:9.

[0073] In a preferred embodiment, step (ii) is carried out at a temperature between 20 and 60°C, more preferably at 40°C.

[0074] In a preferred embodiment, step (ii) is performed for a residence time of between 1 and 23 h.

[0075] In a preferred embodiment, step (ii) is performed at a pH between 5.3 and 8.8, more preferably at a pH of 7.0.

[0076] In a preferred embodiment, in step (ii), an acid corresponding to the carboxylate is added for adjusting the pH.

[0077] In an even more preferred embodiment, in step (ii), citric acid is added for adjusting the pH.

[0078] In a preferred embodiment, the first base added to step (ii) is a carbonate salt selected from sodium carbonate, sodium bicarbonate or magnesium carbonate.

[0079] In an even more preferred embodiment, the first base added to step (iii) is sodium carbonate.

[0080] In a preferred embodiment, the sodium carbonate added to step (iii) is in a stoichiometric ratio of 1 :1 relative to the lead concentration in the third leach solution.

[0081] In a preferred embodiment, in step (iii), the precipitation reaction is carried out at a temperature between 20 and 90°C, more preferably at 70°C.

[0082] In a preferred embodiment, in step (iii), the precipitation reaction is carried out for a duration between 0.5 and 6 h.

[0083] In a preferred embodiment, in step (ii), the precipitation reaction is carried out at a pH between 6 and 9, more preferably at a pH of 7.5.

[0084] In a preferred embodiment, in step (iii), the pH adjustment is carried out with a neutralizing agent such as sodium hydroxide, without taking into account the neutralizing agent that provides hardness to the calcium and / or magnesium-based solution.

[0085] In a preferred embodiment, the first precipitated solution of step (iii) is recycled to step (ii) in order to leach the first leached sludge of step (i).

[0086] In a preferred embodiment, a portion of the first lead concentrate is recycled to step (iii) in order to act as a seed.

[0087] In an even more preferred embodiment, the recycled portion of the first lead concentrate corresponds to 30% of the total of the first precipitated solids.

[0088] In a preferred embodiment, the recycling rate of the first precipitated solution lacking lead of step (iii) is 90%.

[0089] In a preferred embodiment, the first precipitated solution lacking lead recycled to step (ii) requires sodium citrate restitución.

[0090] In a preferred embodiment, the sodium citrate restitución is a solution containing sodium citrate, the sodium citrate being used to prepare said solution in a mass ratio of 0.35:1 with respect to water, and is added so as to obtain a slurry with 10% w / w solids with respect to the second leached sludge.

[0091] In a preferred embodiment, the first precipitated solution lacking lead recycled to step (ii) requires a pH adjustment to 7.0.

[0092] In an even more preferred embodiment, the pH adjustment in step (ii) is carried out with a citric acid solution between 600 and 900 g / L.

[0093] In a preferred embodiment, the first precipitated solution lacking lead obtained from step (ii) does not require a step for removing sodium sulfate.

[0094] In a preferred embodiment, the first lead concentrate consists of lead carbonate.

[0095] In a preferred embodiment, the second base used in the leaching of step (iv) is selected between Mg(OH)2, KOH or NaOH.

[0096] In a preferred embodiment, the second base added in step (iv) is added in a ratio of between 5 and 10% w / w, more preferably in a ratio of 6.0% w / w relative to the total mass of the alkaline leaching solution.

[0097] In a preferred embodiment, the leaching reaction of step (iv) is carried out at a temperature between 70 and 150°C, more preferably at a temperature of 130°C.

[0098] In a preferred embodiment, the leaching reaction of step (iv) is carried out for a residence time of between 1 and 12 hours, more preferably for a residence time of 3 hours.

[0099] In a preferred embodiment, the acid used in the leaching in step (iv) is HCl.

[0100] In a preferred embodiment, HCl is provided in step (v) at a concentration varying from 50 to 140 g / L.

[0101] In a preferred embodiment, a chloride salt is added in step (v).

[0102] In an even more preferred embodiment, the chloride environment is enhanced in step (v) by adding magnesium chloride.

[0103] In a preferred embodiment, the chloride salt is provided in step (v) such that the chloride concentration is within the range of 140 and 240 g / L.

[0104] In a preferred embodiment, step (v) is performed at a pH between -1.5 and -0.25, preferably in the range of -0.73 to -0.65.

[0105] In a preferred embodiment, step (v) is performed at a temperature in the range of 40 to 95°C.

[0106] In a preferred embodiment, the neutralized slurry of step (vi) of metal precipitation is selected from calcium hydroxide, calcium oxide, calcium carbonate, lime, magnesium-containing lime, magnesium carbonate, magnesium hydroxide or magnesium oxide.

[0107] In an even more preferred embodiment, the neutralized slurry of step (vi) of metal precipitation is a magnesium oxide slurry.

[0108] In another preferred embodiment, step (vi) is performed at a temperature between 50 and 95°C.

[0109] In a preferred embodiment, the neutralized slurry added in step (vi) is provided until a pH between 3 and 7 is reached.

[0110] In another preferred embodiment, step (vi) has a residence time of between 0.5 and 3 h.

[0111] In a preferred embodiment, the chloride-rich fifth solution of step (vi) is sent to a crystallization process of magnesium chloride.

[0112] In another preferred embodiment, the chloride-rich fifth solution of step (vi) is recycled to the fourth step of silver precipitation.

[0113] In another preferred embodiment, the sulfuric acid solution of step (vii) has a sulfuric acid concentration between 60 and 275 g / L.

[0114] In another preferred embodiment, the sulfuric acid solution of step (vii) is a sulfuric acid leaching solution of smelter powder.

[0115] In another preferred embodiment, the sulfuric acid solution of step (vii) is the first leaching solution of step (i) rich in copper and iron, and optionally arsenic and bismuth, the acidity of which has been adjusted to between 60 and 275 g / L.

[0116] In a preferred embodiment, step (vii) of leaching the first precipitated solid rich in iron, copper and lead and optionally arsenic is carried out at a temperature between 50 and 95°C.

[0117] In a preferred embodiment, the second lead concentrate is recycled to step (ii).

[0118] In a preferred embodiment, the first leach solution rich in copper, iron and optionally arsenic is fed to a copper leaching process of smelting fines.

[0119] In a preferred embodiment, the first leach solution rich in copper, iron and optionally arsenic is sent to an arsenic elimination process.

[0120] In a preferred embodiment, the sixth leach solution rich in copper, iron and optionally arsenic is fed to a copper leaching process of the smelted powder.

[0121] In a preferred embodiment, the sixth leach solution, rich in copper, iron and optionally arsenic, is sent to an arsenic elimination process.

[0122] In a preferred embodiment, the arsenic elimination processes are selected from those that allow for the production of ferric arsenate.

[0123] In an even more preferred embodiment, the arsenic elimination process is a scorodite production process.

[0124] Application Examples

[0125] The following examples are to be regarded as embodiments of the present invention and are not to be regarded as limitations of the invention in any case, since the different modifications which can be made therein are to be covered within the subject matter claimed by the invention.

[0126] Sulfuric acid leaching

[0127] Examples 1 to 7

[0128] Between 2550 and 2850 g of a sulfuric acid solution containing HSO at a concentration between 150 and 250 g / L was prepared and placed in a 5 L glass reactor. Sludge previously subjected to a copper leaching process was added to a solids content between 5 and 10% w / w. The mineralogical composition of the sludge is shown in Table 1. The reactor was stirred at 85°C and 300 rpm for 3 to 6 hours. Once the reaction time was complete, the slurry was filtered using a Büchner system. The results are shown in Table 2.

[0129] Table 1 Mineralogical composition of mud

[0130] Substance Units Value <![CDATA[PbSO4]]> ％ 12.84 PbS ％ 0.1 PbO ％ 0.1 <![CDATA[CuSO4]]> ％ 2.54 <![CDATA[Cu2S]]> ％ 0.63 CuS ％ 4.02 CuO ％ 0.71 <![CDATA[CuOFe2O3]]> ％ 15.09 <![CDATA[ZnOFe2O3]]> ％ 4.46 ZnS ％ 2.94 <![CDATA[Fe3O4]]> ％ 4.74 Fe2O3 ％ 4.91 <![CDATA[FeS2]]> ％ 6.32 <![CDATA[Ag2S]]> ％ 0.1 <![CDATA[FeAsO4*2H2O]]> ％ 5.18 <![CDATA[Bi2O3]]> ％ 0.59 <![CDATA[Sb2O3]]> ％ 0.5 <![CDATA[KAl3Si3O 10 (OH)2]]> ％ 7.01 <![CDATA[Al2Si2O3(OH)4]]> ％ 2.92 Ge g / tonne 548

[0131] Table 2 Sulfuric acid leaching results of Examples 1 to 7

[0132]

[0133] Examples 8 to 10

[0134] 2550 g of a 250 g / L H2SO4 solution was prepared and placed in a 4 L autoclave. Sludge previously subjected to a copper leaching process was added to a solids content of 15% w / w. The reactor was stirred at 300 rpm and 130°C for 1 to 6 hours. Once the reaction time was complete, the slurry was filtered using a Büchner system. The results are shown in Table 3.

[0135] Table 3 Sulfuric acid leaching results of Examples 8 to 10

[0136] Variable / Example Units 8 9 10 Leach time h 1 3 6 Cu leach yield ％ 75.9 76.1 82.0 Mass loss ％ 35.0 41.0 42.0

[0137] Example 11

[0138] A refined wastewater solution (Table 4) was prepared, its H2SO4 concentration adjusted to 250 g / L, and placed in a 5 L glass reactor. 450 g of sludge previously subjected to a copper leaching process was added to produce a slurry with 15% w / w solids. The reactor was stirred at 300 rpm and 85°C for 6 hours. Once the reaction time was complete, the slurry was filtered using a Büchner system. The results showed a leaching yield of 72.0% for Cu, 62.0% for Fe, 71.5% for As, and 57.0% for Zn, with a mass loss of 38.5%.

[0139] Table 4 Composition of refining wastewater

[0140]

[0141]

[0142] Citric acid leaching

[0143] Example 12

[0144] A solution was prepared using 40 L of water, to which 14 kg of sodium citrate was added, and the pH was adjusted to 7.0 using 800 g / L of citric acid solution. Once the reagents had dissolved, 6 kg of leached sludge was added according to Example 3. The head sludge (La borra de cabeza) had a Pb content of 15.4%. Leaching was carried out at 20°C with stirring at 1000 rpm for 9 hours. A Pb leaching efficiency of 94% was achieved, resulting in a leached sludge with a mass reduction of 24% and a Pb content of 1.19%.

[0145] Examples 13 to 19

[0146] A solution was prepared using 2 L of water with a sodium citrate concentration between 323 and 368 g / L and a pH between 5.3 and 8.8. The pH was adjusted with 800 g / L of citric acid solution. Once the reagents were dissolved, the sludge treated according to Example 3 was added at a ratio of between 1.2 and 2.3 g of sodium citrate per gram of sludge. The sludge had a lead content between 15.0 and 15.1%. Leaching was performed at temperatures between 30 and 60°C and with stirring between 500 and 700 rpm for between 2 and 4 hours. The results are shown in Table 5.

[0147] Table 5 Citric acid leaching results of Examples 13 to 19

[0148]

[0149] Lead precipitation

[0150] Examples 20 to 25

[0151] Take 447 mL of a lead leach solution with a Pb concentration of 15.8 g / L. Add 400 g / L of NaOH solution to adjust the pH to between 7.0 and 8.0. The amount of sodium carbonate added should be equimolar to the amount of Pb present in the leach solution. Precipitation is carried out over a period of 1 hour at a temperature between 40 and 70°C.

[0152] Table 6 Lead precipitation results of Examples 20 to 25

[0153] Variable / Example Units 20 21 22 23 24 25 pH 7 7.5 8 7 7.5 8 Temperature ℃ 40 40 40 70 70 70 Pb precipitation yield ％ 72 89 91 95 96 97

[0154] Example 26

[0155] A solution was prepared using 4.7 L of water, to which 1633 g of sodium citrate was added, and the pH was adjusted to 7.0 with 800 g / L of citric acid solution. Once the reagents had dissolved, 700 g of the previously leached sludge was added according to Example 3. The first sludge had a lead content of 17.4%. Leaching was carried out at 40°C and stirred at 700 rpm for 3 hours. Subsequently, the slurry was filtered and the filtered solution was used for lead carbonate precipitation. The pH of the solution was adjusted to 7.5 with 400 g / L of NaOH solution, and then an equimolar amount of sodium carbonate was added. The precipitation was carried out for 1 hour, and the slurry was then filtered, while 90% of the filtered solution from the precipitation step was used to leach new sludge. For the second leaching cycle, sodium citrate solution was added at a sodium citrate:water ratio of 0.35:1, freshly used to adjust the solids content of the slurry to 10% w / w relative to the sludge content. Additionally, the pH of the solution was adjusted to 7.0 with 800 g / L of citric acid solution. Once the leaching pH was adjusted, a second Pb leaching cycle was performed, followed by a second Pb precipitation cycle using the same precipitation conditions as described at the beginning of Example 18. Recirculation of the filtered solution from the precipitation step was repeated until a total of 15 cycles were completed. For the final five cycles, the precipitated solids of lead carbonate were used as seeds for the precipitation step. The average leaching efficiency was 94%, with the leached solids having an average Pb content of 1.33%. The Pb precipitate obtained in the precipitation step had a Pb content of 76%. The sulfate content reached a maximum of 96 g / L and remained constant for the final five test cycles. Other analytes that concentrated during the various leaching cycles were Fe, reaching 4 g / L, Bi, reaching 1.5 g / L, and K, reaching 1.5 g / L. The element that leached most significantly during the testing was Bi, with an average leaching rate of 66%. Cu, Fe, Ag, Ge, Sb, As, Si, Al, and K had negligible leaching efficiency. In this context, the semi-continuous testing did not require a sodium sulfate removal step, and the concentrations achieved at each run did not affect leaching.

[0156] DRX analysis was performed on the precipitated solid, which showed the presence of lead carbonate (NaPb2(CO3)2OH) bound to sodium hydroxide. In turn, Raman analysis was performed, which showed a peak at 635.6 cm-1 and 1,011.9cm -1 There are two peaks corresponding to lead carbonate.

[0157] Alkaline leaching

[0158] Examples 27 to 35

[0159] A slurry with a solids content between 5.0 and 7.0% w / w was prepared using sodium hydroxide solution with a concentration between 5.4 and 8.7% w / w and leached sludge that had undergone sequential sulfuric acid and citric acid leaching. The slurry was placed in a 4 L autoclave and heated at 600 rpm at a temperature between 100 and 140°C for between 1 and 6 hours. Once the leaching time was complete, the slurry was cooled and filtered in a Büchner system. The results are shown in Table 7.

[0160] Table 7 Results of Examples 27 to 35

[0161]

[0162] Examples 36 to 37

[0163] A slurry was prepared using 6230 mL of water, to which was added 420 g of sodium hydroxide and 350 g of leached sludge from a continuous copper and lead leaching process to achieve a NaOH concentration of 6.0% w / w and 5.0% w / w solids. The slurry was placed in a 10 L glass reactor and heated at 90°C for between 1 and 6 hours with stirring at 900 rpm. Once the leaching time was complete, the slurry was cooled and filtered in a Büchner system.

[0164] Table 8 Results of Examples 36 and 37

[0165] Variable / Example Units 36 37 Residence time h 1 6 Leach yield Si ％ 63.2 63.0

[0166] hydrochloric acid leaching

[0167] Examples 38 to 45

[0168] Solutions were prepared with HCl concentrations between 54 and 160 g / L and chloride concentrations between 140 and 237 g / L. The chloride concentration was increased by adding magnesium chloride hexahydrate. To this solution was added 180 g of sludge that had undergone a sulfuric acid and citric acid leaching process and, on the other hand, a sulfuric acid, citric acid, and alkaline leaching process as described in Experiments 1 to 37. The slurry was added to a 5 L glass reactor, heated at 90°C, and maintained under constant stirring for 6 hours. Once the slurry test was completed, the slurry was filtered in a Büchner system. The results of these tests are shown in Table 9.

[0169] Table 9 Results of Examples 38 and 45

[0170]

[0171] The results show a significant contribution to copper leaching including the alkaline leaching step, which results in an improved copper leaching yield for the overall process. This observation is explained by the presence of chrysocolla in the sludge matrix from step (ii), which is effectively modified by the addition of silicon to remove the alkali in step (ii), making the copper more vulnerable to alkaline attack in step (v), as can be understood from the results presented herein.

[0172] The final residues from the hydrochloric acid leaching test were subjected to stability tests according to the TCLP and SPLP protocols, thereby obtaining cadmium, arsenic and lead release values ​​that were lower than the standard allowable values.

[0173] Metal precipitation

[0174] Examples 46 to 52

[0175] In a 600 mL precipitation flask, 450 g of PLS ​​obtained from the hydrochloric acid leaching test was placed and heated between 25 and 80° C. The hydrochloric acid leaching solution was neutralized using a 15% by volume magnesium oxide slurry until the pH was in the range of 3 to 6. Subsequently, the slurry was filtered using a 45 μm filter paper.

[0176] Table 10 Results of Examples 46 to 52

[0177] Variable / Example Units 46 47 48 49 50 51 52 Neutralisation pH - 6 3 3 5 5 6 6 Temperature ℃ 25 50 80 50 80 50 80 Precipitation yield Pb ％ >99.5 >99.5 >99.5 >99.5 >99.5 >99.5 >99.5 Fe ％ 97 85 80 98 97 100 99 Cu ％ 18 5 2 16 15 20 23

[0178] Lead concentrate

[0179] Examples 53 to 55

[0180] 50 g of a metal precipitate having a content of 17.1% Fe, 1.3% Cu and 0.44% Pb was placed in a 600 mL precipitation flask and sulfuric acid was added to obtain a concentration of H2SO4 between 60 and 257 g / L. The slurry was stirred with a magnetic bar and maintained at a temperature of 60°C for 5 h, and once the slurry leaching time was complete, the slurry was filtered using 45 μm paper.

[0181] Table 1 Results of Examples 53 to 55

[0182] Variable / Example Units 53 54 55 Sulphuric acid concentration - 60 80 120 Temperature ℃ 60 60 60 Residence time h 3 3 3 Concentrate regime Pb ％ 2.0 5.4 14.0 Fe ％ 37.6 29.3 6.2 Cu ％ 0.12 0.09 0.04

[0183] Examples 56 to 57

[0184] 1700 g of a metal precipitate with a content of 17.1% Fe, 1.3% Cu, and 0.44% Pb was placed in a 20 L glass reactor, to which 15300 g of a 275 g / L H2SO4 solution was added. The slurry was maintained between 25 and 80°C and mechanically stirred at 450 rpm for 5 h, and once the leaching time was complete, the slurry was filtered using filter paper No. 42.

[0185] Table 2 Results of Examples 56 to 57

[0186] Variable / Example Units 56 57 Sulphuric acid concentration - 275 275 Temperature ℃ 25 80 Residence time h 5 5 Concentrate regime Fe ％ 4.0 3.5 Cu ％ 0.35 0.15 Pb ％ 29.0 28.6

[0187] Citric acid leaching of lead concentrate

[0188] Examples 58 to 61

[0189] The lead concentrate was recycled to the citric acid leaching step. In a 5-liter glass reactor, 120 grams of lead concentrate from the metal precipitation leaching test and a 0.5- to 1-M sodium citrate solution adjusted to pH 7 with citric acid were added. The slurry was maintained between 20 and 70°C and stirred at 700 rpm for 3 hours. The lead leaching yield varied between 80 and 82%.

[0190] Table 3 Results of Examples 58 to 61

[0191]

[0192]

Claims

1. A method for leaching copper and lead from metallurgical residues of smelting powder which has undergone a copper leaching process and contains copper, iron, lead, silicon, arsenic, antimony and bismuth, characterized in that it comprises: i. leaching copper from the metallurgical residue with a first acid solution so as to obtain a first leaching solution rich in copper and iron, arsenic, antimony and bismuth and a first leached sludge having a reduced copper and iron content, a reduced arsenic content and rich in lead and silicon, ii. leaching a first leachable sludge, wherein said first leachable sludge is treated with a first solution of a carboxylate salt so as to obtain a second leachable sludge deficient in lead and a second leach solution enriched in lead, iii. precipitation, wherein a first base is added to the second leaching solution enriched in lead so as to obtain a first lead concentrate and a first precipitation solution deficient in lead, iv. alkaline leaching of the second leachate sludge, wherein a second base is added to form an alkaline leach solution so as to obtain a third leachate sludge having a reduced silicon content and a third leach solution enriched in silicon and arsenic, v. leaching the third leachable sludge with hydrochloric acid using an acid solution in a chloride environment to obtain a fourth leachable sludge for final disposal and a fourth leach solution rich in copper, lead, iron and arsenic, vi. precipitating metals from the fourth leach solution rich in copper, lead, iron and arsenic using a neutralizing slurry selected from calcium hydroxide, calcium oxide, calcium carbonate, magnesium carbonate, magnesium hydroxide or magnesium oxide to produce a fifth solution rich in chloride and a first precipitated solid rich in iron, copper, lead and arsenic, and vii. Leaching the first precipitated solid rich in iron, copper, lead and arsenic with a sulfuric acid solution to produce a sixth leach solution rich in copper, iron and arsenic and a second lead concentrate.

2. The method according to claim 1, characterized in that the metallurgical residue to be processed is a powder obtained by a metal smelting process.

3. The method according to claim 2, characterized in that the powder obtained by a copper smelting process is a smelting powder.

4. The method according to claim 3, characterized in that the metallurgical residue has been subjected to leaching with H2SO4.

5. The method according to claim 1 , wherein the metallurgical residue to be processed comprises the mineral substances lead vanadium, indigo copper ore, copper-iron spinel in the form of CuOFe2O3, zinc spinel in the form of ZnOFe2O3, magnetite, iron (III) oxide, pyrite, scorodite, muscovite, kaolinite and lead (II) sulfate.

6. The method according to claim 5, characterized in that the copper contained in the metallurgical residue is present as copper sulfate, chalcocite, covellite and copper-iron spinel in the form of CuOFe2O3.

7. The method according to any one of claims 1 to 6, characterized in that the silicon contained in the metallurgical residue is present as muscovite and kaolinite.

8. The method according to claim 1, wherein the lead contained in the metallurgical residue is present as lead(II) sulfate, galena or lead(II) oxide.

9. The method according to claim 8, characterized in that at least 95% of the lead is as lead(II) sulfate.

10. The method according to any one of claims 1 to 9, characterized in that the first acid solution of step (i) comprises H2SO4 and / or refinery wastewater.

11. The method according to any one of claims 1 to 10, characterized in that step (i) is carried out at a H2SO4 concentration between 150 and 300 g / L.

12. The process according to any one of claims 1 to 11, characterized in that step (i) is carried out at a temperature between 50 and 130°C.

13. The method according to any one of claims 1 to 12, characterized in that step (i) is carried out for a time of between 3 and 12 hours.

14. The process according to any one of claims 1 to 13, characterized in that step (i) is carried out at a solids concentration of between 5 and 20% w / w.

15. A method according to any one of claims 1 to 14, characterised in that in the leaching step (ii), the carboxylate salt is sodium citrate.

16. The method according to claim 15, characterized in that in step (ii) the sodium citrate solution has a molar concentration of sodium citrate between 0.5 and 1 M.

17. The method according to any one of claims 1 to 16, characterized in that in step (ii), the first leached sludge is added to the sodium citrate solution at a mass ratio of 1:

9.

18. The process according to any one of claims 1 to 17, characterized in that step (ii) is carried out at a temperature between 20 and 60°C.

19. The process according to any one of claims 1 to 18, characterized in that step (ii) is carried out for a residence time of between 1 and 23 h.

20. The process according to any one of claims 1 to 19, characterized in that step (ii) is carried out at a pH between 5.3 and 8.

8.

21. The process according to any one of claims 1 to 20, characterized in that citric acid is added in step (ii) to adjust the pH.

22. The method according to any one of claims 1 to 21, characterized in that the pH adjustment in step (ii) is performed with 600 and 900 g / L citric acid solutions.

23. The process according to any one of claims 1 to 22, characterized in that the first base added to step (iii) is a carbonate selected from sodium carbonate, sodium bicarbonate or magnesium carbonate.

24. The process according to claim 23, characterized in that the first base added to step (iii) is sodium carbonate.

25. The method of claim 24, wherein the sodium carbonate added to step (iii) is in a stoichiometric ratio of 1:1 relative to the lead concentration in the third leach solution.

26. The process according to any one of claims 1 to 25, characterized in that in step (iii), the precipitation reaction is carried out at a temperature between 20 and 90°C.

27. The method according to any one of claims 1 to 26, characterized in that, in step (iii), the precipitation reaction is carried out for 0.5 to 1 hour.

28. The method according to any one of claims 1 to 27, characterized in that, in step (iii), the precipitation reaction is carried out at a pH between 6 and 9.

29. The method according to claim 28, characterized in that, in step (iii), the pH adjustment is carried out using sodium hydroxide.

30. The method according to any one of claims 1 to 29, characterized in that the first precipitation solution of step (iii) is recycled to step (ii) in order to leach the first leached sludge of step (i).

31. A process according to any one of claims 1 to 29, characterised in that a portion of the first lead concentrate is recycled to step (iii) in order to serve as seed crystals.

32. The method according to claim 31, characterized in that the portion of the first lead concentrate that is recycled corresponds to 30% of the total amount of the first precipitation solids.

33. The method according to any one of claims 1 to 32, characterized in that the recycling rate of the lead-deficient first precipitation solution in step (iii) is 90%.

34. The method of claim 33, wherein the recycling of the lead-deficient first precipitation solution to step (ii) requires the addition of an additional input of sodium citrate.

35. The method of claim 34, wherein the additional input of sodium citrate is a solution containing sodium citrate in a mass ratio of 0.35:1 sodium citrate to water used to prepare the solution, and wherein the solids content of the slurry is adjusted to 10% w / w relative to the amount of the second leachable sludge.

36. The method according to any one of claims 34 to 35, characterized in that the first lead-deficient precipitation solution recycled to step (ii) requires pH adjustment to 7.

0.

37. The method according to claim 36, characterized in that the pH adjustment in step (ii) is performed with a 600 to 900 g / L citric acid solution.

38. The method according to any one of claims 34 to 36, characterized in that the first lead-deficient precipitation solution obtained from step (iii) does not require a step of removing sodium sulfate.

39. A method according to any one of claims 1 to 38, characterised in that the first lead concentrate is lead carbonate.

40. A method according to any one of claims 1 to 39, characterised in that the second base used in the leaching of step (iv) is selected between Mg(OH)2, KOH or NaOH.

41. A process according to any one of claims 1 to 40, characterised in that the second base added in step (iv) is added in a ratio of between 5 and 10% w / w relative to the total mass of the alkaline leaching solution.

42. The method according to claim 41, characterized in that the second base added in step (iv) is added in a ratio of 6.0% w / w relative to the total mass of the alkaline leaching solution.

43. The method according to any one of claims 1 to 42, characterized in that the leaching reaction of step (iv) is carried out at a temperature between 90 and 140°C.

44. The method according to any one of claims 1 to 43, characterized in that the leaching reaction of step (iv) is carried out for a residence time of between 1 and 6 hours.

45. The process according to claim 44, characterized in that in step (v) HCl is provided in a concentration varying from 50 to 140 g / L.

46. ​​The method according to any one of claims 1 to 45, characterized in that in step (v) the chloride environment is increased by adding a chloride salt.

47. The method of claim 46, wherein the chloride environment is increased in step (v) by adding magnesium chloride.

48. A method according to any one of claims 46 to 47, characterised in that in step (v) chloride is provided at a concentration between 140 and 240 g / L.

49. The process according to any one of claims 1 to 48, characterized in that step (v) is carried out at a pH between -1.5 and -0.

25.

50. The process according to any one of claims 1 to 49, characterized in that step (v) is carried out at a temperature between 40 and 95°C.

51. The process according to any one of claims 1 to 50, characterized in that step (vi) is carried out at a temperature between 50 and 95°C.

52. The method according to any one of claims 1 to 50, characterized in that the neutralized slurry added in step (vi) is provided until a pH between 3 and 7 is reached.

53. The process according to any one of claims 1 to 52, characterized in that step (vi) has a residence time of between 0.5 and 3 hours.

54. The process according to any one of claims 1 to 53, characterized in that the chloride-rich fifth solution of step (vi) is sent to a crystallization process of magnesium chloride.

55. The method according to any one of claims 1 to 54, characterized in that the sulfuric acid solution of step (vii) has a sulfuric acid concentration between 60 and 275 g / L.

56. The process according to any one of claims 1 to 55, characterized in that the sulfuric acid solution of step (vii) is the first leaching solution of step (i) rich in copper and iron, and optionally arsenic and bismuth, whose acidity has been adjusted to between 60 and 275 g / L.

57. The method according to any one of claims 1 to 56, wherein the sulfuric acid solution in step (vii) is a sulfuric acid leaching solution of smelting powder.

58. The method according to any one of claims 1 to 57, characterized in that the sulfuric acid solution of step (vii) is the first leaching solution of step (i) rich in copper and iron and optionally arsenic and bismuth.

59. The method according to any one of claims 1 to 58, characterized in that step (vii) of leaching the first precipitated solid rich in iron, copper and lead and optionally arsenic is carried out at a temperature between 50 and 95°C.

60. A process according to any one of claims 1 to 59, characterised in that the second lead concentrate is recycled to step (ii).

61. The method according to any one of claims 1 to 60, characterized in that the first copper-rich leaching solution is sent to a copper leaching process of the smelting powder.

62. The method according to any one of claims 1 to 61, characterized in that the first leaching solution rich in copper is sent to an arsenic elimination process.

63. The method of claim 62, wherein the arsenic elimination process is a scorodite production process.

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