Method for recovering valuable metals from copper anode slime step by step

Abstract

The invention discloses a method for recovering valuable metals from copper anode slime step by step, and belongs to the technical field of metallurgy and chemical industry. According to the method, through control over the oxidation-reduction potential, the oxidation-reduction potential is high in the leaching stage, and it is guaranteed that gold, selenium, tellurium and the like can be oxidized to enter a leaching solution; in the reduction stage, gold, selenium and tellurium products are reduced step by step according to the difference of gold, selenium and tellurium in reduction potentials; in the leaching stage, soluble chlorine salt is added while enough oxidation potential is guaranteed, enough chloride ions are provided for the solution, and therefore gold can enter the solution in the form of AuCl4-(tetrachloroauric acid); meanwhile, in the oxidation leaching stage, a small amount of silver enters the solution in the form of silver sulfate or silver complex, and under the condition that enough chloride ions are provided, silver is precipitated and retained in residues in the form of silver chloride, so that separation of gold and silver is achieved, and meanwhile dispersion of gold and silver in the leaching process is reduced. According to the method, valuable metal is efficiently separated and recycled.

Description

Technical Field

[0001] This invention relates to a method for the stepwise recovery of valuable metals from copper anode mud, belonging to the field of metallurgical and chemical technology. Background Technology

[0002] Copper anode slime is a major byproduct of copper electrolytic refining, primarily originating from inert metallic impurities in the copper anode plate that deposit in solid form at the bottom of the electrolytic cell during electrolysis. Copper anode slime has extremely high recycling value, being rich in various high-value metals such as gold, silver, selenium, and tellurium. Extraction from copper anode slime is one of the world's most important sources of selenium, tellurium, gold, and silver; over 80% of selenium and tellurium and 40% of gold and silver originate from copper anode slime recycling processes.

[0003] Currently, the mainstream copper anode slime recovery processes mainly include pyrometallurgical processes, hydrometallurgical processes, and combined beneficiation and smelting processes. In pyrometallurgical processing, valuable metals are often dispersed in various intermediate products, making the recovery process complex and inefficient. Furthermore, pyrometallurgical processes suffer from severe environmental pollution, high energy consumption, and long production cycles. In comparison, hydrometallurgical processes offer higher recovery efficiency, but they have stricter requirements regarding the composition and phase stability of the copper anode slime. Combined beneficiation and smelting processes can effectively enrich precious metals such as gold and silver through mineral processing, facilitating recovery; however, this process also suffers from high gold and silver content in the beneficiation tailings and low tellurium direct recovery rates.

[0004] Achieving efficient and selective recovery of selenium, tellurium, gold, and silver has been a key focus for smelting companies. Summary of the Invention

[0005] To address the shortcomings of the prior art, this invention provides a method for the stepwise recovery of valuable metals from copper anode mud.

[0006] To achieve the above objectives, the technical solution adopted by the present invention is as follows: A method for stepwise recovery of valuable metals from copper anode slime includes the following steps: (1) Add sulfuric acid, oxidant and soluble chloride salt to copper anode mud, adjust pH and control redox potential to 1050~1300mV. After the reaction is completed, separate solid and liquid to obtain acid leaching filtrate and acid leaching residue.

[0007] (2) Add reducing agent I to the acid leaching filtrate obtained in step (1), adjust the pH, control the oxidation-reduction potential to 730~950mV, and after the reaction is completed, separate the solid and liquid to obtain the reduced liquid and crude gold powder.

[0008] (3) Add reducing agent II to the first reduction solution obtained in step (2), control the pH, and control the redox potential between 550 and 720 mV. After the reaction is completed, separate the solid and liquid to obtain the second reduction solution and crude selenium.

[0009] (4) Add copper powder to the secondary reduction solution obtained in step (3), and after the reaction is completed, separate the solid and liquid to obtain copper sulfate solution and copper telluride.

[0010] (5) Add thiosulfate solution to the acid leaching residue obtained in step (1) to adjust the pH. After the reaction, silver-separated solution and lead-rich slag are obtained.

[0011] (6) Add reducing agent III to the silver separation solution obtained in step (5), and crude silver powder is obtained after the reaction.

[0012] Preferably, in step (1), the liquid-to-solid ratio of sulfuric acid solution to copper anode mud is (3~7):1, with units of mL:g, the mass concentration of sulfuric acid in the sulfuric acid solution is 80~200g / L, the pH is adjusted to below 0.5, the reaction temperature is 50~90℃, and the reaction time is 2~6h.

[0013] Preferably, in step (1), the mass ratio of the amount of oxidant added to the theoretically required oxidant for the oxidation of selenium and tellurium in the copper anode mud is (1.2~2.5):1; the oxidation-reduction potential is controlled by controlling the rate of oxidant addition; wherein the oxidant is one of hydrogen peroxide, ozone, manganese dioxide or nitric acid.

[0014] Preferably, the molar ratio of the amount of soluble chloride added in step (1) to the total amount of gold and silver in the copper anode mud is (8~12):1, wherein the soluble chloride is one of sodium chloride, potassium chloride or ammonium chloride.

[0015] Preferably, in step (2), the mass ratio of the amount of reducing agent I added to the theoretical amount of reducing agent required for gold reduction in the acid leaching filtrate is (1.1~2.0):1. The reducing agent I is one or more of oxalic acid, ascorbic acid, and hydrazine hydrate mixed in any proportion. The pH value is adjusted to 1~2, the reaction temperature is 40~70℃, and the reaction time is 1~3h.

[0016] Preferably, in step (3), the mass ratio of the amount of reducing agent II added to the liquid after the first reduction to the theoretical amount of reducing agent required for the reduction of selenium in the liquid after the first reduction is (1.2~1.8):1. The reducing agent II is one of sulfur dioxide, sodium sulfite or iron powder. The pH is controlled at 1~3, the reaction temperature is 60~80℃, and the reaction time is 0.5~4h.

[0017] Preferably, in step (4), the mass ratio of the amount of copper powder added to the theoretical amount of tellurium replacement required in the solution after secondary reduction is (1.5~4.5):1, the reaction temperature is 60~90℃, and the reaction time is 2~6h.

[0018] Preferably, in step (5), the molar ratio of the amount of thiosulfate added to the silver content in the acid leaching residue is (2.4~4):1, and the liquid-solid ratio of the extract to the acid leaching residue is (3~7):1, with units of mL:g; the thiosulfate is one or more of sodium thiosulfate, potassium thiosulfate, and ammonium thiosulfate mixed in any proportion; the pH is adjusted to 8~10, the reaction temperature is 25~50℃, and the reaction time is 2~5h.

[0019] Preferably, in step (6), the mass ratio of the amount of reducing agent III added to the theoretical amount of silver required for silver reduction in the silver separation solution is (1.2~1.5):1, and the reducing agent III is zinc powder, iron powder, or sodium sulfite; the reaction temperature is 50~80℃, and the reaction time is 1~4h.

[0020] Preferably, sulfuric acid or sodium hydroxide is used to adjust the pH in each step.

[0021] The beneficial effects of this invention are: (1) In order to solve the problem of low direct recovery rate of valuable metals caused by long recovery process and many metal dispersion points in copper anode mud, this invention achieves efficient recovery of valuable metals by controlling the oxidation-reduction potential at different leaching stages and combining the selection of oxidant, reducing agent and chloride salt.

[0022] (2) The process of this invention is short, simple to control, uses few reagents, and causes little pollution. It can selectively recover valuable metals from copper anode mud and has good prospects for industrialization. Attached Figure Description

[0023] Figure 1 This is a flowchart illustrating the method described in this invention. Detailed Implementation

[0024] The technical solutions provided by the present invention will be described in detail below with reference to the embodiments, but they should not be construed as limiting the scope of protection of the present invention.

[0025] Example 1 A method for stepwise recovery of valuable metals from copper anode slime, wherein the main components of the copper anode slime are: Au 0.95wt%, Ag 6.67wt%, Cu 23.01wt%, Te 3.52wt%, Se 6.14wt%, and Pb 6.01%.

[0026] Includes the following steps: (1) In 1L of sulfuric acid solution with a mass concentration of 160g / L, the liquid-to-solid ratio of sulfuric acid solution to copper anode mud is 4:1. 250g of copper anode mud is added. The total amount of gold and silver in the copper anode mud is 0.167mol. 2mol of sodium chloride is added according to the molar ratio of soluble chloride addition to the total gold and silver content in the copper anode mud of 12:1. The pH of the slurry is adjusted to 0.1. The oxidation-reduction potential is monitored by an online oxidation-reduction potential analyzer (ORP). The oxidation-reduction potential is controlled to 1200mV by controlling the rate of hydrogen peroxide addition. The total amount of hydrogen peroxide added is 1.5 times the amount of oxidant (119g) theoretically required for the oxidation of selenium and tellurium in the copper anode mud. Add 178.5g of 30% hydrogen peroxide; place the slurry in a reactor and carry out oxidative leaching at a reaction temperature of 80℃ for 2 hours; after leaching, perform solid-liquid separation to obtain acid leaching residue and acid leaching filtrate, thus achieving efficient leaching of Cu, Au, Se, and Te; send the acid leaching residue and acid leaching filtrate for analysis, and calculate the leaching rate of Cu as 95.47%, Au as 87.03%, Se as 90.99%, Te as 94.31% based on the element content in the acid leaching residue, Ag remains in the residue in the form of silver chloride, with a sedimentation rate of 91.03%, and lead remains in the residue in the form of lead sulfate, with almost no leaching.

[0027] (2) Take 500 ml of the acid leaching filtrate obtained in step (1), adjust the pH of the solution to 1, put the solution into the reaction vessel, and reduce and recover gold by controlling the addition rate of reducing agent at 60℃ and 2h for 2h. Under the condition of controlling the redox potential to 740mV, 1.4g of reducing agent I (theoretical amount of reducing agent is 0.7g) is added. After the reaction, the solid and liquid are separated to obtain the liquid after one reduction and crude gold powder. The liquid after one reduction and crude gold powder are analyzed and tested. The gold reduction recovery rate is calculated to be 97.45%. Only 1.02% of selenium enters the crude gold powder during the reduction process. Tellurium and copper are almost not lost and remain in the liquid. The reducing agent I is oxalic acid (AR).

[0028] (3) Take 450 ml of the first reduction solution obtained in step (2), control the pH of the solution to 2, put the solution into the reaction vessel, and under the conditions of 60℃ reaction temperature and 2h reaction time, control the redox potential to 550mV by controlling the flow rate of reducing agent II to reduce and recover selenium; after the reaction, the solid and liquid are separated to obtain the second reduction solution and crude selenium. The second reduction solution and crude selenium are analyzed and detected, and the selenium reduction recovery rate is calculated to be 97.45%. Only 2.01% of tellurium enters the crude selenium during the reduction process, and copper is almost not lost and remains in the liquid; the reducing agent II is sulfur dioxide.

[0029] (4) The tellurium concentration in the solution after the second reduction in step (3) is 8.13 g / L. Take 400 ml of the solution after the second reduction obtained in step (3), the tellurium content is 3.25 g, and the theoretical copper powder required to replace tellurium with copper telluride is 9.3 g. According to the ratio of actual addition to theoretical value of 1.5:1, add 13.95 g of copper powder and reduce and recover tellurium under the conditions of reaction temperature of 80℃ and reaction time of 3 h. After the reaction, the solid and liquid are separated to obtain copper sulfate solution and copper telluride. The copper sulfate solution and copper telluride are analyzed and detected. The tellurium reduction recovery rate is calculated to be 96.47%, and only 1.36% of selenium enters the copper telluride during the reduction process.

[0030] (5) In step (1), the silver content in the acid leaching residue accounts for 33.86% of the total mass of the acid leaching residue. Take 40g of acid leaching residue, the silver content is 0.126mol (about 13.59g). According to the molar ratio of thiosulfate to silver content of 4:1, 0.506mol of sodium thiosulfate needs to be added. Prepare 200mL of sodium thiosulfate into an extraction solution to leach the acid leaching residue. The leaching reaction temperature is 40℃, the reaction time is 3h, and the pH is 9. After the leaching is completed, solid-liquid separation is performed to obtain lead-rich slag and silver-separated solution. The lead-rich slag and silver-separated solution are analyzed. According to the silver content in the silver-separated solution and the acid leaching residue, the silver leaching rate is calculated to be 96.79%. The lead is almost not leached and is retained in the lead-rich slag.

[0031] (6) Take 100 mL of the silver-containing solution obtained in step (5), in which the silver content is 0.061 mol. Theoretically, 0.0305 mol of reducing agent III is required. According to the ratio of actual addition to theoretical addition of 1.2:1, add 0.0366 mol of reducing agent III and carry out reduction and silver extraction at a reaction temperature of 60℃ and a reaction time of 2h. After the reaction, crude silver powder and silver-extracted liquid are obtained. The crude silver powder and silver-extracted liquid are analyzed and the silver reduction recovery rate is calculated to be 98.03%. The reducing agent III is sodium sulfite.

[0032] Example 2 A method for stepwise recovery of valuable metals from copper anode mud, wherein the main component content of the copper anode mud is consistent with that in Example 1.

[0033] Includes the following steps: (1) In 900 mL of sulfuric acid solution with a mass concentration of 80 g / L, add 300 g of copper anode mud at a liquid-to-solid ratio of 3:1. The total amount of gold and silver in the copper anode mud is 0.2 mol. Add 2 mol of potassium chloride at a molar ratio of soluble chloride to gold and silver content of 10:1. Adjust the pH of the slurry to 0.1. Monitor the oxidation-reduction potential using an online oxidation-reduction potential analyzer (ORP). Control the oxidation-reduction potential to 1050 mV by controlling the rate of manganese dioxide addition. The total amount of manganese dioxide added is 1.2 times the theoretically required oxidant (109.62 g), and 131 g of manganese dioxide needs to be added. 54g; The slurry was placed in a reactor and oxidized and leached at a reaction temperature of 50℃ for 6 hours. After leaching, solid-liquid separation was performed to obtain acid leaching residue and acid leaching filtrate, thus achieving efficient leaching of Cu, Au, Se, and Te. The acid leaching residue and acid leaching filtrate were sent for analysis. Based on the element content in the acid leaching residue, the leaching rate of Cu was calculated to be 96.32%, the leaching rate of Au was 86.33%, the leaching rate of Se was 92.24%, the leaching rate of Te was 95.03%, Ag remained in the residue in the form of silver chloride, the sedimentation rate was 90.12%, and lead remained in the residue in the form of lead sulfate, with almost no leaching.

[0034] (2) Take 500ml of the acid leaching filtrate obtained in step (1), adjust the pH of the solution to 2, put the solution into the reaction vessel, add 1.507g of reducing agent I (theoretical reducing agent amount is 1.37g) according to the mass ratio of the amount of reducing agent I added to the theoretical amount of reducing agent required for reducing gold in the acid leaching filtrate of 1.1:1, and reduce and recover gold under the conditions of 40℃ reaction temperature, 3h reaction time, and controlled redox potential of 730mV by controlling the rate of addition of reducing agent; after the reaction, the solid and liquid are separated to obtain the liquid after one reduction and crude gold powder. The liquid after one reduction and crude gold powder are analyzed and detected. According to the gold content of the liquid after reduction, the gold reduction recovery rate is 96.79%. Only 0.89% of selenium enters the crude gold powder during the reduction process, and tellurium and copper are almost not lost and remain in the liquid; the reducing agent I is ascorbic acid (AR).

[0035] (3) Take 450 ml of the first reduction solution obtained in step (2), control the pH of the solution to 2, put the solution into the reaction vessel, and introduce 40.14 g of reducing agent II (theoretical reducing agent II is 33.45 g) according to the mass ratio of the amount of original agent II added to the theoretical amount of reducing agent required for the reduction of selenium in the first reduction solution of 1.2:1. The reaction temperature is 80℃, the reaction time is 1 h, and the redox potential is controlled by controlling the flow rate of reducing agent II and the redox potential is controlled to recover selenium. After the reaction, the solid and liquid are separated to obtain the second reduction solution and crude selenium. The second reduction solution and crude selenium are analyzed and detected. The selenium reduction recovery rate is calculated to be 98.03%. Only 2.86% of tellurium enters the crude selenium during the reduction process, and copper is almost not lost and remains in the liquid. The reducing agent II is sodium sulfite.

[0036] (4) The tellurium concentration in the solution after the second reduction in step (3) is 8.12 g / L. Take 400 ml of the solution after the second reduction obtained in step (3), the tellurium content is 3.248 g, and the theoretical copper powder required to replace tellurium with copper telluride is 9.3 g. According to the ratio of actual addition to theoretical value of 2:1, add 18.6 g of copper powder and reduce and recover tellurium under the conditions of reaction temperature of 90℃ and reaction time of 2 h. After the reaction, the solid and liquid are separated to obtain copper sulfate solution and copper telluride. The copper sulfate solution and copper telluride are analyzed and detected. The tellurium reduction recovery rate is calculated to be 97.69%, and only 2.01% of selenium enters the copper telluride during the reduction process.

[0037] (5) In step (1), the silver content in the acid leaching residue is 32.16%. Take 40g of acid leaching residue, the silver content is 0.12mol. According to the molar ratio of potassium thiosulfate to silver content of 3:1, 0.36mol of potassium thiosulfate needs to be added. Prepare 200mL of potassium thiosulfate leaching solution to leach the acid leaching residue. The leaching reaction temperature is 50℃, the reaction time is 2h, and the pH is 10. After the leaching is completed, solid and liquid separation is obtained to obtain lead-rich slag and silver-separated solution. The lead-rich slag and silver-separated solution are analyzed. The silver leaching rate is calculated to be 95.98% based on the silver content of the lead-rich slag. The lead is almost not leached and is all retained in the lead-rich slag.

[0038] (6) Take 100 mL of the silver-containing solution obtained in step (5), in which the silver content is 0.057 mol and the theoretical amount required is 0.0285 mol of reducing agent III. Add 0.0387 mol of reducing agent III according to the ratio of actual amount added to theoretical amount added of 1.5:1. Reduce and extract silver under the conditions of reaction temperature of 50℃ and reaction time of 3h. After the reaction, crude silver powder and silver-extracted liquid are obtained. The crude silver powder and silver-extracted liquid are analyzed and the silver reduction recovery rate is calculated to be 97.36%. The reducing agent III is zinc powder.

[0039] Example 3 A method for stepwise recovery of valuable metals from copper anode mud, wherein the main component content of the copper anode mud is consistent with that in Example 1.

[0040] Includes the following steps: (1) In 700 mL of sulfuric acid solution with a mass concentration of 200 g / L, 250 (100) g of copper anode mud was added according to the liquid-solid ratio of sulfuric acid solution to copper anode mud of 7:1. The total amount of gold and silver in the copper anode mud was 0.067 mol. According to the molar ratio of soluble chloride salt added to the total content of gold and silver of 8:1, 28.59 g (0.534 mol) of ammonium chloride was added to adjust the pH of the slurry to 0.1. The oxidation-reduction potential was monitored by an online oxidation-reduction potential analyzer (ORP). The oxidation-reduction potential was controlled to 1300 mV by controlling the ozone introduction rate. The total amount of ozone added was 2.5 times the theoretically required oxidant (6.75 g). Adding ozone requires 16.88g; the slurry is placed in a reactor and oxidative leaching is carried out at a reaction temperature of 90℃ for 3 hours; after leaching, solid-liquid separation is performed to obtain acid leaching residue and acid leaching filtrate, thus achieving efficient leaching of Cu, Au, Se, and Te; the acid leaching residue and acid leaching filtrate are sent for analysis, and the leaching rate of Cu is calculated to be 97.01%, the leaching rate of Au is 85.13%, the leaching rate of Se is 93.21%, the leaching rate of Te is 95.16%, Ag remains in the residue in the form of silver chloride, the sedimentation rate is 91.01%, and lead remains in the residue in the form of lead sulfate, with almost no leaching.

[0041] (2) Take 500 ml of the acid leaching filtrate obtained in step (1), adjust the pH of the solution to 2, put the solution into the reaction vessel, add 0.3 g of reducing agent I (theoretical reducing agent amount is 0.2 g) according to the mass ratio of the amount of reducing agent I added to the theoretical amount of reducing agent required for gold reduction in the acid leaching filtrate of 1.5:1, and reduce and recover gold under the conditions of 70℃ reaction temperature, 1h reaction time, and controlled redox potential of 950mV by controlling the rate of addition of reducing agent; after the reaction, the solid and liquid are separated to obtain the liquid after one reduction and crude gold powder. The liquid after one reduction and crude gold powder are analyzed and detected. According to the gold content of the liquid after reduction, the gold reduction recovery rate is 95.83%. Only 0.54% of selenium enters the crude gold powder during the reduction process, and tellurium and copper are almost not lost and remain in the liquid; the reducing agent I is hydrazine hydrate.

[0042] (3) Take 450 ml of the first reduction solution obtained in step (2), control the pH of the solution to 2, put the solution into the reaction vessel, the mass ratio of the amount of reducing agent II added in the first reduction solution to the theoretical amount of reducing agent required for the reduction of selenium in the first reduction solution is 1.8:1, and introduce 10.44 g of reducing agent II (the theoretical amount of reducing agent II is 5.8 g) to reduce and recover selenium; after the reaction is completed under the conditions of reaction temperature of 80℃, reaction time of 1 h, control of the flow rate of reducing agent II, and control of the redox potential of 600 mV, the solid and liquid are separated to obtain the second reduction solution and crude selenium. The first reduction solution and crude selenium are analyzed and detected, and the selenium reduction recovery rate is calculated to be 97.12%. Only 3.06% of tellurium enters the crude selenium during the reduction process, and copper is almost not lost and remains in the liquid; the reducing agent II is iron powder.

[0043] (4) The tellurium concentration in the solution after the second reduction in step (3) is 4.63 g / L. Take 400 ml of the solution after the second reduction obtained in step (3), the tellurium content is 1.85 g, and the theoretical copper powder required to replace tellurium with copper telluride is 5.29 g. According to the ratio of actual addition to theoretical value of 4.5:1, add 23.82 g of copper powder and reduce and recover tellurium under the conditions of reaction temperature of 60℃ and reaction time of 6 h. After the reaction, the solid and liquid are separated to obtain copper sulfate solution and copper telluride. The copper sulfate solution and copper telluride are analyzed and detected. The tellurium reduction recovery rate is calculated to be 95.83%, and only 1.76% of selenium enters the copper telluride during the reduction process.

[0044] (5) In step (1), the silver content in the acid leaching residue is 31.86%. Take 20g of acid leaching residue, the silver content is 0.06mol. Add 0.144mol of ammonium thiosulfate according to the molar ratio of ammonium thiosulfate to silver content of 2.4:1. Prepare 100mL of leaching solution with ammonium thiosulfate to leach the acid leaching residue. The leaching reaction temperature is 25℃, the reaction time is 5h, and the pH is 8. After leaching, solid and liquid separation is obtained to obtain lead-rich slag and silver-separated solution. Analyze the lead-rich slag and silver-separated solution. The silver leaching rate is calculated to be 96.48% based on the silver content of the lead-rich slag. The lead is almost not leached and is retained in the lead-rich slag.

[0045] (6) Take 100 mL of the silver-containing solution obtained in step (5), in which the silver content is 0.058 mol and the theoretical amount required is 0.029 mol of reducing agent III. Add 0.038 mol of reducing agent III according to the ratio of actual amount added to theoretical amount added of 1.3:1. Reduce and extract silver under the conditions of reaction temperature of 80℃ and reaction time of 1 h. After the reaction, crude silver powder and silver-extracted liquid are obtained. Analyze the crude silver powder and silver-extracted liquid and calculate that the silver reduction recovery rate is 96.12%. The reducing agent III is iron powder.

[0046] Comparative Example 1 A method for stepwise recovery of valuable metals from copper anode slime, wherein the main component content of the copper anode slime is consistent with that in Example 1.

[0047] Includes the following steps: (1) In 1L of sulfuric acid solution with a mass concentration of 160g / L, the liquid-solid ratio of sulfuric acid solution to copper anode mud is 4:1. 250g of copper anode mud is added. The total amount of gold and silver in the copper anode mud is 0.167mol. According to the molar ratio of chloride salt addition to gold and silver content of 12:1, 116.88g of sodium chloride is added. The total amount of hydrogen peroxide added is 1.5 times the theoretical required oxidant (119g). 178.5g of 30% hydrogen peroxide is added. The redox potential of the system is not controlled during the process. The slurry is put into the reactor and reacted at a temperature of 80℃. Oxidative leaching was carried out over a period of 2 hours. After leaching, solid-liquid separation was performed to obtain acid leaching residue and acid leaching filtrate, thus achieving efficient leaching of Cu, Au, Se, and Te. The acid leaching residue and acid leaching filtrate were sent for analysis. Based on the element content in the acid leaching residue, the leaching rates of Cu, Au, Se, and Te were calculated to be 95.23%, 76.33%, 91.32%, and 93.41%, respectively. Ag remained in the residue as silver chloride, with a sedimentation rate of 89.21%. Lead remained in the residue as lead sulfate, with virtually no leaching.

[0048] (2) Take 500 ml of the acid leaching filtrate obtained in step (1), adjust the pH of the solution to 1, put the solution into the reaction vessel, and add 1.4 g of reducing agent I (theoretical reducing agent dosage is 0.7 g) at a time at a reaction temperature of 60℃ and a reaction time of 2 h. The gold is recovered by reducing without controlling the redox potential. After the reaction, the solid and liquid are separated to obtain the liquid after the first reduction and the crude gold powder. The liquid after the first reduction and the crude gold powder are analyzed and tested. The gold reduction recovery rate is calculated to be 93.15%. 9.75% of the selenium enters the crude gold powder during the reduction process, 2.26% of the tellurium enters the crude gold powder, and the copper is almost not lost and remains in the liquid. The reducing agent I is oxalic acid (AR).

[0049] (3) Take 450 ml of the first reduction solution obtained in step (2), control the pH of the solution to 2, put the solution into the reaction vessel, and pass 9.15 L of reducing agent II (theoretical amount of reducing agent II is 6.1 L) under the condition of reaction temperature of 60℃, reaction time of 2 h, and no control of redox potential to reduce and recover selenium; after the reaction, the solid and liquid are separated to obtain the second reduction solution and crude selenium. The first reduction solution and crude selenium are analyzed and detected, and the reduction recovery rate of selenium is calculated to be 97.31%. More than 8.61% of tellurium enters the crude selenium during the reduction process, and copper is almost not lost and remains in the liquid; the reducing agent II is sulfur dioxide.

[0050] (4) The tellurium concentration in the solution after the second reduction in step (3) is 7.34 g / L. Take 400 ml of the solution after the second reduction obtained in step (3), the tellurium content is 2.94 g, and the theoretical copper powder required to replace tellurium with copper telluride is 8.4 g. According to the ratio of actual addition to theoretical value of 1.5:1, add 12.62 g of copper powder and reduce and recover tellurium under the conditions of reaction temperature of 80℃ and reaction time of 3 h. After the reaction, the solid and liquid are separated to obtain copper sulfate solution and copper telluride. The copper sulfate solution and copper telluride are analyzed and detected. The tellurium reduction recovery rate is calculated to be 94.53%, and 2.55% of selenium enters the copper telluride during the reduction process.

[0051] (5) In step (1), the silver content in the acid leaching residue is 33.11%. Take 40g of acid leaching residue, the silver content is 0.125mol. Add 0.5mol of sodium thiosulfate according to the molar ratio of sodium thiosulfate to silver content of 4:1. Prepare 200mL of sodium thiosulfate into an extraction solution to leach the acid leaching residue. The leaching is carried out at a reaction temperature of 40℃ and a reaction time of 3h. After the leaching is completed, solid-liquid separation is performed to obtain lead-rich slag and silver-separated solution. The lead-rich slag and silver-separated solution are analyzed. The silver leaching rate is calculated to be 95.81% based on the silver content of the lead-rich slag. The lead is almost not leached and is retained in the lead-rich slag.

[0052] (6) Take 100 mL of the silver-containing solution obtained in step (5), in which the silver content is 0.06 mol and theoretically 0.03 mol of reducing agent III is required. Add 0.036 mol of reducing agent III according to the ratio of actual addition to theoretical addition of 1.2:1. Reduce and extract silver under the conditions of reaction temperature of 60℃ and reaction time of 2h. After the reaction, crude silver powder and silver-extracted liquid are obtained. Analyze the crude silver powder and silver-extracted liquid and calculate that the silver reduction recovery rate is 96.78%. The reducing agent III is sodium sulfite.

[0053] Compared with Example 1, it was found that the redox potential of the control system has a significant impact on metal recovery. In the stepwise recovery of gold, selenium and tellurium, the reducing agent has a reduction effect on all three metals. Under the condition of not controlling the redox potential, the separation effect of gold, selenium and tellurium is weakened, and the content of other elements doped in the reduced product will be significantly increased, resulting in a decrease in product grade and a decrease in metal recovery rate.

[0054] Finally, it should be noted that the above embodiments are only used to illustrate the technical solutions of the present invention and are not intended to limit the scope of protection of the present invention. Although the present invention has been described in detail with reference to preferred embodiments, those skilled in the art should understand that modifications or equivalent substitutions can be made to the technical solutions of the present invention without departing from the essence and scope of the technical solutions of the present invention.

Claims

1. A method for stepwise recovery of valuable metals from copper anode mud, characterized in that: Includes the following steps: (1) Add sulfuric acid, oxidant and soluble chloride salt to copper anode mud, adjust pH and control redox potential to 1050~1300mV. After the reaction is completed, separate solid and liquid to obtain acid leaching filtrate and acid leaching residue. (2) Add reducing agent I to the acid leaching filtrate obtained in step (1), adjust the pH, control the redox potential to 730~950mV, and after the reaction is completed, separate the solid and liquid to obtain the reduced liquid and crude gold powder. (3) Add reducing agent II to the first reduction solution obtained in step (2), control the pH, control the redox potential to be between 550 and 720 mV, and after the reaction is completed, separate the solid and liquid to obtain the second reduction solution and crude selenium. (4) Add copper powder to the secondary reduction solution obtained in step (3), and after the reaction is completed, separate the solid and liquid to obtain copper sulfate solution and copper telluride; (5) Add the leaching solution prepared with thiosulfate to the acid leaching residue obtained in step (1), adjust the pH, and after the reaction, obtain silver-separated solution and lead-rich slag. (6) Add reducing agent III to the silver separation solution obtained in step (5), and crude silver powder is obtained after the reaction.

2. The method for stepwise recovery of valuable metals from copper anode mud according to claim 1, characterized in that: In step (1), the liquid-solid ratio of sulfuric acid solution to copper anode mud is (3~7):1, with units of mL:g. The mass concentration of sulfuric acid in the sulfuric acid solution is 80~200g / L. The pH is adjusted to below 0.5, the reaction temperature is 50~90℃, and the reaction time is 2~6h.

3. The method for stepwise recovery of valuable metals from copper anode mud according to claim 1, characterized in that: In step (1), the mass ratio of the amount of oxidant added to the theoretically required oxidant for the oxidation of selenium and tellurium in the copper anode mud is (1.2~2.5):1; the redox potential is controlled by controlling the rate of oxidant addition. The oxidant is one of hydrogen peroxide, ozone, manganese dioxide, or nitric acid.

4. The method for stepwise recovery of valuable metals from copper anode mud according to claim 1, characterized in that: In step (1), the molar ratio of the amount of soluble chloride added to the total amount of gold and silver in the copper anode mud is (8~12):1, wherein the soluble chloride is one of sodium chloride, potassium chloride or ammonium chloride.

5. The method for stepwise recovery of valuable metals from copper anode slime according to claim 1, characterized in that: In step (2), the mass ratio of the amount of reducing agent I added to the theoretical amount of reducing agent required for gold reduction in the acid leaching filtrate is (1.1~2.0):

1. The reducing agent I is one or more of oxalic acid, ascorbic acid, and hydrazine hydrate mixed in any proportion. The pH value is adjusted to 1~2, the reaction temperature is 40~70℃, and the reaction time is 1~3h.

6. The method for stepwise recovery of valuable metals from copper anode slime according to claim 1, characterized in that: In step (3), the mass ratio of the amount of reducing agent II added to the liquid after the first reduction to the theoretical amount of reducing agent required for selenium reduction in the liquid after the first reduction is (1.2~1.8):

1. The reducing agent II is one of sulfur dioxide, sodium sulfite or iron powder. The pH is controlled at 1~3, the reaction temperature is 60~80℃, and the reaction time is 0.5~4h.

7. The method for stepwise recovery of valuable metals from copper anode slime according to claim 1, characterized in that: In step (4), the mass ratio of the amount of copper powder added to the theoretical amount of tellurium replacement required in the solution after secondary reduction is (1.5~4.5):1, the reaction temperature is 60~90℃, and the reaction time is 2~6h.

8. The method for stepwise recovery of valuable metals from copper anode slime according to claim 1, characterized in that: In step (5), the molar ratio of the amount of thiosulfate added to the silver content in the acid leaching residue is (2.4~4):1, and the liquid-solid ratio of the extract to the acid leaching residue is (3~7):1, with units of mL:g; the thiosulfate is one or more of sodium thiosulfate, potassium thiosulfate, and ammonium thiosulfate mixed in any proportion; the pH is adjusted to 8~10, the reaction temperature is 25~50℃, and the reaction time is 2~5h.

9. The method for stepwise recovery of valuable metals from copper anode slime according to claim 1, characterized in that: In step (6), the mass ratio of the amount of reducing agent III added to the theoretical amount required for silver reduction in the silver separation solution is (1.2~1.5):

1. The reducing agent III is zinc powder, iron powder, or sodium sulfite. The reaction temperature is 50~80℃ and the reaction time is 1~4h.

10. The method according to claim 1, characterized in that: Sulfuric acid or sodium hydroxide is used to adjust the pH in each step.

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