Process for producing silver concentrates from metallurgical residues - Patents.com

Abstract

A method for producing a silver concentrate from metallurgical residues, particularly from residues containing copper, iron, lead, silicon, silver, and antimony, and which may optionally contain elements such as arsenic and bismuth, comprises (i) copper leaching the metallurgical residue with a first acidic solution to produce a first leaching solution that is rich in copper and iron, and optionally rich in arsenic, and a first leached solution that is reduced in copper and iron, optionally reduced in arsenic, and has a content rich in lead, germanium, silver, and silicon. (ii) leaching the first leached sludge with a first solution of carboxylate to obtain a second leached sludge depleted in lead and a second leaching solution enriched in lead; (iii) adding a base to form an alkaline leaching solution and alkaline leaching the second leached sludge to obtain a third leached sludge having a reduced silicon content and a third leached sludge enriched in silicon and optionally enriched in arsenic. (iv) hydrogen chloride leaching the third leached sludge using an acidic solution in a chloride environment to obtain a fourth leached sludge for final treatment and a fourth leaching solution rich in silver, copper, lead, and iron, and optionally rich in arsenic; (v) silver precipitation from the fourth leaching solution rich in silver, copper, and iron, and optionally rich in arsenic, using a neutralized slurry to obtain a fifth solution rich in chlorides and a fifth solution rich in iron, copper, lead, and silver. (vi) leaching the first precipitated solids rich in iron, copper, lead, and iron, optionally rich in arsenic, with a sulfuric acid solution to produce a sixth leaching solution rich in copper and iron, optionally rich in arsenic, and a first silver-lead concentrate; and (vii) leaching the first silver concentrate with a second carboxylate solution to produce a seventh leaching solution and a second silver concentrate.

Description

Detailed Description of the Invention

[0001] [Field of the Invention] The present invention relates to a method (procedure) for producing a silver concentrate from metallurgical residues, in particular from residues containing copper, iron, lead, silicon and silver, and which may optionally contain elements (components) such as arsenic, antimony and bismuth.

[0002] In a more particular embodiment, the metallurgical residue is a powder obtained from a metal smelting process.

[0003] In a more particular embodiment, the metallurgical residue is a powder obtained from a copper smelting process.

[0004] In a more specific embodiment, metallurgical residue (or, in particular, smelter powder) contemplates material that has already been subjected to a leaching process, such as sulfuric acid leaching.

[0005] In this disclosure, metallurgical residues that have been subjected to a prior leaching process are considered as sludge.

[0006] [Latest technology] <Copper leaching> The copper in the sludge is mainly composed of species such as ferrite and / or spinel in the form CuFe2O4, zinc as ZnFe2O4, and the associated part of iron, i.e. FeFe2O4. The leaching of these species is based 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). BS Boyanov et al. carried out a study on the leaching of synthetic ferrites of zinc, copper, and cadmium, evaluating each of the variables mentioned above. The results of this study show that at high temperatures and high acid concentrations, ferrites are better dissolved by HCl and H2SO4.

[0007] At high acid concentrations, copper leaching has been observed to have an asymptotic behavior with respect to leaching temperature: after a reaction time of 60 minutes in sulfur medium, copper leaching yields of over 90% are reached at temperatures of 85-90 °C.

[0008] Lead leaching and precipitation The world lead consumption in 2011 was over 10 million tons, of which about 80% was for the production of lead-acid batteries. These batteries contain lead amounts in the form of Pb, PbO2, and PbSO4. The most traditional method of recovering lead is pyrometallurgy, characterized by the addition of reducing agents (e.g. carbon powder, iron scrap, and sodium oxalate). This operation is carried out in ovens at temperatures above 1000 °C. As shown in He et al., Minerals 7, no. 6 (2017): 93, this leads to a high demand for energy processes.

[0009] On the other hand, hydrometallurgical processes for recovering lead allow operation at lower temperatures and with lower energy consumption. Hydrometallurgical processes then do not produce sulfur dioxide, a gas that is harmful to the environment. Hydrometallurgical processes use desulfurizing agents such as sodium carbonate, ammonium carbonate, sodium bicarbonate, ammonium bicarbonate, sodium hydroxide, sodium citrate, acetic acid, and sodium acetate. The purpose of these processes is to exchange sulfate ions with other anions to form insoluble salts. Once recovered, lead salts such as lead citrate can be calcined to produce lead oxide (Zarate-Gutierrez and Lapidus, Hydrometallurgy 144 (2014): 124-128).

[0010] <Desulfurization using citrate> In the specific case of using citrate, a mixture of citric acid and sodium citrate is beneficial for the leaching of lead sulfate and the subsequent crystallization of lead citrate.

[0011] Lead leaching in citrate solutions The solubility product constant of anglesite at 20°C is 6.31·10 -7 This indicates that the solubility of PbSO4 is significantly reduced. However, in the presence of concentrated citrate solutions, lead forms a series of soluble complexes. 2+ In solutions having a pH of 4.6 to 11.5, a wide variety of citrate complex species are present. Below a pH of 4.6, lead sulfate dominates, whereas above a pH of 11.5, lead hydroxide dominates.

[0012] He et al., Minerals 7, no. 6 (2017): 93, studied lead leaching for a paste with a lead sulfate to water weight ratio of 1:10 by adding 650 g / L sodium citrate at 35°C. These conditions allowed the conversion of more than 99% of the lead sulfate to lead citrate after 60 minutes of reaction time. Increasing the temperature to 95°C at a sodium citrate concentration of 300 g / L allowed an efficiency of nearly 99% to be achieved after 60 minutes of reaction time. However, a decrease in the production of lead citrate was observed when citric acid was introduced into the mixture. The optimum pH for the production of lead citrate was in the range of 6-7. A high lead leaching efficiency was also obtained from acid and lead-acid batteries at pH 5.5 with citric acid and ammonium agents. In the pH range of 5.2-5.5, the presence of lead citrate trihydrate ([Pb3(C6H5O7)2]·[3H2O]) was reported as the main species. In the higher pH range of 8-10, the recovery of lead 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 is carried out at 20 °C for 15-60 min with molar ratios of citric acid to lead(II) and (IV) oxides of 1:1 and 4:1. In this case, leaching efficiencies of more than 99 wt.% are reached, yielding Pb(C6H6O7)·H2O as the main species (Sonmez and Kumar, Hydrometallurgy 95, no. 1-2 (2009), 82-86).

[0013] Pulp density is another important parameter for lead leaching with citrate solutions. In the range of 10-50 g / L anglesite pulp, a higher level of lead extraction of 90-94% was achieved at 10 g / L pulp consistency for a leaching solution of 1M sodium citrate solution, pH 7, at 600 rpm and 25°C. At higher pulp consistency, the amount of lead extracted was less.

[0014] Therefore, the hydrometallurgical desulfurization process is influenced by the diffusion of citrate ions in the lead paste within the reactor due to the high density of the lead paste, and it is therefore important to design the reactor to maximize the mass transfer in the system.

[0015] A technology based on lead recovery from lead waste using citric acid has been developed by Cambridge Enterprises Limited (WO2008056125A1). Essentially, the technology involves treating lead residues, including lead(II) oxide, lead(IV) oxide, and lead sulfate, with citric acid solution. Alternatively, the lead residues may be treated in combination with sodium citrate at pHs ranging from 1.4 to 6. Finally, hydrogen peroxide can be added as a reducing agent in a basic environment to accelerate the lead(IV) oxide leaching reaction to produce lead citrate (Sonmez and Kumar, Hydrometallurgy 95, no. 1-2 (2009), 82-86).

[0016] The present invention differs from the above patent (WO2008056125A1) in that the pH required for leaching ranges from 5.33 to 8.8. In the present invention, a pH equal to 7 is preferably used. Furthermore, the present invention proposes recirculating the citrate solution obtained after the precipitation step with sodium carbonate to re-leaching the output metallurgical residue from the sulfuric acid leaching step.

[0017] <Alkaline leaching> Mufakhir et al. studied silicon leaching in the presence of sodium hydroxide from slag obtained from the ferronickel acquisition process in IOP Conf. Series: Materials Science and Engineering 285 (2017) 012003. Alkaline leaching results show a maximum silicon leaching yield of 31% when the slag is evaluated at NaOH concentrations of 6-14 mol / L, solid contents of 5-25% w / w, and temperatures of 25-110°C. For the purposes of the present invention, a method is disclosed for maximizing copper and lead leaching. The method includes sulfuric acid leaching and citric acid leaching steps with the objective of removing Cu and Pb present in the sludge, followed by alkaline leaching. Removing Cu, Fe, and Pb in an early stage allows the sludge to be chemically modified, leaving silicon species more vulnerable to leaching, as shown in the results obtained in this application.

[0018] <Hydrogen leaching> Patent US7329396 describes a process for leaching precious metals of oxidized materials such as lateritic nickel minerals. The process includes leaching the mineral with a lixiviant containing a cationic salt (e.g. magnesium chloride) and hydrochloric acid. Additional oxidizing agents or metal chlorides (such as those obtained from the leaching operation) may be added. In one embodiment, the process includes recovery of precious metal minerals, including (i) leaching the mineral with a lixiviant, (ii) separating the leachate value rich in metal minerals in a first solid-liquid separation, (iii) oxidizing and neutralizing the metal-rich leachate value thus obtained, and (iv) separating the magnesium chloride solution from the leachate thus obtained in a second solid-liquid separation. In another embodiment, the lixiviant solution is regenerated from the magnesium chloride solution. In yet another embodiment, the regeneration of the leachate solution includes a step for producing magnesium oxide from the magnesium chloride solution.

[0019] The difference between the present invention and patent application US7329396 is that the latter indicates a preferred pH above 0.4 to precipitate hematite. In the present case, it is advantageous to work at a low pH, preferably below pH-0.25, with the aim of obtaining ferric ions in the solution. This favors the use of the leaching solution in other leaching processes, such as the molten powder leaching process. Moreover, in the present invention, whenever the silver to iron concentration ratio reaches 0.01 g Ag / g Fe, the precipitation of iron hydroxide is completely disadvantaged. As a result, the precipitation of iron hydroxide can drag the silver out of the solution.

[0020] Patent application CA2820631A1 relates to a process that may be efficient for treating multiple materials containing many different metals. These materials may be leached with HCl to obtain a leachate and solids. The leachate and solids may then be separated from each other and the first leach metal may be separated. The second metal may then be separated from the leachate. The first metal and the second metal may each be substantially separated from the leachate. This may be done by controlling the leachate temperature, adjusting the pH, and further reacting the leachate with HCl, etc. The metals that may be recovered in the form of metal chlorides may finally be converted to the corresponding metal oxides. Thus, recovery of HCl is possible. The multiple metals may be selected from among aluminum, iron, zinc, copper, gold, silver, molybdenum, cobalt, magnesium, lithium, manganese, nickel, palladium, platinum, thorium, phosphorus, uranium, titanium, rare earth and rare metal elements.

[0021] The difference between the present invention and patent application CA2820631A1 is that the former does not require temperatures above 90°C to perform silver leaching efficiently, while the latter requires temperatures above 125°C. Furthermore, leaching of aluminum-containing materials is performed starting from a hydrochloric acid concentration of 18%, whereas in the present invention, a hydrochloric acid concentration of less than 140g / L (or less than 11% w / w) is required.

[0022] [Drawing Description] FIG. 1 shows a process diagram of the method disclosed by the present invention.

[0023] [Description of the invention] In one broad aspect, the present invention describes a method for producing a silver concentrate from metallurgical residues, in particular from residues containing copper, iron, lead, silicon, silver and antimony (which may optionally contain elements such as arsenic and bismuth), the method comprising: Step (I): subjecting the metallurgical residue (1) to copper leaching using a first acidic solution (2), a first leaching solution (3) rich in copper and iron, and optionally rich in arsenic and bismuth; obtaining a first leached sludge (4) having a copper and iron reduced content, optionally arsenic reduced content, and a lead and silicon rich content; Step (II): leaching the first leached sludge (4) with a first solution of a carboxylate (5), A second leached sludge depleted in lead (6); obtaining a second leaching solution (7) rich in lead; Step (III): Adding a base to generate an alkaline leaching solution (8) and subjecting the second leached sludge (6) to alkaline leaching; a third leached sludge (9) having a reduced silicon content; obtaining a third leaching solution (10) rich in silicon and optionally rich in arsenic; step (iv): silver leaching the third leached sludge (9) using an acidic solution (11) in a chloride environment, A fourth leached sludge (12) for final treatment; obtaining a fourth leaching solution (13) enriched in silver, copper, lead, and iron, and optionally enriched in arsenic; step (v): silver precipitation from said fourth leaching solution (13), rich in silver, copper and iron, and optionally rich in arsenic, using a neutralized slurry (14); a fifth chloride-rich solution (15); a first precipitated solid (16) rich in iron, copper, lead, and silver, and optionally rich in arsenic; step (vi): leaching the first precipitated solid (16), rich in iron, copper, lead and iron, and optionally rich in arsenic, with a sulfuric acid solution (17), a sixth leaching solution (16) rich in copper and iron, and optionally rich in arsenic; producing a first silver-lead concentrate (first silver and lead concentrate) (19); and step (vii): leaching the first silver concentrate (19) with a second carboxylate solution (20), A seventh leaching solution (21); a second silver concentrate (22); Contains:

[0024] In one preferred option, the metallurgical residue to be treated is a powder obtained by a metal smelting process.

[0025] In a further preferred option, said powder obtained by a copper smelting process is a smelted powder.

[0026] In an even more preferred option, said metallurgical residue is subjected to a copper leaching process.

[0027] In an even more preferred option, the metallurgical residue is subjected to leaching with sulphuric acid.

[0028] In one preferred option, the metallurgical residue to be treated contains the mineral species anglesite, coberlin, cuprospinel in the form CuOFe2O3, zinc spinel in the form ZnOFe2O3, magnetite, iron(III) oxide, pyrite, scorodite, mucovite, kaolinite and lead(II) sulfate.

[0029] In an even more preferred option, the copper contained in said metallurgical residue is present as copper sulfate, chalcosine, cobalt and cuprospinel in the form CuOFe2O3.

[0030] In an even more preferred option, the copper contained in said metallurgical residue is present to at least 50% in the form of cuprospinel of the CuOFe2O3 form.

[0031] In one preferred option, the silicon contained in the metallurgical residue is present as muscovite and kaolinite.

[0032] In another preferred option, the lead contained in the metallurgical residue is present as lead(II) sulfate, galena or lead(II) oxide.

[0033] In an even more preferred option, at least 95% of the lead is found as lead (II) sulfate.

[0034] In one preferred option, the first H2SO4 solution may contain sulfuric acid and / or refinery wastewater.

[0035] In one preferred option, step (i) is carried out at a sulfuric acid concentration of 150-300 g / L, more preferably at a sulfuric acid concentration of 250 g / L.

[0036] In one preferred option, step (i) is carried out at a temperature between 50 and 130°C, more preferably at a temperature of 85°C.

[0037] In one preferred option, step (i) is carried out for a period of 3 to 12 hours, more preferably for a residence time of 6 hours.

[0038] In one preferred option, step (i) is carried out at a solids concentration of 5-20% w / w, more preferably at a solids concentration of 15% w / w.

[0039] In one preferred option, the carboxylate in leaching step (ii) is sodium citrate.

[0040] In one preferred option, in step (ii), the sodium citrate solution has a molarity of sodium citrate between 0.5 and 1 M.

[0041] In one preferred option, in step (ii), the first leached sludge is fed to a sodium citrate solution in a mass ratio of 1:9.

[0042] In one preferred option, step (ii) is carried out at a temperature between 20 and 60°C, more preferably at a temperature of 40°C.

[0043] In one preferred option, step (ii) is carried out for a residence time of from 1 to 23 hours.

[0044] In one preferred option, step (ii) is carried out at a pH of 5.3 to 8.8, more preferably at pH 7.0.

[0045] In one preferred option, in step (ii) an acid corresponding to the carboxylate is added to adjust the pH.

[0046] In an even more preferred option, in step (ii) citric acid is added to adjust the pH.

[0047] In an even more preferred option, the pH adjustment in step (ii) is carried out using a 600-900 g / L citric acid solution.

[0048] In one preferred option, the base used for leaching in step (iii) may be selected from potassium hydroxide, magnesium hydroxide, or sodium hydroxide.

[0049] In one preferred option, the base added in step (iii) is added in a ratio of 5-10% w / w, based on the total mass of the alkaline leaching solution, more preferably in a ratio of 6.0% w / w, based on the total mass of the alkaline leaching solution.

[0050] In one preferred option, the leaching reaction in step (iii) is carried out at a temperature between 70 and 150°C, more preferably at a temperature of 130°C.

[0051] In one preferred option, the leaching reaction in step (iii) is carried out for a residence time of 1 to 12 hours, more preferably for a residence time of 3 hours.

[0052] In one preferred option, the acid used for leaching in step (iv) is hydrochloric acid.

[0053] In one preferred option, in step (iv), hydrochloric acid is provided at a concentration of 50-140 g / L.

[0054] In one preferred option, in step (v), the chloride environment is increased by addition of a chloride salt.

[0055] In one preferred option, in step (iv), the chloride environment is increased by the addition of magnesium chloride.

[0056] In one preferred option, chloride is provided in step (iv) at a concentration of 140-240 g / L.

[0057] In one preferred option, step (iv) is carried out at a pH of -1.5 to 0, preferably in the range of pH -0.73 to -0.65.

[0058] In one preferred option, step (iv) is carried out at a temperature between 40 and 95°C.

[0059] In one preferred option, the neutralisation slurry in step (v) of silver deposition is selected from calcium hydroxide, calcium oxide, calcium carbonate, lime, dolomitic lime, magnesium carbonate, magnesium hydroxide or magnesium oxide.

[0060] In an even more preferred option, the neutralised slurry in step (vi) of the silver deposition is a magnesium oxide slurry.

[0061] In another preferred option, step (v) is carried out at a temperature between 50 and 95°C.

[0062] In another preferred option, the neutralized slurry added in step (v) is fed until a pH of 3-7 is reached.

[0063] In another preferred option, step (v) has a residence time of from 0.5 to 3 hours.

[0064] In one preferred option, the chloride-rich fifth solution in step (v) is sent to a magnesium chloride crystallization process.

[0065] In another preferred option, the fifth chloride-rich solution in step (v) is recycled to the silver deposition step (iv).

[0066] In another preferred option, the sulfuric acid solution in step (vi) has a sulfuric acid concentration of 60-275 g / L.

[0067] In another preferred option, the sulphuric acid solution in step (vi) is a sulphuric acid leaching solution of smelting powder.

[0068] In another preferred option, the sulphuric acid solution in step (vi) is the first leaching solution in step (i) having an acidity adjusted to between 60 and 275 g / L and enriched in copper and iron, and optionally enriched in arsenic and bismuth.

[0069] In one preferred option, step (vi) of leaching the silver precipitate is carried out at a temperature comprised between 50 and 95°C.

[0070] In one preferred option, the carboxylate in step (vii) of leaching said first silver concentrate is preferably sodium citrate.

[0071] In an even more preferred option, the sodium citrate concentration is between 0.5 and 1M.

[0072] In one preferred option, step (vii) of leaching the first silver concentrate is carried out at a temperature between 25 and 90°C.

[0073] In one preferred option, step (vii) of leaching the first silver concentrate is carried out at a solids content of 5-10%.

[0074] In one preferred option, step (vii) of leaching the first silver concentrate is carried out for a period of from 1 to 6 hours.

[0075] In one preferred option, the seventh leaching solution is recycled to the leaching in step (ii).

[0076] In one preferred option, the second silver concentrate is constituted by silver antimonate.

[0077] In an even more preferred option, the second silver concentrate is composed of silver antimonate and lead antimonate.

[0078] In one preferred option, the copper-rich first leaching solution is sent to a copper leaching process of smelting powder.

[0079] In one preferred option, the sixth leaching solution, rich in copper and iron, and optionally rich in arsenic, is sent to a copper leaching process of smelting powder.

[0080] In one preferred option, the copper-rich first leaching solution is sent to an arsenic removal process.

[0081] In one preferred option, the sixth leaching solution (16), rich in copper and iron, and optionally rich in arsenic, is sent to an arsenic removal process.

[0082] In one preferred option, the arsenic removal process is selected from among those intended to produce ferric arsenate.

[0083] In an even more preferred option, the arsenic removal process is a scorodite production process.

[0084] [Example of application] The following examples should be considered as embodiments of the present invention, and they should not be considered in any way as limiting the embodiments, since various adaptations (modifications) that can be performed to the embodiments are included within the scope of the subject matter claimed by the present invention.

[0085] [Sulfuric acid leaching] <Examples 1 to 7> 2.550-2.850 g of sulfuric acid solution with H2SO4 concentration of 150-250 g / L was prepared. The solution was placed in a 5 L glass reactor. Sludge previously subjected to a copper leaching process was added to a solid content of 5-10% w / w. Table 1 shows the mineralogy of the sludge. The reactor was stirred at 300 rpm at 85°C for 3-6 hours. After the reaction time was over, the pulp was filtered by a Buchner system. Table 2 shows the results. [Table 1] [Table 2]

[0086] <Examples 8 to 10> 2,550 g of 250 g / L H2SO4 solution was prepared. The solution was placed in a 4 L autoclave. Sludge previously subjected to a copper leaching process was added to a solid content of 15% w / w. The reactor was stirred at 300 rpm at 130°C for 1-6 hours. After the reaction time was over, the pulp was filtered by a Buchner system. Table 3 shows the results. [Table 3]

[0087] <Example 11> A refinery wastewater solution was prepared (Table 4). The H2SO4 concentration of the solution was adjusted to 250 g / L. The solution was placed in a 5 L glass reactor. 450 g of sludge previously subjected to a copper leaching process was added to produce a pulp with a solid content of 15% w / w. The reactor was stirred at 300 rpm for 6 hours at 85°C. After the reaction time was over, the pulp was filtered by a Buchner system. The results showed a Cu leaching yield of 72.0%, Fe leaching yield of 62.0%, As leaching yield of 71.5%, Zn leaching yield of 57.0%, and a mass loss (mass defect) of 38.5%. [Table 4]

[0088] [Citric acid leaching] <Example 12> A solution was prepared with 40 L of water. 14 kg of sodium citrate was added to the solution and the pH was adjusted to 7.0 with 800 g / L of citric acid solution. After the test solution (reaction solution) had dissolved, 6 kg of leached sludge was added according to Example 3. The head sludge had a Pb content of 15.4%. Leaching was carried out at 20°C and stirred at 1,000 rpm for 9 hours. A Pb leaching efficiency of 94% was obtained. A leached sludge was thus obtained with a Pb content of 1.19% and a mass reduction of 24%.

[0089] <Examples 13 to 19> A solution was prepared using 2 L of water with sodium citrate at a concentration of 323-368 g / L with a pH of 5.3-8.8. The pH was adjusted with 800 g / L citric acid solution. After the reagent solution was dissolved, sludge treated according to Example 3 was added at a ratio of 1.2-2.3 g of sodium citrate per gram of sludge. The head sludge had a Pb content of 15.0-15.1%. Leaching was carried out at 30-60°C and stirred at 500-700 rpm for 2-4 hours. The results are shown in Table 5. [Table 5]

[0090] [Alkaline leaching] <Examples 20 to 28> Pulps were prepared using sodium hydroxide solutions with concentrations ranging from 5.4 to 8.7% w / w and leached sludge from a continuous copper and lead leaching process with a solid content of 5.0 to 7.0% w / w. The pulps were placed in a 4 L autoclave and heated to a temperature of 100 to 140°C at 600 rpm for 1 to 6 hours. After the leaching time was over, the pulps were cooled and filtered in a Buchner system. The results are shown in Table 6. [Table 6]

[0091] <Examples 29 to 30> A pulp was prepared by adding 420 g of sodium hydroxide and 350 g of leached sludge from a continuous copper and lead leaching process to 6.230 mL of water. This resulted in a concentration of 6.0% w / w NaOH and 5.0% w / w solids. The pulp was placed in a 10 L glass reactor and heated to 90°C and stirred at 900 rpm for 1-6 hours. After the leaching time was completed, the pulp was cooled and filtered in a Buchner system. [Table 7]

[0092] [Hydrogen leaching] <Examples 31 to 38> Solutions with HCl concentrations of 54-160 g / L and chloride concentrations of 140-237 g / L were prepared. The chloride concentration was increased by adding magnesium chloride. 180 g of sludge that had been subjected to sulfuric acid leaching process and citric acid leaching process was added to the solution. Meanwhile, the sludge was subjected to sulfuric acid leaching process, citric acid leaching process, and alkaline leaching process as described in Examples 31-37. The pulp was fed into a 5 L glass reactor, heated at 90°C, and kept under constant stirring for 6 hours. After the pulp test was completed, the pulp was filtered in a Buchner system. The results of these tests are shown in Table 8. [Table 8]

[0093] The results show a clear contribution to copper leaching involving an alkaline leaching step, which allows to increase the copper leaching yield in the overall process. This observation is explained by the presence of chrysocolla in the matrix of the sludge originating from step ii. These are effectively modified by the addition of a silicon-removing base in step ii, leaving the copper more vulnerable to alkaline attack in step 4, as recognized in the results presented herein.

[0094] The final residue from the hydrochloric acid leaching test was subjected to stability testing according to the TCLP and SPLP protocols, which resulted in values ​​of cadmium, arsenic and lead released below the values ​​allowed by the specifications.

[0095] <Example 39> The silver precipitation solution with a magnesium chloride concentration of 300 g / L was subjected to an evaporation process until the broth was concentrated to reach 500 g / L magnesium chloride. 358 mL of the evaporation solution, 385 mL of concentrated hydrochloric acid, and 94 mL of water were added to 274 g of sludge subjected to the continuous leaching process described in Examples 1-28 in a 5 L reactor. The reactor was heated at 90°C and kept under constant stirring for 6 hours. After the pulp test was completed, the pulp was filtered in a Buchner system. The results showed a silver leaching yield of 86%, an iron leaching yield of 95%, and a copper leaching yield of 94%. At least 98% of the leached iron was divalent ferric ion.

[0096] [Silver deposits] <Examples 40 to 46> 450 g of PLS ​​obtained from the hydrochloric acid leaching test was obtained with 508 mg / L Ag in a 600 mL precipitation flask and the PLS was heated at 25-80° C. Neutralization of the silver leaching solution was carried out with 15 vol.% magnesium oxide slurry until the pH reached a range of 3-6. The pulp was then filtered through a 45 μm filter paper. [Table 9]

[0097] [Silver concentrate] <Examples 47 to 49> 50 g of silver precipitate with 3,950 g / ton Ag was placed in a 600 mL precipitation flask and sulfuric acid was added to obtain 60-257 g / L H2SO4. The pulp was stirred using a magnetic bar and subjected to a temperature of 60°C for 5 hours. After the pulp leaching time was completed, the pulp was filtered through 45 μm paper. [Table 10]

[0098] <Examples 50 to 51> 1,700 g of 4,598 g / ton silver precipitate was placed in a 20 ml glass reactor, and 15,300 g of 275 g / L H2SO4 solution was added to the glass reactor. The pulp was kept at 25-80°C and mechanically stirred at 450 rpm for 5 hours. After the leaching time was over, the pulp was filtered through filter paper No. 42. [Table 11]

[0099] [Washing of silver concentrate] <Examples 52 to 55> Due to the increase in the amount of lead in the silver concentrate, a silver concentrate leaching test was performed using sodium citrate. In a 5 L glass reactor, 120 g of silver concentrate obtained from the silver precipitation leaching test and 0.5-1 M sodium citrate solution with pH 7 adjusted using citric acid were added. The pulp was kept at 20-70°C and stirred at 700 rpm for 3 hours. The lead leaching yields ranged from 80-82. The results show that no silver leaching occurs in this step. The Sb leaching yields ranged from 7-10%. [Table 12]

[0100] Quemscan analysis performed on the silver-enriched product revealed the presence of compounds such as silver antimonate and lead antimonate in each of the examples of the present invention. [Brief description of the drawings]

[0101] [Figure 1] 1 shows a process diagram of the method disclosed by the present invention.

Claims

1. A method for producing a silver concentrate from metallurgical residues containing copper, iron, lead, silicon, silver, antimony, arsenic and bismuth, comprising the steps of: i. copper leaching the metallurgical residue with a first acidic solution; a first leaching solution rich in copper, iron and arsenic; obtaining a first leached sludge having reduced copper, iron and arsenic contents and rich contents of lead, germanium, silver and silicon; ii. leaching the first leached sludge with a first solution of a carboxylate salt; a second leached sludge that is lead-depleted; and a second leaching solution rich in lead; iii. Adding a base to generate an alkaline leaching solution, and alkaline leaching the second leached sludge; a third leached sludge having a reduced silicon content; a third leaching solution rich in silicon and arsenic; iv. Hydrogen chloride leaching the third leached sludge using an acidic solution in a chloride environment; a fourth leached sludge for final treatment; a fourth leaching solution enriched in silver, copper, lead, iron and arsenic; v. Precipitating silver from the fourth leaching solution, which is rich in silver, copper, iron and arsenic, using a neutralized slurry; a fifth solution rich in chloride; a first precipitated solid enriched in iron, copper, lead, silver and arsenic; vi. leaching the first precipitated solids enriched in iron, copper, lead, silver and arsenic with a sulfuric acid solution; a sixth leaching solution rich in copper, iron and arsenic; a first silver-lead concentrate; and vii. leaching the first silver-lead concentrate with a second carboxylate solution; A seventh leaching solution; and a second silver concentrate; The method includes:

2. 2. The method of claim 1, wherein the metallurgical residue to be treated is a powder obtained by a metal smelting process.

3. The method of claim 2 , wherein the powder obtained by a copper smelting process is a smelted powder.

4. 4. The method according to claim 1, wherein the metallurgical residue has been subjected to a copper leaching process.

5. 5. The method of claim 4, wherein the metallurgical residue is subjected to leaching with sulfuric acid.

6. The metallurgical residue to be treated contains the mineral species anglesite, coberlin, CuOFe ２ O ３ Cuprospinel in the form of ZnOFe ２ O ３ 4. The method of claim 1, wherein the inorganic filler includes zinc spinel in the form of zinc sulphate, magnetite, iron(III) oxide, pyrite, scorodite, mucovite, kaolinite, and lead(II) sulphate.

7. The copper contained in the metallurgical residue is present in the form of copper sulfate, chalcone, cobalt, and CuOFe. ２ O ３ 7. The method of claim 6, wherein the cuprospinel is in the form of cuprospinal.

8. 8. The method according to claim 1, wherein the silicon contained in the metallurgical residue is present as muscovite and kaolinite.

9. 9. 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.

10. 10. The method of claim 9, wherein at least 95% of the lead is found as lead (II) sulfate.

11. A method according to any one of claims 1 to 10, wherein the first acidic solution may contain sulfuric acid and / or refinery waste liquid.

12. The method according to claim 11, wherein step (i) is carried out at a sulfuric acid concentration of 150 to 300 g / L.

13. The method according to any one of claims 1 to 12, wherein step (i) is carried out at a temperature of from 50 to 130°C.

14. 14. The method of any one of claims 1 to 13, wherein step (i) is carried out for a period of from 3 to 12 hours.

15. 15. A method according to any one of the preceding claims, wherein step (i) is carried out such that the metallurgical residue has a solids concentration of 5-20% w / w.

16. 16. The method of any one of claims 1 to 15, wherein the carboxylate in step (ii) of leaching is sodium citrate.

17. 17. The method of claim 16, wherein in step (ii) the sodium citrate solution has a molarity of sodium citrate of 0.5 to 1 M.

18. 17. The method of claim 16, wherein in step (ii), the first leached sludge is fed to the sodium citrate solution in a mass ratio of first leached sludge:sodium citrate solution of 1:

9.

19. The method according to any one of claims 1 to 18, wherein step (ii) is carried out at a temperature of from 20 to 60°C.

20. 20. The method of any one of claims 1 to 19, wherein step (ii) is carried out for a residence time of from 1 to 23 hours.

21. 21. The method of any one of claims 1 to 20, wherein step (ii) is carried out at a pH of 5.3 to 8.

8.

22. 22. The method of any one of claims 1 to 21, wherein in step (ii) citric acid is added to adjust the pH.

23. 23. The method of claim 22, wherein the pH adjustment in step (ii) is carried out using a citric acid solution of 600 and 900 g / L.

24. The base used in the leaching in step (iii) is Mg(OH) ２ 24. The method of claim 1, wherein the aqueous solution is selected from the group consisting of KOH, and NaOH.

25. 25. The method according to any one of claims 1 to 24, wherein the base added in step (iii) is added in a ratio of 5 to 10% w / w relative to the total mass of the alkaline leaching solution.

26. 26. The method according to any one of claims 1 to 25, wherein the leaching reaction in step (iii) is carried out at a temperature of 90 to 140°C.

27. 27. The method according to any one of claims 1 to 26, wherein the leaching reaction in step (iii) is carried out for a residence time of from 1 to 6 hours.

28. 28. The method of any one of claims 1 to 27, wherein the acid used for leaching in step (iv) is hydrochloric acid.

29. 29. The method of claim 28, wherein in step (iv), hydrochloric acid is provided at a concentration of 50 to 140 g / L.

30. 30. The method of claim 29, wherein in step (iv) the chloride environment is increased by addition of a chloride salt.

31. 31. The method of claim 30, wherein in step (iv) the chloride environment is increased by the addition of magnesium chloride.

32. 32. The method of claim 30 or 31, wherein in step (iv) chloride is provided at a concentration of 140 to 240 g / L.

33. 33. The method of any one of claims 1 to 32, wherein step (iv) is carried out at a pH of -1.5 to -0.

25.

34. 34. The method of any one of claims 1 to 33, wherein step (iv) is carried out at a temperature of 40 to 95°C.

35. 35. The method of any one of claims 1 to 34, wherein the neutralized slurry in silver deposition step (v) is selected from calcium hydroxide, calcium oxide, calcium carbonate, lime, dolomitic lime, magnesium carbonate, magnesium hydroxide, or magnesium oxide.

36. 36. The method of claim 35, wherein the neutralized slurry in silver deposition step (vi) is a magnesium oxide slurry.

37. 37. The method of any one of claims 1 to 36, wherein step (v) is carried out at a temperature of 50 to 95°C.

38. 38. The method of any one of claims 1 to 37, wherein the neutralized slurry added in step (v) is fed until a pH of 3 to 7 is reached.

39. 39. The method of any one of claims 1 to 38, wherein step (v) has a residence time of from 0.5 to 3 hours.

40. 40. The method of any one of claims 1 to 39, wherein the fifth chloride-rich solution in step (v) is sent to a magnesium chloride crystallization process.

41. 40. The method of any one of claims 1 to 39, wherein the fifth chloride-rich solution in step (v) is recycled to step (iv).

42. The method of any one of claims 1 to 41, wherein the sulfuric acid solution in step (vi) has a sulfuric acid concentration of 60 to 275 g / L.

43. 43. The method according to any one of claims 1 to 42, wherein the sulfuric acid solution in step (vi) is the first leaching solution in step (i) whose acidity has been adjusted to 60 to 275 g / L and which is rich in copper, iron, arsenic and bismuth.

44. 44. The method of any one of the preceding claims, wherein the step (vi) of leaching the silver precipitate is carried out at a temperature of from 50 to 95°C.

45. 45. The method of any one of claims 1 to 44, wherein the carboxylate in step (vii) of leaching the first silver-lead concentrate is sodium citrate.

46. 46. ​​The method of claim 45, wherein the concentration of sodium citrate in step (vii) is 0.5 to 1 M.

47. 47. The method of any one of claims 1 to 46, wherein the step (vii) of leaching the first silver-lead concentrate is carried out at a temperature of 25 to 70°C.

48. 48. The method of any one of the preceding claims, wherein step (vii) of leaching the first silver-lead concentrate is carried out such that the first silver-lead concentrate has a solids content of 5-10% w / w.

49. 49. The method of any one of claims 1 to 48, wherein the step (vii) of leaching the first silver-lead concentrate is carried out for a period of 1 to 6 hours.

50. 50. The method of any one of claims 1 to 49, wherein the seventh leaching solution in step (vii) is recycled to the leaching in step (ii).

51. 51. The method of any one of claims 1 to 50, wherein the second silver concentrate comprises silver antimonate.

52. The method of any one of claims 1 to 51, wherein the second silver concentrate is comprised of silver antimonate and lead antimonate.

53. 53. The method of any one of claims 1 to 52, wherein the copper-rich first leaching solution is delivered to a copper leaching process of smelting powder.

54. 53. The method of any one of claims 1 to 52, wherein the first leaching solution rich in copper is sent to an arsenic removal process.

55. 55. The method of claim 54, wherein the arsenic removal process is a scorodite production process.

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