Method for reducing copper content in copper electrolysis anode slime and retaining rare and precious elements in copper electrolysis anode slime

Abstract

The invention discloses a method for reducing the content of copper in copper electrolysis anode slime and retaining rare and precious elements in the copper electrolysis anode slime, which comprises the following steps: (1) putting the copper electrolysis anode slime and electrolyte into a thickener for concentration, and pumping out supernate for later use; (2) pumping anode mud at the bottom of a thickener into a reaction kettle, adding a waste acid solution, carrying out stirring reaction for 3-5 hours, then carrying out primary solid-liquid separation, combining primary filtrate with supernatant of the thickener, carrying out secondary solid-liquid separation, directly using secondary filtrate for preparing electrolyte, and entering the next step by using two filter cakes; (3) uniformly mixing the filter cake and concentrated sulfuric acid, and roasting; (4) adding a waste acid solution and industrial salt into the roasted anode mud, stirring and leaching for 3-5 hours, then carrying out solid-liquid separation, conveying a filter cake to a rare and precious process, and entering the next step of leaching liquid; (5) concentrating the leachate, cooling to 25-30 DEG C, crystallizing to obtain copper sulfate, and carrying out the next step on crystallization mother liquor; and (6) heating the crystallization mother liquor to 80-90 DEG C, introducing sulfur dioxide to reduce and precipitate rare and precious elements, filtering to obtain a rare and precious element mixture, and sending the rare and precious element mixture to a rare and precious process.

Description

Technical Field

[0001] The invention relates to the technical field of copper electrolysis anode mud treatment, in particular to a method for reducing the copper content in copper electrolysis anode mud and retaining rare and precious elements therein. Background Art

[0002] Copper electrolytic anode mud is an indispensable by-product in the copper electrolytic refining process. It is produced by the dissolution and precipitation of anode metal during the electrolysis process. It is mainly formed at the bottom of the electrolytic cell and is composed of metals and compounds in the anode crude metal that are insoluble in the electrolyte. The composition of copper anode mud depends on the composition of the copper anode, the casting quality and the control of electrolytic technical conditions. Its yield is generally 0.2% to 0.8%. Copper anode mud usually contains Au, Ag, Pb, Se, Te, As, Sb, Bi, Fe, S, Sn, Ni, Cu, SiO2, A12O3, platinum group elements and water. In the copper electrolytic production process, the average copper content of the anode mud produced by the permanent cathode copper electrolysis system reaches 25% or more.

[0003] The copper content in the anode mud of copper electrolysis mainly comes from two aspects: one is the copper impurities in the raw materials, which are brought into the anode mud during the electrolysis process; the other is that during the electrolysis process, some copper ions fail to react completely with the cathode plate and are deposited in the anode mud. Since the average copper content of the anode mud produced in the permanent cathode copper electrolysis system reaches more than 25%, the excessive copper content of the anode mud not only reduces the processing capacity of the subsequent anode mud treatment system, but also increases the production cost of the subsequent rare and precious process (because copper is an impurity element for rare and precious processes).

[0004] When the anode copper dissolves, the impurities contained in it undergo complex reactions to form a layer of black anode mud attached to the surface of the anode copper. When the copper anode containing impurities begins to dissolve, the concentration of copper ions on the anode surface in the electrolyte begins to increase. When the concentration of copper ions on the anode surface reaches saturation, it will gradually form in the pores of the anode mud in the form of copper sulfate. The formation of CuSO4•5H2O causes the copper concentration in the electrolyte to decrease, which will eventually lead to the re-dissolution of copper sulfate. Due to the uneven distribution of current on the anode surface, if the current in a certain area is too large, it will cause the copper sulfate generation in that area to be greater than the re-dissolution, the copper sulfate generation in that area will increase, and there will be a "passivation" trend in that area, which will reduce the current in that area and make the copper sulfate generation less than the re-dissolution. This cycle is repeated to achieve a balance between the generation and re-dissolution of the entire anode plate surface. As electrolysis proceeds, the anode mud layer on the anode surface gradually thickens. When the anode mud layer exceeds a certain thickness, the anode mud and copper sulfate are wrapped together and begin to fall to the bottom of the electrolytic cell. There is no circulating stirring of electrolyte at the bottom of the electrolytic cell, and the natural diffusion of copper ions is suppressed by gravity. In this way, the copper sulfate in the anode mud is basically retained until it enters the filtration system during the loading operation.

[0005] At present, the treatment of copper electrolysis anode mud is mainly carried out through the treatment system of the copper electrolysis system. The copper electrolysis anode mud treatment system includes a cache pit, a thickener, a primary filter, a secondary filter and a low-level tank. During the electrolysis loading operation, the copper anode mud passes through the thickener, primary filtration and secondary filtration together with the electrolyte. The filtered liquid is stored in the low-level tank and returned to the copper electrolysis cell for use as electrolyte. The filter cake is sent to the rare and precious process to further recover rare and precious elements such as lead, antimony, nickel and tin. After being treated by the above-mentioned treatment system, the copper content in the filter cake is still around 25%, which has basically not decreased and is far from the expected value of copper recovery. As a result, when the subsequent rare and precious process recovers other rare and precious elements, the copper content is too high, which brings about an increase in production costs. At the same time, for the permanent cathode copper electrolysis system, the copper content loss is relatively large. Therefore, improving the anode mud treatment process of the existing permanent cathode copper electrolysis system is the key to solving this problem. Summary of the invention

[0006] The purpose of the present invention is to provide a method for reducing the copper content in copper electrolysis anode mud and retaining the rare and precious elements therein, in order to address the problem that the current copper electrolysis anode mud treatment system cannot effectively reduce the copper content in the anode mud, which leads to a significant reduction in the treatment capacity of the subsequent anode mud treatment system and increases the production cost of the subsequent rare and precious elements treatment process.

[0007] A method for reducing the copper content in copper electrolysis anode mud and recovering rare elements therein of the present invention comprises the following steps: (1) Concentration: The copper electrolysis anode mud and electrolyte are placed in a thickener for concentration. After 3-4 hours of concentration, the supernatant in the thickener enters the supernatant buffer pool for standby use; (2) Rough extraction: The anode mud at the bottom of the thickener is pumped into a jacketed atmospheric pressure sealed reactor through a mud pump, and waste acid liquid containing 20-40% sulfuric acid by mass fraction is added. The solid-liquid ratio of anode mud: waste acid liquid is 1:3-5. The reactor is heated to 60-65°C, stirred for reaction for 3-5 hours, and pumped into a plate-frame filter for solid-liquid separation. The first filtrate is combined with the supernatant of the thickener and pumped into a plate-frame filter for secondary solid-liquid separation. The secondary filtrate can be directly used to prepare the electrolyte and stored in the electrolyte buffer pool for standby use. The two filter cakes are combined and entered into the next process. The filter cake obtained by rough extraction contains 15-17% copper by mass. (3) Calcination: After the filter cake and concentrated sulfuric acid are mixed evenly at a solid-liquid ratio of 1.25-1.67:1, they are transported to a calcination rotary kiln for calcination. The calcination temperature is set at 450°C-650°C and the calcination time is 5-8 hours. After calcination, the particle size of the anode mud is controlled at 10-200 meshes. (4) Leaching: Add the roasted anode mud into waste acid solution containing 20-40% sulfuric acid by mass, the solid-liquid ratio of anode mud to waste acid solution is 1:3-5, the leaching temperature is 80℃-90℃, and 5-7.5kg / m3 of industrial salt is added. 3 , stir and leach for 3~5h, solid-liquid separation, filter cake is sent to rare and precious process to extract rare and precious elements, and the leaching liquid enters the next process. The mass fraction of copper in the filter cake obtained by roasting-leaching solid-liquid separation is 0.1~0.3%; (5) Vacuum high-efficiency evaporation: The leaching solution is concentrated using a vacuum high-efficiency evaporator to a Baume degree of 36-42, and then cooled to 25-30°C to crystallize copper sulfate. The crystallization mother liquor enters the next process; (6) Reduction: The crystallization mother liquor is heated to 80°C to 90°C, and sulfur dioxide, a highly efficient reducing agent, is introduced at a flow rate of 50 to 80 L / min to reduce and precipitate the precious elements. The reactor is kept at a slightly positive pressure, and the reaction is carried out for 0.5 to 1 h. The precious elements mixture is filtered to obtain a precious element mixture, which is then sent to the precious element process for further separation of the precious elements. The filtrate is returned to step (5) for vacuum concentration and cooling crystallization to produce copper sulfate.

[0008] The mass fraction of copper in the copper electrolysis anode mud described in the present invention is ≥25%.

[0009] The waste acid solution containing 20-40% of sulfuric acid by mass in the present invention is the returned waste acid from the production of nickel sulfate (nickel may or may not be removed), which does not increase the process cost. Of course, only for the purpose of pursuing the copper removal effect, a newly prepared sulfuric acid solution or waste acid containing 20-40% of sulfuric acid by mass produced by other processes may also be used.

[0010] The atmospheric pressure sealed reaction kettle of the present invention is lined with titanium alloy.

[0011] The slight positive pressure in step (6) of the present invention is 100-500 Pa.

[0012] The process principle of the present invention is: First, the concentrated anode mud is passed through a sealed reactor with a jacket, and waste sulfuric acid solution is added. It is reacted under heating and stirring conditions at 60-65°C for 3-5 hours. At this time, about 8-10% of the copper in the anode mud is transferred to the solution. After solid-liquid separation, the anode mud contains 15-17% copper, and the rare and precious elements in the anode mud are not dissolved into the solution. The filtrate after the solid-liquid separation and the supernatant of the thickener are combined for secondary filtration. The secondary filtrate can be directly used as water for preparing the electrolyte, and the copper element can be directly recovered; the main reaction is 2Cu+O2+2H2SO4=2CuSO4+2H2O.

[0013] Then the filter cakes from the first filtration and the second filtration are combined, and concentrated sulfuric acid is added in proportion for roasting. Some elements in the anode mud can be oxidized into salts or oxides by roasting. The roasted anode mud is then treated with waste sulfuric acid solution and industrial salt under heating and stirring conditions of 80-90°C for 3-5 hours. At this time, almost all the copper elements in the anode mud enter the leaching solution, and there are still rare and precious elements in the filter cake that have not been leached. They are directly sent to the rare and precious process to extract the rare and precious elements. After the leaching solution is concentrated, cooled and crystallized to remove copper sulfate, sulfur dioxide, a high-efficiency reducing agent, is introduced to reduce and precipitate the rare and precious elements, and the rare and precious elements are sent to the rare and precious process for further extraction of the rare and precious elements. The main reaction is TeO2+2SO2+2H2O=2H2SO4+Te. At this point, the copper element contained in the anode mud is recovered to the maximum extent, and the rare and precious elements therein are also effectively recovered, without waste of resources, and the entire process flow is simple and easy to operate.

[0014] The beneficial effects of the present invention are: The crude extraction-roasting-re-leaching-concentration crystallization-reduction process of the present invention effectively reduces the copper content in the anode mud through crude extraction, and this part of copper directly enters the filtrate and is recovered by preparing an electrolyte. Since there is still 15-17% copper in the filter cake after crude extraction, the copper content in the anode mud is further reduced through subsequent roasting-re-leaching, and then sent to the rare and precious process to extract rare and precious elements, which reduces the work task of extracting precious metals in the rare and precious process. After the leachate is concentrated-cooled and crystallized to remove copper, reducing gas sulfur dioxide is introduced to precipitate the rare and precious elements, without causing waste of rare and precious elements. The method of the present invention has a simple process flow, simple operation, high element recovery rate, and no waste liquid or waste residue is discharged in the whole process, which is safe and environmentally friendly. DETAILED DESCRIPTION

[0015] The present invention will be further described below in conjunction with specific embodiments. It should be emphasized that the following description is merely exemplary and is not intended to limit the scope of the present invention.

[0016] Example 1 This embodiment takes the treatment method of copper electrolysis anode mud with a copper mass fraction of 25.1% as an example to explain the present invention in detail.

[0017] A method for reducing the copper content in copper electrolysis anode mud and recovering rare and precious elements therein comprises the following steps: (1) Concentration: The copper electrolysis anode mud and electrolyte are placed in a thickener for concentration. After 3.5 hours of concentration, the supernatant in the thickener enters the supernatant buffer tank for standby use; (2) Rough extraction: The anode mud at the bottom of the thickener is pumped into a jacketed atmospheric pressure sealed reactor through a mud pump, and waste acid liquid containing 25% sulfuric acid by mass is added. The solid-liquid ratio of anode mud: waste acid liquid is 1:4. The reactor is heated to 60-65°C, stirred and reacted for 4 hours, and pumped into a plate and frame filter for a solid-liquid separation. The first filtrate is combined with the supernatant of the thickener and pumped into a plate and frame filter for a second solid-liquid separation. The secondary filtrate can be directly used to prepare the electrolyte and stored in the electrolyte buffer pool for standby use; the two filter cakes are combined and entered into the next process. The mass fraction of copper in this filter cake is 16.2%; (3) Calcination: After the filter cake and concentrated sulfuric acid are mixed evenly at a solid-liquid ratio of 1.67:1, they are transported to a calcination rotary kiln for calcination. The calcination temperature is set at 500°C~600°C and the calcination time is 6.5 hours. After calcination, the particle size of the anode mud is controlled to be below 70 mesh; (4) Leaching: Add the roasted anode mud into waste acid solution containing 25% sulfuric acid by mass, the solid-liquid ratio of anode mud to waste acid solution is 1:4, the leaching temperature is 85℃~90℃, and 6kg / m3 of industrial salt is added. 3 , stir and leach for 3~5h, separate the solid and liquid, send the filter cake to the rare and precious process to extract rare and precious elements (the mass fraction of copper in this filter cake is 0.25%), and the leaching liquid enters the next process; (5) Vacuum high-efficiency evaporation: The leaching solution is concentrated by a vacuum high-efficiency evaporator until the solution reaches a Baume degree of 40, and then cooled to 25-28°C to crystallize copper sulfate. The crystallization mother liquor enters the next process; (6) Reduction: The crystallization mother liquor is heated to 85°C-90°C, and sulfur dioxide, a highly efficient reducing agent, is introduced at a flow rate of 65 L / min to reduce and precipitate the precious elements. The reactor is maintained at a slightly positive pressure (300-500 Pa). The reaction is carried out for 0.75 h, and the precious element mixture is filtered to obtain the precious element mixture, which is then sent to the precious element process for further separation of the precious elements. The filtrate is returned to step (5) for vacuum concentration and cooling crystallization to produce copper sulfate.

[0018] The waste acid solution containing 25% by mass of sulfuric acid described in this embodiment is the returned waste acid from the production of nickel sulfate.

[0019] The atmospheric pressure sealed reactor described in this embodiment is lined with titanium alloy (the same below).

[0020] Example 2 This embodiment takes the treatment method of copper electrolysis anode mud with a copper mass fraction of 26.3% as an example to explain the present invention in detail.

[0021] A method for reducing the copper content in copper electrolysis anode mud and recovering rare and precious elements therein comprises the following steps: (1) Concentration: The copper electrolysis anode mud and electrolyte are placed in a thickener for concentration. After 4 hours of concentration, the supernatant in the thickener enters the supernatant buffer tank for standby use; (2) Crude extraction: The anode mud at the bottom of the thickener is pumped into a jacketed atmospheric pressure sealed reactor through a mud pump, and waste acid liquid containing 30% sulfuric acid by mass is added. The solid-liquid ratio of anode mud: waste acid liquid is 1:5. The reactor is heated to 60-65°C, stirred and reacted for 5 hours, and pumped into a plate and frame filter for a solid-liquid separation. The primary filtrate is combined with the supernatant of the thickener and pumped into a plate and frame filter for a secondary solid-liquid separation. The secondary filtrate can be directly used to prepare the electrolyte and stored in the electrolyte buffer pool for standby use; the two filter cakes are combined and entered into the next process. The mass fraction of copper in this filter cake is 15.4%; (3) Calcination: After the filter cake and concentrated sulfuric acid are mixed evenly at a solid-liquid ratio of 1.25:1, they are transported to a roasting rotary kiln for roasting. The roasting temperature is set at 550°C~650°C and the roasting time is 5 hours. After roasting, the particle size of the anode mud is controlled to be below 100 mesh; (4) Leaching: Add the roasted anode mud into waste acid solution containing 30% sulfuric acid by mass, the solid-liquid ratio of anode mud to waste acid solution is 1:3, the leaching temperature is 80℃~85℃, and 5kg / m3 of industrial salt is added. 3 , stir and leach for 3 hours, separate the solid and liquid, send the filter cake to the rare and precious process to extract the rare and precious elements (the mass fraction of copper in this filter cake is 0.18%), and the leaching liquid enters the next process; (5) Vacuum high-efficiency evaporation: The leaching solution is concentrated using a vacuum high-efficiency evaporator, and the solution reaches a Baume degree of 42, and then cooled to 28-30°C to crystallize copper sulfate, and the crystallization mother liquor enters the next process; (6) Reduction: The crystallization mother liquor is heated to 80°C-85°C, and sulfur dioxide, a highly efficient reducing agent, is introduced at a flow rate of 80 L / min to reduce and precipitate the precious elements. The reactor is maintained at a slightly positive pressure (300-500 Pa). The reaction is carried out for 0.5 h, and the precious element mixture is filtered to obtain the precious element mixture, which is then sent to the precious element process for further separation of the precious elements. The filtrate is returned to step (5) for vacuum concentration and cooling crystallization to produce copper sulfate.

[0022] The waste acid solution containing 30% sulfuric acid by mass described in this embodiment is the returned waste acid from the production of nickel sulfate.

[0023] Example 3 This embodiment takes the treatment method of copper electrolysis anode mud with a copper mass fraction of 25.7% as an example to explain the present invention in detail.

[0024] A method for reducing the copper content in copper electrolysis anode mud and recovering rare and precious elements therein comprises the following steps: (1) Concentration: The copper electrolysis anode mud and electrolyte are placed in a thickener for concentration. After 3 hours of concentration, the supernatant in the thickener enters the supernatant buffer tank for standby use; (2) Rough extraction: The anode mud at the bottom of the thickener is pumped into a jacketed atmospheric pressure sealed reactor through a mud pump, and waste acid liquid containing 20% ​​sulfuric acid by mass is added. The solid-liquid ratio of anode mud: waste acid liquid is 1:3. The reactor is heated to 60-65°C, stirred and reacted for 3 hours, and pumped into a plate and frame filter for a solid-liquid separation. The first filtrate is combined with the supernatant of the thickener and pumped into a plate and frame filter for a second solid-liquid separation. The secondary filtrate can be directly used to prepare the electrolyte and stored in the electrolyte buffer pool for standby use; the two filter cakes are combined and entered into the next process. The mass fraction of copper in this filter cake is 17%; (3) Calcination: After the filter cake and concentrated sulfuric acid are mixed evenly at a solid-liquid ratio of 1.45:1, they are transported to a roasting rotary kiln for roasting. The roasting temperature is set at 450°C~550°C and the roasting time is 8 hours. After roasting, the particle size of the anode mud is controlled to be less than 10 mesh; (4) Leaching: Add the roasted anode mud into waste acid solution containing 20% ​​sulfuric acid by mass, the solid-liquid ratio of anode mud to waste acid solution is 1:5, the leaching temperature is 83℃~87℃, and 7.5kg / m3 of industrial salt is added. 3 , stir and leach for 5 hours, separate the solid and liquid, send the filter cake to the rare and precious process to extract rare and precious elements (the mass fraction of copper in this filter cake is 0.29%), and the leaching liquid enters the next process; (5) Vacuum high-efficiency evaporation: The leaching solution is concentrated using a vacuum high-efficiency evaporator to a solution degree of 36, and then cooled to 26-28°C to crystallize copper sulfate, and the crystallization mother liquor enters the next process; (6) Reduction: The crystallization mother liquor is heated to 82°C~87°C, and sulfur dioxide, a highly efficient reducing agent, is introduced at a flow rate of 70L / min to reduce and precipitate the precious elements. The reactor is maintained at a slightly positive pressure (200~400Pa). The reaction is carried out for 1 hour and filtered to obtain a mixture of precious elements. The mixture is sent to the precious process to further separate the precious elements. The filtrate is returned to step (5) for vacuum concentration and cooling crystallization to produce copper sulfate.

[0025] The waste acid solution containing 30% sulfuric acid by mass described in this embodiment is the returned waste acid from the production of nickel sulfate.

[0026] Example 4 This embodiment takes the treatment method of copper electrolysis anode mud with a copper mass fraction of 25.4% as an example to explain the present invention in detail.

[0027] A method for reducing the copper content in copper electrolysis anode mud and recovering rare and precious elements therein comprises the following steps: (1) Concentration: The copper electrolysis anode mud and electrolyte are placed in a thickener for concentration. After 3.2 hours of concentration, the supernatant in the thickener enters the supernatant buffer tank for standby use; (2) Rough extraction: The anode mud at the bottom of the thickener is pumped into a jacketed atmospheric pressure sealed reactor through a mud pump, and waste acid liquid containing 40% sulfuric acid by mass is added. The solid-liquid ratio of anode mud: waste acid liquid is 1:3. The reactor is heated to 60-65°C, stirred and reacted for 3 hours, and pumped into a plate and frame filter for a solid-liquid separation. The first filtrate is combined with the supernatant of the thickener and pumped into a plate and frame filter for a second solid-liquid separation. The secondary filtrate can be directly used to prepare the electrolyte and stored in the electrolyte buffer pool for standby use; the two filter cakes are combined and entered into the next process. The mass fraction of copper in this filter cake is 15%; (3) Calcination: After the filter cake and concentrated sulfuric acid are mixed evenly at a solid-liquid ratio of 1.56:1, they are transported to a roasting rotary kiln for roasting. The roasting temperature is set at 500°C~550°C and the roasting time is 7 hours. After roasting, the particle size of the anode mud is controlled to be below 30 mesh; (4) Leaching: Add the roasted anode mud into waste acid solution containing 40% sulfuric acid by mass, the solid-liquid ratio of anode mud to waste acid solution is 1:3, the leaching temperature is 80℃~85℃, and 6.5kg / m3 of industrial salt is added. 3 , stir and leach for 3 hours, separate the solid and liquid, send the filter cake to the rare and precious process to extract rare and precious elements (the mass fraction of copper in this filter cake is 0.12%), and the leaching liquid enters the next process; (5) Vacuum high-efficiency evaporation: The leaching solution is concentrated using a vacuum high-efficiency evaporator, and the solution reaches a Baume degree of 38, and then cooled to 25-27°C to crystallize copper sulfate, and the crystallization mother liquor enters the next process; (6) Reduction: The crystallization mother liquor is heated to 83°C~88°C, and sulfur dioxide, a highly efficient reducing agent, is introduced at a flow rate of 50L / min to reduce and precipitate the precious elements. The reactor is maintained at a slightly positive pressure (100~300Pa). The reaction is carried out for 1 hour and filtered to obtain a mixture of precious elements. The mixture is sent to the precious process to further separate the precious elements. The filtrate is returned to step (5) for vacuum concentration and cooling crystallization to produce copper sulfate.

[0028] The waste acid solution containing 40% sulfuric acid by mass described in this embodiment is the returned waste acid for producing nickel sulfate.

[0029] Through the treatment of the above four embodiments, the copper in the copper electrolysis anode mud is effectively reduced, and the rare elements therein are recovered, thereby achieving the purpose of recovering copper and further recovering rare elements. At the same time, the production method of the present invention does not discharge waste liquid and waste residue, the production process is environmentally friendly, and the economic benefit is high.

[0030] The above embodiments of the present invention are only examples for further illustrating the specific implementation of the present invention, and are not intended to limit the implementation of the present invention. Any replacement or modification made by a person skilled in the art within the technical scope disclosed by the present invention and in accordance with the content and spirit of the present invention shall be included in the protection scope of the present invention.

Claims

1. A method for reducing the copper content in copper electrolysis anode mud and recovering rare and precious elements therein, characterized in that The steps include: (1) Concentration: The copper electrolysis anode mud and electrolyte are placed in a thickener for concentration. After 3-4 hours of concentration, the supernatant in the thickener enters the supernatant buffer tank for standby use; (2) Rough extraction: The anode mud at the bottom of the thickener is pumped into a jacketed atmospheric pressure sealed reactor through a mud pump, and waste acid containing 20-40% sulfuric acid by mass fraction is added. The solid-liquid ratio of anode mud: waste acid is 1:3-5. The reactor is heated to 60-65°C, stirred for reaction for 3-5 hours, and pumped into a plate-frame filter for primary solid-liquid separation. The primary filtrate is combined with the supernatant of the thickener and pumped into a plate-frame filter for secondary solid-liquid separation. The secondary filtrate can be directly used to prepare the electrolyte and stored in the electrolyte buffer pool for standby use; the two filter cakes are combined and entered into the next process; (3) Calcination: After the filter cake and concentrated sulfuric acid are mixed evenly at a solid-liquid ratio of 1.25-1.67:1, they are transported to a calcination rotary kiln for calcination. The calcination temperature is set at 450°C-650°C and the calcination time is 5-8 hours. After calcination, the particle size of the anode mud is controlled at 10-200 meshes. (4) Leaching: Add the roasted anode mud into waste acid solution containing 20-40% sulfuric acid by mass, the solid-liquid ratio of anode mud to waste acid solution is 1:3-5, the leaching temperature is 80℃-90℃, and 5-7.5kg / m3 of industrial salt is added. 3 , stir and leach for 3~5h, separate the solid and liquid, send the filter cake to the rare and precious process to extract the rare and precious elements, and the leachate enters the next process; (5) Vacuum high-efficiency evaporation: The leaching solution is concentrated using a vacuum high-efficiency evaporator to a Baume degree of 36-42, and then cooled to 25-30°C to crystallize copper sulfate. The crystallization mother liquor enters the next process; (6) Reduction: The crystallization mother liquor is heated to 80°C to 90°C, and sulfur dioxide, a highly efficient reducing agent, is introduced at a flow rate of 50 to 80 L / min to reduce and precipitate the precious elements. The reactor is kept at a slightly positive pressure, and the reaction is carried out for 0.5 to 1 h. The precious elements mixture is filtered to obtain a precious element mixture, which is then sent to the precious element process for further separation of the precious elements. The filtrate is returned to step (5) for vacuum concentration and cooling crystallization to produce copper sulfate.

2. The method for reducing the copper content in copper electrolysis anode mud and recovering rare and precious elements therein according to claim 1, characterized in that: The mass fraction of copper in the copper electrolysis anode mud is ≥25%.

3. The method for reducing the copper content in copper electrolysis anode mud and recovering rare and precious elements therein according to claim 1, characterized in that: The waste acid liquid containing 20-40% sulfuric acid by mass is the returned waste acid from the production of nickel sulfate.

4. The method for reducing the copper content in copper electrolysis anode mud and recovering rare and precious elements therein according to claim 1, characterized in that: The atmospheric pressure sealed reactor is lined with titanium alloy.

5. The method for reducing the copper content in copper electrolysis anode mud and recovering rare and precious elements therein according to claim 1, characterized in that: The mass fraction of copper in the two filter cakes combined in step (2) is 15-17%.

6. The method for reducing the copper content in copper electrolysis anode mud and recovering rare and precious elements therein according to claim 1, characterized in that: The mass fraction of copper in the filter cake obtained by solid-liquid separation in step (4) is 0.1-0.3%.

7. The method for reducing the copper content in copper electrolysis anode mud and recovering rare and precious elements therein according to claim 1, characterized in that: The slight positive pressure in step (6) is 100-500 Pa.