Method for recovering precious metal from high platinum group metal copper anode slime by wet process

Abstract

The invention discloses a method for wet recovery of precious metals from high platinum group metal copper anode slime, which comprises the following steps: heating and reacting the high platinum group metal copper anode slime in the presence of dilute sulphuric acid and air, and after the reaction is finished, carrying out solid-liquid separation to obtain a first filtrate and a first filter cake; dispersing and diluting the first filter cake into first slurry by adopting a sulfuric acid solution, heating the first slurry in an oxygen atmosphere, carrying out high-pressure reaction at the reaction temperature of not higher than 170 DEG C, and after the reaction is finished, carrying out solid-liquid separation to obtain second filtrate and a second filter cake; preparing the second filter cake into second slurry by adopting a sulfuric acid solution, heating to 90-110 DEG C, adding sodium chlorate in batches, distilling, absorbing ruthenium in distilled flue gas by adopting a hydrochloric acid solution, absorbing osmium in the flue gas by adopting sodium hydroxide, and recovering osmium and ruthenium from distillate; and recovering gold, platinum, palladium, rhodium, iridium, selenium and tellurium from the residual distilled slurry.

Description

Technical Field

[0001] The invention belongs to the technical field of copper anode mud treatment, and particularly relates to a method for wet-process recovery of precious metals from high-platinum group metal copper anode mud. Background Art

[0002] The statements herein merely provide background information related to the present invention and do not necessarily constitute prior art.

[0003] Copper anode mud is a by-product produced during the copper electrolytic refining process. Its main components are metals and their compounds that are insoluble in the electrolyte.

[0004] Currently, pyrometallurgical processes are widely used to treat copper anode slime. These processes typically include roasting for selenium removal, acid leaching for copper removal, reduction smelting, precious lead oxidation refining, and silver and gold electrolysis. These processes offer advantages such as high throughput and simple operation, making them popular with large-scale copper smelters. However, pyrometallurgical treatment processes have numerous drawbacks: high slag return, low direct gold and silver recovery rates, long production cycles, and significant backlogs of precious metals.

[0005] Furthermore, when using pyrometallurgical methods to treat high-PGM anode slime, secondary PGMs such as rhodium, iridium, and ruthenium cannot be dissolved in aqua regia or chlorination systems after being exposed to high temperatures, making subsequent processing difficult. Furthermore, secondary PGMs such as ruthenium and osmium volatilize into the flue gas at high temperatures, making them difficult to recover and resulting in the loss of valuable metals. Summary of the Invention

[0006] In view of the shortcomings of the prior art, the object of the present invention is to provide a method for wet recovery of precious metals from high-platinum group metal copper anode mud.

[0007] In order to achieve the above object, the present invention is implemented through the following technical solutions:

[0008] A method for wet recovery of precious metals from high-platinum group metal copper anode mud comprises the following steps:

[0009] The high-platinum group metal copper anode mud is heated to react in the presence of dilute sulfuric acid and air, and after the reaction is completed, the solid and liquid are separated to obtain a first filtrate and a first filter cake;

[0010] The first filter cake is dispersed and diluted with a sulfuric acid solution to form a first slurry. The first slurry is heated in an oxygen atmosphere and reacted under high pressure at a reaction temperature not exceeding 170° C. After the reaction is completed, the solid and liquid are separated to obtain a second filtrate and a second filter cake.

[0011] The second filter cake is prepared into a second slurry with a sulfuric acid solution, heated to 90-110° C., sodium chlorate is added in batches, and distillation is performed. Ruthenium in the distillation flue gas is absorbed with a hydrochloric acid solution, osmium in the flue gas is absorbed with sodium hydroxide, and osmium and ruthenium are recovered from the distillate;

[0012] Gold, platinum, palladium, rhodium, iridium, selenium and tellurium can be recovered from the remaining distillation slurry.

[0013] The purpose of pressure leaching is to separate base metals from precious metals. The base metals are leached and oxidized into soluble sulfates that enter the solution, while the precious metals remain in the insoluble residue. When the temperature is above 170°C and under pressure, some of the precious metals will be oxidized and dissolved into the solution, resulting in precious metal loss.

[0014] In some embodiments, the high-platinum group metal copper anode slime is heated in the presence of dilute sulfuric acid and air at a reaction temperature of 70-80° C. for 4-6 hours. Under these reaction conditions, a portion of the copper in the copper anode slime is oxidized and dissolved.

[0015] In some embodiments, the first filtrate is subjected to a copper electrolysis process to purify the copper.

[0016] In some embodiments, the concentration of sulfuric acid in the first slurry is 350-370 g / L.

[0017] In some embodiments, the pressure of the heated, high-pressure reaction is 0.5 to 1 MPa, preferably 0.7 to 0.9 MPa.

[0018] In some embodiments, the second filter cake is washed and blown, and then prepared into a second slurry using a sulfuric acid solution.

[0019] Preferably, the washing and blowing is performed using compressed air to remove moisture from the filter cake and dry the filter cake.

[0020] In some embodiments, the second filtrate is heated to 90-100° C., sulfur dioxide is introduced, and copper powder is added to perform selenium and tellurium precipitation.

[0021] The ruthenium and osmium vapors volatilized by oxidation pass through the ruthenium absorption system and the osmium absorption system in turn. The scrubbing liquid of the ruthenium absorption system is mainly hydrochloric acid solution, which recovers ruthenium in the flue gas. The following reactions mainly occur:

[0022] 2RuO4+20HCl=2H2RuCl4+8H2O+4Cl2;

[0023] The scrubbing liquid of the osmium absorption system is mainly a liquid alkali solution. The following reactions mainly occur when osmium in the flue gas is recovered:

[0024] 2OsO4+4NaOH=2Na2OsO4+2H2O+O2.

[0025] In some embodiments, the specific method for recovering osmium ruthenium from the distillate is:

[0026] Ruthenium Recovery: The hydrochloric acid absorption solution of ruthenium will contain some OsO4, so the osmium must be removed first. The hydrochloric acid absorption solution of ruthenium is placed in a distiller, and the exhaust pipe is connected to the osmium absorption system filled with liquid alkali solution. The ruthenium absorption solution is heated to boiling, causing the OsO4 in the ruthenium absorption solution to evaporate and be absorbed by the alkaline absorption solution. The solution after removing the osmium is further concentrated, and solid ammonium chloride is added while it is hot, producing the following reaction:

[0027] H2RuCl4+2NH4Cl=(NH4)2RuCl6↓+2HCl;

[0028] The ammonium hexachlororuthenate is precipitated. After the precipitation is complete, it is cooled, filtered, washed, dried, and then calcined and hydrogen reduced to obtain the finished ruthenium powder.

[0029] Recovery of osmium: Cool the osmium absorption solution, add solid ammonium chloride to the solution while it is cold and stirring, and the following reaction occurs:

[0030] Na2OsO4+4NH4Cl=[OsO2(NH3)4]Cl2↓+2NaCl+2H2O;

[0031] Tetraammineosmium anhydride dichloride is a precipitate, which is filtered, dried, and then subjected to hydrogen reduction, washed, and dried to obtain a finished osmium powder.

[0032] In some embodiments, chlorine gas is introduced into the remaining distilled slurry to perform chlorination leaching and solid-liquid separation to obtain a third filtrate and a third filter cake;

[0033] The third filter cake is reduced, cast and electrolyzed to obtain silver powder;

[0034] Add barium chloride to the third filtrate to precipitate barium selenate precipitate, and the filtrate is a gold, platinum, palladium, rhodium and iridium solution.

[0035] Preferably, the barium selenate precipitate is replaced by sulfuric acid to obtain barium sulfate precipitate, and the filtrate is reduced by sulfur dioxide to obtain crude selenium.

[0036] Preferably, the gold, platinum, palladium, rhodium and iridium solution is precipitated with sodium bisulfite to recover gold;

[0037] The post-immersion gold solution uses extractant S 201 (Diisoamyl sulfide) is used to extract palladium, and the palladium strip solution is deoiled, concentrated acid, ammonia complexed, and acidified to form crude dichlorodiamine palladium. The crude dichlorodiamine palladium is purified and reduced with hydrazine hydrate to obtain sponge palladium.

[0038] More preferably, the palladium extraction solution is treated with ammonium chloride to precipitate platinum, which is then purified by reacting hydrochloric acid and ammonium chloride to obtain pure ammonium hexachloroplatinate, which is then calcined to obtain sponge platinum.

[0039] More preferably, zinc and magnesium powders are added to the platinum precipitate to reduce it to obtain rhodium-iridium concentrate, the rhodium-iridium concentrate is subjected to chlorination leaching, iridium is extracted using TAPO (trialkylphosphine oxide), the extract is washed and stripped, and refined to obtain iridium powder; the raffinate is decontaminated, concentrated, and refined to obtain rhodium powder.

[0040] The beneficial effects achieved by one or more embodiments of the present invention are as follows:

[0041] The present invention uses atmospheric pressure acid leaching and high pressure oxidation leaching to leach copper, some selenium, and tellurium from the copper anode mud into solution for separate recovery. The copper-removed mud is then digested with sulfuric acid to recover valuable metals such as osmium and ruthenium. Chlorination leaching then proceeds to chlorinate gold, platinum, palladium, ruthenium, rhodium, and selenium into solution. Silver is recovered as silver chloride. Selenium in the solution is recovered using barium chloride to remove selenium. The remaining precious metals are then reduced or extracted for separate recovery.

[0042] Osmium and ruthenium are recovered using sulfate digestion technology. Chloride leaching technology chlorinates gold, platinum, palladium, ruthenium, rhodium, and selenium into a water solution, and barium chloride is used to recover selenium. Precious metals are recovered separately using fractional reduction or extraction technology. BRIEF DESCRIPTION OF THE DRAWINGS

[0043] The accompanying drawings, which constitute a part of the present invention, are used to provide a further understanding of the present invention. The exemplary embodiments of the present invention and their descriptions are used to explain the present invention and do not constitute improper limitations on the present invention.

[0044] Figure 1 This is a process flow chart for wet recovery of precious metals from copper anode mud according to an embodiment of the present invention. DETAILED DESCRIPTION

[0045] It should be noted that the following detailed description is illustrative and is intended to provide further explanation of the present invention. Unless otherwise specified, all technical and scientific terms used in the present invention have the same meaning as commonly understood by those skilled in the art to which the present invention belongs.

[0046] The present invention is described in detail below with reference to the embodiments.

[0047] Example 1

[0048] A method for wet recovery of precious metals from high-platinum group metal copper anode mud comprises the following steps:

[0049] 1. Atmospheric pressure leaching

[0050] In the prepreg tank, add 5t copper anode mud to 20m 3In a dilute sulfuric acid solution (pre-soak solution with a sulfuric acid concentration of 100g / l), the reaction temperature is set at 75°C under mechanical stirring, steam heating, and air bubbling to partially oxidize and dissolve the copper. Following the reaction, the filtrate is filtered and returned to the copper electrolysis system. The filter cake enters a slurry tank, where water and sulfuric acid are added to a concentration of 350g / l before being pumped into an autoclave.

[0051] 2. High pressure leaching

[0052] The anode slurry is pressurized and leached in an autoclave. The oxidant used is process oxygen with a flow rate of 40 Nm 3 / h, leaching pressure 0.86Mpa, temperature not higher than 170℃. After high-pressure leaching, the slurry enters the transfer tank. The slurry is filtered through a filter press. After washing and blowing, the filter cake (copper-removed mud) is used to recover precious metals, and the filtrate is stored in a selenium-tellurium precipitation tank.

[0053] 3. Selenium and tellurium precipitation

[0054] In a selenium-tellurium precipitation tank, heated to 100°C, sulfur dioxide and copper powder are used to reduce and precipitate selenium-tellurium. The selenium-tellurium filter cake is obtained by pressure filtration and the selenium-tellurium is recovered. The filtrate is returned to the electrolysis workshop.

[0055] 4. Osmium-Ruthenium Distillation

[0056] The copper-removed filter cake is mixed with sulfuric acid and placed in a distillation kettle. The mixture is heated to 100°C and sodium chlorate is slowly added for distillation. Hydrochloric acid is used to absorb ruthenium from the flue gas, while sodium hydroxide is used to absorb osmium from the flue gas. The distillate is regularly processed to recover osmium and ruthenium.

[0057] The ruthenium and osmium vapors volatilized by oxidation pass through the ruthenium absorption system and the osmium absorption system in turn. The scrubbing liquid of the ruthenium absorption system is mainly hydrochloric acid solution, which recovers ruthenium in the flue gas. The following reactions mainly occur:

[0058] 2RuO4+20HCl=2H2RuCl4+8H2O+4Cl2;

[0059] The scrubbing liquid of the osmium absorption system is mainly a liquid alkali solution. The following reactions mainly occur when osmium in the flue gas is recovered:

[0060] 2OsO4+4NaOH=2Na2OsO4+2H2O+O2.

[0061] In some embodiments, the specific method for recovering osmium ruthenium from the distillate is:

[0062] Ruthenium Recovery: The hydrochloric acid absorption solution of ruthenium will contain some OsO4, so the osmium must be removed first. The hydrochloric acid absorption solution of ruthenium is placed in a distiller, and the exhaust pipe is connected to the osmium absorption system filled with liquid alkali solution. The ruthenium absorption solution is heated to boiling, causing the OsO4 in the ruthenium absorption solution to evaporate and be absorbed by the alkaline absorption solution. The solution after removing the osmium is further concentrated, and solid ammonium chloride is added while it is hot, producing the following reaction:

[0063] H2RuCl4+2NH4Cl=(NH4)2RuCl6↓+2HCl;

[0064] The ammonium hexachlororuthenate is precipitated. After the precipitation is complete, it is cooled, filtered, washed, dried, and then calcined and hydrogen reduced to obtain the finished ruthenium powder.

[0065] Recovery of osmium: Cool the osmium absorption solution, add solid ammonium chloride to the solution while it is cold and stirring, and the following reaction occurs:

[0066] Na2OsO4+4NH4Cl=[OsO2(NH3)4]Cl2↓+2NaCl+2H2O;

[0067] Tetraammineosmium anhydride dichloride is a precipitate, which is filtered, dried, and then subjected to hydrogen reduction, washed, and dried to obtain a finished osmium powder.

[0068] 5. Chlorination leaching

[0069] After the distillation is completed, chlorine gas is introduced into the distilled material for chlorination leaching, and metals such as gold, platinum, palladium, rhodium, iridium, selenium, and tellurium are chlorinated into the solution, and silver forms silver chloride and precipitates into the filter cake.

[0070] Silver chloride is reduced with iron powder, and silver anodes are cast for electrolysis to obtain qualified silver powder, which is then cast into silver ingots. The anode mud is silver anode mud, which is then leached by chlorination.

[0071] Barium chloride is added to the filtrate to form a barium selenate precipitate. Filter the filtrate to obtain a gold, platinum, palladium, rhodium, and iridium solution. The filter cake, barium selenate, is displaced with sulfuric acid and filtered to obtain a barium sulfate cake. Sulfur dioxide is then introduced into the filtrate to reduce the solution and obtain crude selenium.

[0072] 6. Gold, platinum, palladium, rhodium and iridium solution treatment

[0073] 6.1 Gold Recovery

[0074] When the gold, platinum, palladium, rhodium and iridium solution is heated to 55-60℃, sodium bisulfite is used as a reducing agent to precipitate gold to obtain primary gold powder, which can be cast into gold ingots. It is further reduced to secondary gold powder and then leached by chlorination.

[0075] 6.2 Palladium Recovery

[0076] The post-immersion gold solution is mainly platinum palladium rhodium iridium solution, using extraction method, using extraction agent S 201(diisoamyl sulfide) to extract the palladium and further purify it.

[0077] The palladium stripping solution is deoiled, concentrated acid, ammonia complexed and acidified to form crude dichlorodiamine palladium. After multiple dissolutions and precipitations, pure dichlorodiamine palladium is obtained, which is then reduced with hydrazine hydrate to obtain sponge palladium.

[0078] The specific method is:

[0079] The organic phase (containing 25% S 201 The kerosene (kerosene as a diluent) was mixed with the palladium-containing liquid at a ratio (O / A) of 3:1 and stirred at room temperature for 10 minutes. The palladium was selectively extracted into the organic phase, as shown in the following reaction formula:

[0080] PdCl4 2- +2S 201 →[PdCl2·2S 201 ]PdCl4 2- +2S 201 →[PdCl2·2S 201 ];

[0081] Phase separation, standing for 20 minutes, separate the loaded organic phase (containing Pd) and the raffinate (containing other impurities. Wash the loaded organic phase with 1 mol / L HCl to remove a small amount of Pt, Fe and other impurities that are co-extracted. Then use 0.8 mol / L ammonia water to extract palladium, mix at room temperature for 5 minutes, and compare O / A = 1:2. Palladium is converted to Pd(NH3)4 2+ The form enters the water phase and the following reactions occur:

[0082] [PdCl2·2S 201 ]+4NH3→Pd(NH3)4 2+ +2Cl - +2S 201 ;

[0083] Organic phase (containing 25% S 201 The organic phase after stripping is washed with dilute hydrochloric acid (1 mol / L) to remove residual ammonia or thiourea, and then it can be recycled.

[0084] Ammonia stripping solution: Pd(NH3)2Cl2 precipitate can be generated by acidification (adjusting pH to 1) or adding hydrochloric acid.

[0085] The crude dichlorodiammine palladium is dissolved in ammonia water several times, acidified and precipitated to remove impurities, and pure dichlorodiammine palladium precipitate is obtained, which is then reduced by adding hydrazine hydrate to obtain a sponge palladium product.

[0086] 6.3 Platinum Recovery

[0087] The palladium extraction solution is treated with ammonium chloride to precipitate platinum, which is then purified by repeated ammonium chloride precipitation and reaction with hydrochloric acid and ammonium chloride to obtain pure ammonium hexachloroplatinate, which is then calcined to obtain sponge platinum.

[0088] The specific steps are: after palladium extraction, platinum in the liquid is mainly (PtCl6) 2- It exists in the form of (hexachloroplatinate). At room temperature, after adding 1.5 times the amount of NH4Cl to the palladium extraction solution according to the amount of Pt, a yellow (NH4)2PtCl6 precipitate is generated, while other metals (such as Rh, etc.) usually remain in the solution, and the following reaction occurs:

[0089] PtCl6 2- +2NH4 + →(NH4)2PtCl6↓;

[0090] Stir until completely dissolved, and a yellow (NH4)2PtCl6 precipitate gradually forms in the solution. Allow to react for 2 hours to allow the precipitation to complete. Filter and wash the precipitate with saturated NH4Cl solution or 10% HCl to remove adsorbed impurities such as Rh.

[0091] Dissolve the (NH4)2PtCl6 precipitate with a small amount of dilute ammonia (5% NH4OH) and the following reaction occurs:

[0092] (NH4)2PtCl6+2NH3→Pt(NH3)4 2+ +2NH4 + +6Cl - ;

[0093] Add concentrated HCl dropwise to the solution until the pH is about 1 to regenerate the (NH4)2PtCl6 precipitate. Repeat the above process twice to improve the purity. Filter and dry to obtain a pure (NH4)2PtCl6 precipitate.

[0094] Calcinate (NH4)2PtCl6 at 500-600℃ and decompose it into sponge platinum product, and the following reaction occurs:

[0095] (NH4)2PtCl6→Pt+2NH4Cl+2Cl2↑.

[0096] 6.4 Rhodium and Iridium Recovery

[0097] Zinc and magnesium powders are added to the platinum precipitate to reduce it to obtain rhodium-iridium concentrate. The rhodium-iridium concentrate is leached with chloride, and iridium is extracted using TAPO (trialkylphosphine oxide). The extract is washed and stripped, and refined to obtain iridium powder. The raffinate is decontaminated, concentrated, and refined to obtain rhodium powder.

[0098] The solution after platinum precipitation (mainly containing (RhCl6) 3- 、(IrCl6) 2-And a small amount of Pt, Pd, etc., use zinc powder or magnesium powder at 1.5 times the theoretical amount to add, the following reaction occurs:

[0099] 2RhCl6 3- +3Zn→2Rh↓+3Zn 2+ +12Cl - ;

[0100] IrCl6 2- +Zn→Ir↓+Zn 2+ +6Cl - .

[0101] Stir at 90°C for 1 hour to completely reduce Rh and Ir to metal powder. Filter and wash with dilute HCl (1 mol / L) to obtain Rh-Ir concentrate (containing small amounts of Pt, Pd, Cu, etc.).

[0102] Chloride leaching of Rh-Ir concentrate, oxidation by chlorine gas in HCl medium, reaction at 90℃ for 2 hours, converting Rh and Ir into [RhCl6] 3- and [IrCl6] 2- .

[0103] TAPO extraction separation of iridium, using kerosene containing 10-20% TAPO as the extractant, O / A = 1:1, stirring at room temperature for 10 minutes. TAPO preferentially extracts [IrCl6] 2- , Rh remains in the aqueous phase and the following reaction occurs:

[0104] IrCl6 2- +2TAPO→[TAPO·H2IrCl6];

[0105] The loaded organic phase was washed with 3 mol / L HCl to remove a small amount of co-extracted Rh and Pt.

[0106] Stripping iridium, using 1mol / L NaOH solution to strip iridium, the following reaction occurs:

[0107] [TAPO·H2IrCl6]+2OH - →IrO2·nH2O↓+6Cl - +2TAPO;

[0108] After filtration, the Ir precipitate is dissolved with HCl and refined to obtain H2IrCl6 solution, which is further reduced to iridium powder.

[0109] The raffinate is treated to prepare rhodium powder, and the raffinate (containing RhCl6 3-) is added with NH₄Cl to precipitate residual (NH₄)₂PtCl₆ or Pd(NH₃)₂Cl₂, which is removed by filtration. The mixture is concentrated by evaporation, and NH₄Cl is added to form red crystals of (NH₄)₃RhCl₆. (NH₄)₃RhCl₆ is calcined at 600°C to obtain rhodium sponge.

[0110] The recovery rates of various metal elements in copper anode mud are as follows:

[0111] Selenium: 90%; Tellurium: 90%; Osmium: 96.5%; Ruthenium: 98%; Gold: 99.3%; Platinum: 98.5%; Palladium: 98.5%; Rhodium: 98%; Iridium: 98%; Silver: 99%.

[0112] The foregoing description is merely a preferred embodiment of the present invention and is not intended to limit the present invention. Those skilled in the art will readily appreciate that various modifications and variations of the present invention are possible. Any modifications, equivalent substitutions, or improvements made within the spirit and principles of the present invention are intended to be within the scope of protection of the present invention.

Claims

1. A method for wet recovery of precious metals from high-platinum group metal copper anode mud, characterized by: The steps include: The high-platinum group metal copper anode mud is heated to react in the presence of dilute sulfuric acid and air, and after the reaction is completed, the solid and liquid are separated to obtain a first filtrate and a first filter cake; The first filter cake is dispersed and diluted with a sulfuric acid solution to form a first slurry. The first slurry is heated in an oxygen atmosphere and reacted under high pressure at a reaction temperature not exceeding 170° C. After the reaction is completed, the solid and liquid are separated to obtain a second filtrate and a second filter cake. The second filter cake is prepared into a second slurry with a sulfuric acid solution, heated to 90-110° C., sodium chlorate is added in batches, and distillation is performed. Ruthenium in the distillation flue gas is absorbed with a hydrochloric acid solution, osmium in the flue gas is absorbed with sodium hydroxide, and osmium and ruthenium are recovered from the distillate; Gold, platinum, palladium, rhodium, iridium, selenium and tellurium can be recovered from the remaining distillation slurry.

2. The method for wet recovery of precious metals from high-platinum group metal copper anode slime according to claim 1, characterized in that: The high platinum group metal copper anode mud is heated in the presence of dilute sulfuric acid and air to react at a temperature of 70 to 80° C. for 4 to 6 hours.

3. The method for wet recovery of precious metals from high-platinum group metal copper anode slime according to claim 1, characterized in that: In the first slurry, the sulfuric acid concentration is 350-370 g / L.

4. The method for wet recovery of precious metals from high-platinum group metal copper anode slime according to claim 1, characterized in that: The pressure of the heating and high-pressure reaction is 0.5 to 1 MPa, preferably 0.7 to 0.9 MPa.

5. The method for wet recovery of precious metals from high-platinum group metal copper anode slime according to claim 1, characterized in that: After the second filter cake is washed and blown, a sulfuric acid solution is used to prepare a second slurry.

6. The method for wet recovery of precious metals from high-platinum group copper anode slime according to claim 5, characterized in that: The washing and blowing is to use compressed air to take away the moisture in the filter cake and blow the filter cake dry.

7. The method for wet recovery of precious metals from high-platinum group metal copper anode slime according to claim 1, characterized in that: The second filtrate is heated to 90-100°C, sulfur dioxide is introduced, copper powder is added, and selenium and tellurium precipitation is carried out.

8. The method for wet recovery of precious metals from high-platinum group copper anode slime according to claim 1, characterized in that: Chlorine gas is introduced into the remaining distilled slurry to perform chlorination leaching and solid-liquid separation to obtain a third filtrate and a third filter cake; The third filter cake is reduced, cast and electrolyzed to obtain silver powder; Add barium chloride to the third filtrate to precipitate barium selenate precipitate, and the filtrate is a gold, platinum, palladium, rhodium and iridium solution.

9. The method for wet recovery of precious metals from high-platinum group copper anode slime according to claim 1, characterized in that: The barium selenate precipitate is replaced by sulfuric acid to obtain barium sulfate precipitate, and the filtrate is reduced by sulfur dioxide to obtain crude selenium; Preferably, the gold, platinum, palladium, rhodium and iridium solution is precipitated with sodium bisulfite to recover gold; The gold precipitation liquid is extracted with palladium using diisoamyl sulfide as an extractant. The palladium strip solution is deoiled, concentrated acid, ammonia complexed, and acidified to form crude dichlorodiamine palladium. The crude dichlorodiamine palladium is purified and reduced with hydrazine hydrate to obtain sponge palladium.

10. The method for wet recovery of precious metals from high-platinum group copper anode slime according to claim 9, characterized in that: The palladium extraction solution is treated with ammonium chloride to precipitate platinum, which is then purified by reacting nitric acid, hydrochloric acid and ammonium chloride to obtain pure ammonium hexachloroplatinate, which is then calcined to obtain sponge platinum. Preferably, zinc and magnesium powders are added to the platinum precipitate to reduce the rhodium-iridium concentrate, the rhodium-iridium concentrate is leached by chlorination, iridium is extracted using TAPO, the extract is washed and stripped, and iridium powder is refined; The raffinate is decontaminated, concentrated, and refined to obtain rhodium powder.

Images