Treatment process of high platinum group metal copper anode slime

Abstract

The invention discloses a high platinum group metal copper anode slime treatment method which comprises the following steps: placing high platinum group metal copper anode slime in dilute sulphuric acid, blowing air, heating, oxidizing and leaching copper and nickel in the high platinum group metal copper anode slime, and carrying out solid-liquid separation to obtain a first filtrate and a first filter cake; dispersing the first filter cake in dilute sulfuric acid, carrying out heating and pressurizing reaction in an oxygen atmosphere, dissolving out part of metal in the first filter cake, carrying out solid-liquid separation to obtain a second filtrate and a second filter cake, mixing the second filter cake, calcium oxide and coke powder, pressing into blocks, smelting at 1450-1550 DEG C for a set time, atomizing the smelted alloy to obtain fine particles, and recycling the fine particles. Gold-platinum-palladium alloy powder is obtained; and after the second filtrate is subjected to flash evaporation and rapid cooling, ferrous sulfate is added into the second filtrate for palladium precipitation, the filtrate obtained through filtering is sequentially subjected to the steps of silver precipitation and selenium and tellurium precipitation, thiourea is added into the remaining filtrate, a reaction is conducted at high temperature and high pressure, thiourea complex precipitation is generated, solid-liquid separation is conducted, third filtrate and a third filter cake are obtained, and platinum family concentrate is obtained.

Description

Technical Field

[0001] The invention belongs to the technical field of copper anode mud treatment, and in particular relates to a treatment process for 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] In the copper smelting industry, copper anode slime is a byproduct of the electrolytic refining and purification process of anode copper. It is rich in valuable metals such as gold, silver, and platinum group metals (platinum, palladium, and rhodium). PGMs (platinum, palladium, and rhodium) are generally present in low concentrations in copper concentrate, and the amount of these metals concentrated in copper anode slime is less than 1%, or even lower, representing trace amounts. Through the copper anode slime treatment process, these metals are continuously enriched into platinum and palladium concentrate for refining and recovery.

[0004] Currently, the most common method for treating copper anode slime is pyrometallurgy. This involves deep processing of the anode slime through high-temperature electrolysis to separate the different metallic elements. This method, with its advantages of high throughput and simple operation, is widely adopted by large-scale copper smelters. However, pyrometallurgy has several drawbacks: high slag return, low direct recovery of gold and silver, long production cycles, and a significant backlog of precious metals.

[0005] Fire treatment of high-PGM anode slime has its drawbacks. Some secondary PGMs (such as rhodium, iridium, and ruthenium) cannot be dissolved in aqua regia or chlorination systems after exposure to high temperatures, making subsequent processing difficult. Some secondary PGMs (such as ruthenium and osmium) volatilize into the flue gas at high temperatures, resulting in the loss of valuable metals. Currently, there is no comprehensive and comprehensive treatment process for anode slime containing high PGMs (such as platinum, palladium, ruthenium, rhodium, and iridium). Summary of the Invention

[0006] In view of the shortcomings of the existing technology, the purpose of the present invention is to provide a method for treating high-platinum group metal copper anode mud. The anode mud treatment method provided by the present invention can make the high-platinum group metal copper anode mud treatment process reasonable, ensure the high recovery rate of metals such as ruthenium, rhodium, and iridium, and simplify the process, which is conducive to the recovery of valuable metals in the anode mud, simplifies the recovery operation process, and reduces the loss of rare and precious metals.

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

[0008] A method for treating high-platinum group metal copper anode mud comprises the following steps:

[0009] Placing high-platinum group metal copper anode mud in dilute sulfuric acid, blowing air into it, and heating it to oxidize and leach the copper and nickel therein, and separating the solid and liquid to obtain a first filtrate and a first filter cake;

[0010] The first filter cake is dispersed in dilute sulfuric acid, heated and pressurized in an oxygen atmosphere to react, dissolving part of the metal therein, and performing solid-liquid separation to obtain a second filtrate and a second filter cake. The second filter cake, calcium oxide, and coke powder are mixed, pressed into blocks, and smelted at 1450-1550°C for a set time. The smelted alloy is atomized to obtain fine particles for recovery, thereby obtaining gold-platinum-palladium alloy powder.

[0011] After the second filtrate is flash evaporated and rapidly cooled (to prevent the volatilization of ruthenium and the loss of valuable metals), ferrous sulfate is added thereto to precipitate palladium. The filtered filtrate is sequentially subjected to the steps of silver precipitation and selenium and tellurium precipitation. Thiourea is added to the remaining filtrate and reacted under high temperature and high pressure to generate a thiourea complex precipitate. The solid and liquid are separated to obtain a third filtrate and a third filter cake, thereby obtaining a platinum group concentrate.

[0012] In some embodiments, the high-platinum group metal copper anode slime is heated in dilute sulfuric acid at a temperature of 70-80° C., and the reaction time is 3-5 hours.

[0013] In some embodiments, the temperature of the heating and pressurizing reaction is 180-200° C., the oxygen pressure is 0.4-0.6 MPa, and the reaction time is 5-7 hours.

[0014] Preferably, the heating and pressurizing reaction is carried out in a continuous horizontal reactor, which has a horizontal structure and is divided into 5-7 reaction chambers by overflow partitions. Each reaction chamber is provided with an agitator and an oxygen inlet distributor, and the reaction system overflows from one side of the horizontal continuous autoclave in sequence until it reaches the outlet.

[0015] The inventors have discovered through experiments that, compared to single-chamber autoclaves, this continuous horizontal reactor offers greater leaching selectivity and a more concentrated metal distribution. Copper, nickel, silver, palladium, and platinum-group metals are oxidized into solution and separated from gold. The process is continuous and automated, requiring fewer personnel, reducing labor intensity, and lowering energy consumption. This significantly improves oxidation efficiency, further promoting the oxidative removal of copper and nickel, and leaching metals such as silver, ruthenium, rhodium, and iridium into solution.

[0016] Preferably, during the heating and pressurizing reaction, the concentration of sulfuric acid is 200-250 g / l.

[0017] Preferably, during the heating and pressurizing reaction process, gas is continuously removed to maintain a set oxygen partial pressure and prevent accumulation of inert gas.

[0018] In some embodiments, the first filtrate enters a copper electrolysis system for electrolysis.

[0019] In some embodiments, the silver precipitation is carried out by precipitating silver with sodium chloride.

[0020] In some embodiments, in the process of precipitating selenium and tellurium, copper powder is used for reduction to obtain selenium and tellurium precipitate.

[0021] In some embodiments, the mass ratio of the second filter cake, calcium oxide, and coke powder is 1:0.3-0.4:0.03-0.05. Calcium oxide serves as a slag-forming agent, and coke powder serves as a reducing agent.

[0022] Preferably, during the smelting process, slag is discharged at 1550°C and alloy is discharged at 1450°C.

[0023] The beneficial effects achieved by one or more embodiments of the present invention are as follows:

[0024] The present invention performs atmospheric pressure acid leaching and continuous high-pressure oxidation (multiple (4 to 6) oxidations are performed in one autoclave, mainly for strong oxidation to remove copper and nickel, and to leach metals such as silver, ruthenium, rhodium, and iridium into a solution) on the copper anode mud, so that silver and platinum group metals (ruthenium, rhodium, and iridium) in the copper anode mud are leached into a solution and recovered separately, thereby separating the platinum group metals (ruthenium, rhodium, and iridium) from gold, platinum, and palladium in advance, facilitating the subsequent pyrometallurgical treatment of the copper-removed anode mud, improving the recovery rate of the platinum group metals (ruthenium, rhodium, and iridium), and facilitating the subsequent recovery of gold, platinum, and palladium. BRIEF DESCRIPTION OF THE DRAWINGS

[0025] 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.

[0026] Figure 1 It is a process flow chart of an embodiment of the present invention. DETAILED DESCRIPTION

[0027] 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.

[0028] The present invention will be further described below with reference to the embodiments.

[0029] Example 1

[0030] like Figure 1 As shown, a method for treating high-platinum group metal copper anode mud comprises the following steps:

[0031] 1) Atmospheric pressure leaching

[0032] In the prepreg tank, add 5t of copper anode mud to 20m 3In a dilute sulfuric acid solution (concentration 146g / L), under mechanical stirring, steam heating, and air bubbling, the reaction temperature is 80°C and the reaction time is 4 hours, causing partial oxidation and dissolution of copper and nickel. After the reaction, the filtrate is filtered and returned to the copper electrolysis system. The filter cake enters the first slurry tank, where water and sulfuric acid are added to a concentration of 250g / L. It is then pumped into an autoclave, where the exhaust gas from the autoclave is used to heat the first slurry tank.

[0033] 2) High pressure leaching

[0034] The anode slurry was pressure-leached at a temperature of 200°C. The oxidant used was process oxygen (95% O2 content) at an oxygen pressure of 0.4 MPa. High-pressure leaching was carried out in a six-stage autoclave (effective volume 20 m3) equipped with a stirrer. 3 The reaction was carried out continuously in a continuous horizontal reactor, which was evenly divided into six reaction chambers by overflow partitions. The reaction system overflowed from one side and reacted sequentially until it flowed out from the outlet. Each reaction chamber was equipped with an agitator and oxygen was introduced. The continuous reaction time was 6 hours. The autoclave was equipped with a jacket for heating the slurry.

[0035] Due to the large amount of heat released during the pressure leaching process (head), the autoclave is equipped with two heat exchangers, into which cooling water is added to maintain the set temperature state.

[0036] In order to maintain the set oxygen partial pressure in the autoclave and prevent the accumulation of inert gas, gas is continuously removed during the leaching process and discharged into the first slurry tank.

[0037] The leached slurry enters the flash tank, where the slurry is rapidly cooled to below the boiling point and then flows to the transfer tank.

[0038] 3) Palladium precipitation and filtration

[0039] 196 kg of ferrous sulfate is added to the slurry in a transfer tank, which precipitates palladium in metallic form. The slurry undergoes two stages of filtration through a membrane filter press. After washing and air-blowing, the filter cake is sent to the pyrometallurgical area in step 13 to produce precious metal concentrate. The filtrate is stored in a silver precipitation tank equipped with a heat exchanger.

[0040] 4) Silver precipitation and filtration

[0041] In the silver precipitation tank, 87kg of sodium chloride is added to precipitate silver, and then filtered. The filter cake silver chloride is dried and refined, and the filtrate is stored in the selenium tellurium precipitation tank.

[0042] 5) Selenium and tellurium precipitation

[0043] In a selenium-tellurium precipitation tank, copper powder is used for reduction precipitation to obtain a selenium-tellurium precipitate. The selenium-tellurium filter cake is obtained by pressure filtration and the selenium-tellurium is recovered. The filtrate is used for step 6) of platinum group (rhodium, ruthenium, iridium) concentrate precipitation.

[0044] 6) Platinum group (rhodium, ruthenium, iridium) concentrate precipitation

[0045] 20kg of thiourea is added to the platinum group (rhodium, ruthenium, iridium) concentrate precipitation tank and then pumped into a tubular autoclave. Under high temperature and pressure (222°C, 1.8 MPa), the thiourea reacts with the platinum group (rhodium, ruthenium, iridium) elements to form a platinum group (rhodium, ruthenium, iridium) thiourea complex precipitate, along with copper. The slurry is then filtered, and the filtrate is returned to copper electrolysis. The filter cake, which is then used to remove copper from the platinum group (rhodium, ruthenium, iridium) concentrate, is then removed.

[0046] 7) Decoppering of platinum group (rhodium, ruthenium, iridium) concentrates

[0047] In the copper removal tank, water and sulfuric acid are added to the platinum group (rhodium, ruthenium, and iridium) concentrate. The diluted sulfuric acid concentration is 89 g / L, and air is introduced for oxidative copper removal. After copper removal, the concentrate is filtered, resulting in a filter cake, the platinum group (rhodium, ruthenium, and iridium) concentrate. This is then dried and refined. The filtrate is returned to the selenium and tellurium precipitation tank in step 5, and the filter cake proceeds to step 8).

[0048] 8) Selenium and tellurium filter cake treatment

[0049] The selenium tellurium filter cake is added to the alkaline leaching tank, and sodium hydroxide solution is added for alkaline leaching. The concentration of the sodium hydroxide solution is 110 g / L. After the reaction is completed, the filter cake is filtered and returned to the high-pressure leaching in step 2), and the filtrate enters the arsenic removal tank.

[0050] 9) Arsenic removal

[0051] Calcium oxide is added to the arsenic removal tank, and after the reaction, the filter cake is the arsenic filter cake to be processed. The filtrate is sent to the hydrolysis tank.

[0052] 10) Tellurium hydrolysis

[0053] Sulfuric acid is added to the hydrolysis tank to adjust the pH to 4. Sodium tellurate is hydrolyzed to form a crude tellurium dioxide precipitate. Slurry filtration produces a filter cake of tellurium dioxide. The filtrate is used to recover selenium.

[0054] 11) Secondary leaching of crude tellurium dioxide

[0055] The crude tellurium dioxide filter cake is slurried and alkaline leached with sodium hydroxide at a concentration of 110 g / L. After leaching, sulfuric acid is added to adjust the pH to 4, hydrolyzing the tellurium dioxide. The filter cake is then filtered and dried to produce refined tellurium dioxide. The filtrate is then sent to a selenium recovery tank.

[0056] 12) Selenium recovery

[0057] After the two hydrolysis steps, 1.35t of sodium sulfite is added to the selenium recovery tank. After the reaction is complete, the crude selenium slurry is filtered, and the filter cake is crude selenium. The filtrate is returned to wastewater treatment.

[0058] 13) High-pressure leaching filter cake pyrolysis

[0059] The filter cake obtained in step 3) is dried in a rotary drum dryer, and then calcium oxide and coke powder are added and pressed into blocks using a double-roll high-pressure mineral powder briquetting machine. The mass ratio of filter cake, calcium oxide and coke powder is 1:0.4:0.04. The exhaust gas generated by drying is discharged after passing through a bag filter. The agglomerated material is smelted in a DC arc furnace, and the smelting temperature is controlled at 1450℃-1550℃. Slag is discharged at 1550℃, and alloy is discharged at 1450℃. The alloy enters the atomizer for atomization, and spherical particles with a particle size of less than 100μm are collected in a silo to obtain gold, platinum and palladium alloy powder for processing. The slag is accumulated in one furnace and then sent to the electric arc furnace for smelting, and the secondary slag is returned to the copper system.

[0060] The recoveries of each element were: Pd 98.64%, Pt 98.78%, Rh 98.31%, Ru 98.22%, Ir 98.13%, Au 99.54%, and Ag 99.23%.

[0061] Example 2

[0062] like Figure 1 As shown, a method for treating high-platinum group metal copper anode mud comprises the following steps:

[0063] 1) Atmospheric pressure leaching

[0064] In the prepreg tank, 4.8 tons of copper anode mud is added to 19.5 m³ of dilute sulfuric acid solution (concentration: 135 g / L). Under mechanical stirring, steam heating, and air bubbling, the reaction temperature is 70-80°C and the reaction time is 4 hours to partially oxidize and dissolve the copper and nickel. After the reaction, the filtrate is filtered and returned to the copper electrolysis system. The filter cake enters the first slurry tank, where water and sulfuric acid are added to a concentration of 250 g / L. It is then pumped into an autoclave, where the exhaust gas from the autoclave is used to heat the first slurry tank.

[0065] 2) High pressure leaching

[0066] Anode slurry was pressure-leached at 190°C. The oxidant used was process oxygen (95% O₂ content) at a pressure of 0.5 MPa. High-pressure leaching was carried out continuously in a six-stage autoclave (same as in Example 1) equipped with a stirrer for 7 hours. The autoclave was equipped with a jacket for heating the slurry.

[0067] Due to the large amount of heat released during the pressure leaching process (head), the autoclave is equipped with two heat exchangers, into which cooling water is added to maintain the set temperature state.

[0068] In order to maintain the set oxygen partial pressure in the autoclave and prevent the accumulation of inert gas, gas is continuously removed during the leaching process and discharged into the first slurry tank.

[0069] The leached slurry enters the flash tank, where the slurry is rapidly cooled to below the boiling point and then flows to the transfer tank.

[0070] 3) Palladium precipitation and filtration

[0071] 175 kg of ferrous sulfate is added to the slurry in a transfer tank, which precipitates palladium in metallic form. The slurry undergoes two stages of filtration through a membrane filter press. After washing and air-blowing, the filter cake is sent to the pyrometallurgical area in step 13 to produce precious metal concentrate. The filtrate is stored in a silver precipitation tank equipped with a heat exchanger.

[0072] 4) Silver precipitation and filtration

[0073] In the silver precipitation tank, 87kg of sodium chloride is added to precipitate silver, and then filtered. The filter cake silver chloride is dried and refined, and the filtrate is stored in the selenium tellurium precipitation tank.

[0074] 5) Selenium and tellurium precipitation

[0075] In a selenium-tellurium precipitation tank, copper powder is used for reduction precipitation to obtain a selenium-tellurium precipitate. The selenium-tellurium filter cake is obtained by pressure filtration and the selenium-tellurium is recovered. The filtrate is used for step 6) of platinum group (rhodium, ruthenium, iridium) concentrate precipitation.

[0076] 6) Platinum group (rhodium, ruthenium, iridium) concentrate precipitation

[0077] 18kg of thiourea is added to the platinum group (rhodium, ruthenium, iridium) concentrate precipitation tank and then pumped into a tubular autoclave. Under high temperature and pressure (210°C, 1.6 MPa), the thiourea reacts with the platinum group (rhodium, ruthenium, iridium) elements to form a platinum group (rhodium, ruthenium, iridium) thiourea complex precipitate, along with copper. The slurry is then filtered, and the filtrate is returned to copper electrolysis. The filter cake, which is then used to remove copper from the platinum group (rhodium, ruthenium, iridium) concentrate, is then removed.

[0078] 7) Decoppering of platinum group (rhodium, ruthenium, iridium) concentrates

[0079] In the copper removal tank, water and sulfuric acid are added to the platinum group (rhodium, ruthenium, and iridium) concentrate. The diluted sulfuric acid concentration is 100 g / L, and air is introduced for oxidative copper removal. After copper removal, the concentrate is filtered, and the filter cake, representing the platinum group (rhodium, ruthenium, and iridium) concentrate, is dried and refined. The filtrate is returned to the selenium and tellurium precipitation tank in step 5), and the filter cake is sent to step 8).

[0080] 8) Selenium and tellurium filter cake treatment

[0081] The selenium tellurium filter cake is added to an alkaline leaching tank and sodium hydroxide solution is added for alkaline leaching. The concentration of the sodium hydroxide solution is 100 g / L. After the reaction is completed, the filter cake is filtered and returned to the high-pressure leaching in step 2). The filtrate enters the arsenic removal tank.

[0082] 9) Arsenic removal

[0083] Calcium oxide is added to the arsenic removal tank, and after the reaction, the filter cake is the arsenic filter cake to be processed. The filtrate is sent to the hydrolysis tank.

[0084] 10) Tellurium hydrolysis

[0085] Sulfuric acid is added to the hydrolysis tank to adjust the pH to 4.2. Sodium tellurate is hydrolyzed to form a crude tellurium dioxide precipitate. Slurry filtration produces a filter cake of tellurium dioxide. The filtrate is used to recover selenium.

[0086] 11) Secondary leaching of crude tellurium dioxide

[0087] The crude tellurium dioxide filter cake is slurried and alkaline leached with sodium hydroxide at a concentration of 100g / L. After leaching, sulfuric acid is added to adjust the pH to 4.1, hydrolyzing the tellurium dioxide. The filter cake is then filtered and dried to produce refined tellurium dioxide. The filtrate is then sent to a selenium recovery tank.

[0088] 12) Selenium recovery

[0089] After the two hydrolysis steps, 1.12 t of sodium sulfite is added to the selenium recovery tank. After the reaction is complete, the crude selenium slurry is filtered, and the filter cake is crude selenium. The filtrate is returned to wastewater treatment.

[0090] 13) High-pressure leaching filter cake pyrolysis

[0091] The filter cake obtained in step 3) is dried in a rotary drum dryer, and then calcium oxide and coke powder are added and pressed into blocks using a double-roll high-pressure mineral powder briquetting machine. The mass ratio of filter cake, calcium oxide and coke powder is 1:0.3:0.035. The exhaust gas generated by drying is discharged after passing through a bag filter. The agglomerated material is smelted in a DC arc furnace, and the smelting temperature is controlled at 1450℃-1550℃. Slag is discharged at 1550℃, and alloy is discharged at 1450℃. The alloy enters the atomizer for atomization to obtain spherical particles with a particle size of less than 100μm, which are collected in a silo to obtain gold-platinum-palladium alloy powder for processing. The slag is accumulated in one furnace and then sent to the electric arc furnace for smelting, and the secondary slag is returned to the copper system.

[0092] The recoveries of each element are: Pd 98.73%, Pt 98.69%, Rh 98.12%, Ru 98.20%, Ir 98.19%, Au 99.39%, and Ag 99.19%.

[0093] Comparative Example 1

[0094] The difference from Example 1 is that the six-stage autoclave is replaced by a common single-chamber autoclave, and the rest is the same as Example 1.

[0095] The recoveries of each element are: Pd 95.5%, Pt 95.4%, Rh 95.3%, Ru 70%, Ir 95.67%, Au 98.39%, and Ag 98.25%.

[0096] 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 shall be included within the scope of protection of the present invention.

Claims

1. A method for treating high-platinum group metal copper anode mud, characterized by: The steps include: Placing high-platinum group metal copper anode mud in dilute sulfuric acid, blowing air into it, and heating it to oxidize and leach the copper and nickel therein, and separating the solid and liquid to obtain a first filtrate and a first filter cake; The first filter cake is dispersed in dilute sulfuric acid, heated and pressurized in an oxygen atmosphere to react, dissolving part of the metal therein, and performing solid-liquid separation to obtain a second filtrate and a second filter cake. The second filter cake, calcium oxide, and coke powder are mixed, pressed into blocks, and smelted at 1450-1550°C for a set time. The smelted alloy is atomized to obtain fine particles for recovery, thereby obtaining gold-platinum-palladium alloy powder. After the second filtrate is flash evaporated and rapidly cooled, ferrous sulfate is added thereto to precipitate palladium. The filtered filtrate is sequentially subjected to the steps of silver precipitation and selenium and tellurium precipitation. Thiourea is added to the remaining filtrate and reacted under high temperature and high pressure to generate a thiourea complex precipitate. The solid and liquid are separated to obtain a third filtrate and a third filter cake, thereby obtaining a platinum group concentrate.

2. The method for treating high-platinum group metal copper anode slime according to claim 1, characterized in that: The temperature of heating high platinum group metal copper anode mud in dilute sulfuric acid is 70-80°C, and the reaction time is 3-5 hours.

3. The method for treating high-platinum group metal copper anode slime according to claim 1, characterized in that: The oxygen pressure of the heating and pressurizing reaction is 0.4-0.6 MPa, the temperature is 180-200° C., and the reaction time is 5-7 hours.

4. The method for treating high-platinum group metal copper anode slime according to claim 3, characterized in that: The heating and pressurizing reaction is carried out in a continuous horizontal reactor. The continuous horizontal reactor has a horizontal structure and is divided into 5-7 reaction chambers by overflow partitions. Each reaction chamber is equipped with an agitator and an oxygen inlet distributor. The reaction system overflows from one side of the horizontal continuous autoclave in sequence until it reaches the outlet.

5. The method for treating high-platinum group metal copper anode slime according to claim 3, characterized in that: During the heating and pressurizing reaction, the concentration of sulfuric acid is 200-250 g / l.

6. The method for treating high-platinum group metal copper anode slime according to claim 5, characterized in that: During the heating and pressurizing reaction process, the gas is continuously removed.

7. The method for treating high-platinum group metal copper anode slime according to claim 1, characterized in that: The first filtrate enters the copper electrolysis system for electrolysis.

8. The method for treating high-platinum group metal copper anode slime according to claim 1, characterized in that: The silver precipitation is carried out by using sodium chloride to precipitate silver.

9. The method for treating high-platinum group metal copper anode slime according to claim 1, characterized in that: In the treatment of precipitated selenium and tellurium, copper powder is used for reduction to obtain selenium and tellurium precipitate.

10. The method for treating high-platinum group metal copper anode slime according to claim 1, characterized in that: The mass ratio of the second filter cake, calcium oxide and coke powder is 1:0.3-0.4:0.03-0.05.

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