Method for recovering tin and indium from lead ore refining slag

Abstract

A method for recovering tin and indium from lead ore refining slag, comprising: adding limestone and crushed coal to lead ore refining slag, uniformly mixing same, then feeding same into a side-blown furnace for smelting under oxygen-enriched air conditions, cooling collected slag, and then adding a sulfuric acid solution and sodium dodecyl sulfate for grinding to obtain slag powder; adding an acid solution to the slag powder for acid leaching, filtering and separating same to obtain a filtrate A and a residue B, drying and pulverizing the residue B, and then smelting and recovering same to obtain crude tin; adding sodium sulfite to the filtrate A for stirring treatment, then adding calcium carbonate to adjust the pH to 4.5-5.5, performing stirring treatment, and then filtering to obtain a filtrate C and a residue D; adding an acid solution to the residue D for secondary acid leaching to obtain a leaching solution, adding zinc powder into the leaching solution for displacement and recovery to obtain crude indium, and at the same time, collecting the post-displacement solution and mixing same with the filtrate C to prepare zinc sulfate. The method can improve the recovery efficiency of tin and indium elements in refining slag, thereby facilitating recovery and utilization of lead ore refining slag.

Description

A method for recovering tin fumes from lead ore refining waste Technical Field

[0001] This invention belongs to the field of lead ore refining waste recycling technology, specifically relating to a method for recovering tin and indium from lead ore refining waste. Background Technology

[0002] Lead ore refining is a complex chemical and physical process aimed at extracting pure lead metal from lead ore. Pyrometallurgical refining is one of the common methods for lead ore refining. It involves blowing air or oxygen into molten crude lead, or adding an oxidizing agent, to oxidize impurity metals and form oxide slag. Due to differences in density and chemical reaction, the oxide slag and other impurities form a slag layer. This slag layer either floats on the surface of the molten lead (forming scum) or sinks to the bottom (forming bottom slag). The scum and bottom slag are then separated from the molten lead by mechanical skimming or by draining the molten lead, resulting in high-purity molten lead. The separated scum and bottom slag form refining waste. The resulting refining waste also contains oxides and silicates of metals such as iron, copper, tin, antimony, arsenic, and silver, as well as sulfides of lead and other metals and precious metals. China is the world's largest producer of lead concentrate, accounting for approximately 43.7% of global production. Therefore, recycling this refining waste not only contributes to resource conservation and environmental protection but also creates economic benefits. However, the commonly used methods such as pyrometallurgical recovery, hydrometallurgical recovery, and thermal treatment-magnetic separation recovery are not very efficient at recovering tin and indium from refining waste, which is not conducive to the large-scale treatment of lead ore refining waste. Summary of the Invention

[0003] To address the aforementioned shortcomings, this invention discloses a method for recovering tin and indium from lead ore refining waste, which improves the recovery efficiency of tin and indium elements in the refining waste and is beneficial for the recycling and utilization of lead ore refining waste.

[0004] This invention is achieved using the following technical solution:

[0005] A method for recovering tin and indium from lead ore refining waste includes the following steps:

[0006] (1) Take lead ore refining waste residue, add limestone and crushed coal and mix evenly, then send it into a side-blown furnace. Then, oxygen-enriched air with an oxygen content of 50% is introduced and heated to 600-700℃ for smelting for 1-2 hours to obtain crude lead liquid and slag. After casting the crude lead liquid into ingots, it is sent to the electrolysis process for refining. At the same time, the slag is collected and rapidly cooled. The side-blown furnace uses oxygen-enriched side-blowing for smelting, which can reduce the amount of flue gas during the smelting process.

[0007] (2) Take the slag that has been rapidly cooled in step (1), add sulfuric acid solution and sodium dodecyl sulfate, mix them evenly, and then send them to a grinder for grinding to obtain slag powder. The weight ratio of the slag, sulfuric acid solution and sodium dodecyl sulfate is 100:(2~3):(1~2).

[0008] (3) Take the slag powder obtained in step (2), add acid solution to mix and obtain slurry for acid leaching. The liquid-solid ratio of the acid solution and slag powder is (10-15) L:1kg. The acid solution is a sulfuric acid solution with a concentration of 50-100g / L. The acid leaching pressure is 0.5-1.0MPa, the acid leaching temperature is 60-80℃, the pH value of the slurry is controlled at 1.5-2.0 during acid leaching, and the acid leaching time is 1-3h. After acid leaching, filter to separate filtrate A and filter residue B. Dry and crush filter residue B and then smelt and recover crude tin.

[0009] (4) Take the filtrate A from step (3) and add sodium sulfite and stir for 10-20 min. Then add calcium carbonate to adjust the pH of filtrate A to 4.5-5.5 and stir for 45-60 min. Filter to obtain filtrate C and filter residue D. Take the filter residue D and add the acid solution mentioned in step (3) to mix until the pH of the solution is 1.5-2.0. Continue stirring for 10-30 min and filter to obtain filtrate E. Take the filtrate E and add zinc powder to replace and recover crude indium. At the same time, collect the liquid phase after replacement to obtain the replacement liquid.

[0010] (5) Take the filtrate C obtained in step (4) and the replacement liquid, mix them evenly, and use them to prepare zinc sulfate.

[0011] As a further technical solution, in step (1) above, the weight ratio of lead ore refining waste residue, limestone and crushed coal is 100:(5-10):(20-30). By adjusting the ratio of lead ore refining waste residue, limestone and crushed coal, the smelting effect is improved, ensuring that the metal is fully melted, which is conducive to the separation of lead from other metals, and can also improve the fluidity of slag and improve the slag-forming effect.

[0012] As a further technical solution, in step (1) above, the slag is rapidly cooled at a rate of 10-20°C / min. Rapid cooling destroys the internal structure of the slag, which helps to grind and crush the slag and facilitates the leaching of metals such as tin and indium during the subsequent acid leaching process.

[0013] As a further technical solution, in step (2) above, the concentration of the sulfuric acid solution is 10-20%.

[0014] As a further technical solution, in step (2) above, the particle size of the slag powder is 80-100 mesh. Grinding the slag to 80-100 mesh can help the leaching of metals such as tin and indium in the subsequent acid leaching process and improve the leaching efficiency.

[0015] As a further technical solution, in step (3) above, potassium permanganate is added to the slurry during acid leaching, and the amount of potassium permanganate added is 1 to 3% of the weight of the slurry. Adding potassium permanganate can oxidize the existing elemental tin and form a precipitate that enters the slag, thereby improving the tin recovery rate.

[0016] As a further technical solution, in step (3) above, the slurry is stirred at a speed of 200-300 r / min during acid leaching. By controlling the stirring speed, the slag powder and acid solution can be fully contacted, thereby improving the metal leaching efficiency.

[0017] As a further technical solution, in step (4) above, filtrate A from step (3) is added to sodium sulfite and stirred for 10-20 minutes at a speed of 50-80 r / min. Adding sodium sulfite can reduce the residual ferric iron in filtrate A to ferrous iron, and after adding calcium carbonate to neutralize and adjust the pH value, indium can precipitate out before ferrous iron, thereby avoiding the influence of iron on indium precipitation and improving the indium recovery efficiency.

[0018] As a further technical solution, in step (5) above, the filtrate C obtained in step (4) and the replacement liquid are mixed evenly to obtain a mixed solution. Potassium permanganate is added to the mixed solution and the pH value of the mixed solution is controlled to be 5.0 to 5.5. After stirring for 20 to 30 minutes, zinc powder is added and reacted for 10 to 15 minutes. During the reaction process, the pH value is controlled to be 5.0 to 5.5. After the reaction is completed, the solution is filtered. The filtered liquid phase is concentrated by evaporation, crystallization, drying and pulverization to prepare zinc sulfate.

[0019] Compared with existing technologies, this technical solution has the following advantages:

[0020] 1. This invention employs a combination of pyrometallurgical and hydrometallurgical processes to treat lead ore refining waste, thereby achieving the goal of fully recovering tin and indium from the waste. First, pyrometallurgical smelting separates the high-content lead from the waste and enriches tin and indium in the slag. Then, the slag is crushed and acid-leached. By controlling the temperature, pressure, pH value, and other conditions of acid leaching, the leaching of metallic indium is promoted, while tin is also enriched in the acid-leached slag, thus achieving the effect of tin recovery. Then, calcium carbonate is added to the indium-containing leachate to adjust the pH value, causing indium to precipitate. The indium-containing precipitate is then subjected to a second acid leaching, followed by the addition of zinc powder for replacement. This not only effectively recovers indium but also improves the purity of indium.

[0021] 2. The method of the present invention is simple in process and easy to operate. It can effectively recover metals such as lead, zinc, tin and indium from lead ore refining waste, realize resource conservation and harmless treatment of waste, and help protect the environment and reduce pollution. Detailed Implementation

[0022] The present invention is further illustrated by the following examples, but these are not intended to limit the invention. Specific experimental conditions and methods not specified in the following examples are generally conventional methods well known to those skilled in the art. Example 1:

[0023] A method for recovering tin and indium from lead ore refining waste includes the following steps:

[0024] (1) Take lead ore refining waste residue, add limestone and crushed coal and mix evenly, then send it into a side-blown furnace. Then, oxygen-enriched air with an oxygen content of 50% is introduced and heated to 640℃ for smelting for 1.5h to obtain crude lead liquid and slag. After casting the crude lead liquid into ingots, send it to the electrolysis process for refining. At the same time, collect the slag and rapidly cool it at a rate of 18℃ / min. The weight ratio of lead ore refining waste residue, limestone and crushed coal is 100:6:25.

[0025] (2) Take the slag that has been rapidly cooled in step (1), add a 15% sulfuric acid solution and sodium dodecyl sulfate, mix them evenly, and then send them to a grinder for grinding to obtain slag powder with a particle size of 85 mesh. The weight ratio of slag, sulfuric acid solution and sodium dodecyl sulfate is 100:2.2:1.5.

[0026] (3) Take the slag powder obtained in step (2), add acid solution to mix and obtain slurry for acid leaching. The liquid-solid ratio of acid solution and slag powder is 12L:1kg. The acid solution is a sulfuric acid solution with a concentration of 50-100g / L. The acid leaching pressure is 0.8MPa, the acid leaching temperature is 65℃, the pH value of the slurry is controlled at 1.8 during acid leaching, and the acid leaching time is 2h. After acid leaching, filter to separate filtrate A and filter residue B. Dry and crush filter residue B and then smelt and recover crude tin. Add potassium permanganate to the slurry during acid leaching. The amount of potassium permanganate added is 2% of the weight of the slurry. Stir the slurry at a speed of 250r / min during acid leaching.

[0027] (4) Take the filtrate A from step (3) and add sodium sulfite. Stir and mix for 15 min at a speed of 60 r / min. Then add calcium carbonate to adjust the pH of filtrate A to 5.0 and stir for 50 min. Filter to obtain filtrate C and filter residue D. Take the filter residue D and add it to the acid solution from step (3) until the pH of the solution is 1.8. Continue stirring for 15 min and filter to obtain filtrate E. Take the filtrate E and add zinc powder to replace and recover crude indium. At the same time, collect the liquid phase after replacement to obtain the replacement liquid.

[0028] (5) Take the filtrate C obtained in step (4) and the replacement liquid and mix them evenly to obtain a mixture. Add potassium permanganate to the mixture and control the pH value of the mixture to 5.2. Stir for 25 minutes and then add zinc powder for 12 minutes. During the reaction, the pH value is controlled at 5.2. After the reaction is completed, filter the liquid phase. After evaporation, crystallization, drying and pulverization, zinc sulfate is prepared. Example 2:

[0029] A method for recovering tin and indium from lead ore refining waste includes the following steps:

[0030] (1) Take lead ore refining waste residue, add limestone and crushed coal and mix evenly, then send it into a side-blown furnace. Then, oxygen-enriched air with an oxygen content of 50% is introduced and heated to 600℃ for 2 hours to obtain crude lead liquid and slag. After casting the crude lead liquid into ingots, it is sent to the electrolysis process for refining. At the same time, the slag is collected and rapidly cooled at a rate of 10℃ / min. The weight ratio of lead ore refining waste residue, limestone and crushed coal is 100:5:20.

[0031] (2) Take the slag after rapid cooling in step (1), add a 10% sulfuric acid solution and sodium dodecyl sulfate, mix them evenly, and then send them to a grinder for grinding to obtain slag powder with a particle size of 80 mesh. The weight ratio of slag, sulfuric acid solution and sodium dodecyl sulfate is 100:2:1.

[0032] (3) Take the slag powder obtained in step (2), add acid solution to mix and obtain slurry for acid leaching. The liquid-solid ratio of acid solution and slag powder is 10L:1kg. The acid solution is a sulfuric acid solution with a concentration of 50g / L. The acid leaching pressure is 0.5MPa, the acid leaching temperature is 60℃, the pH value of the slurry is controlled at 1.5 during acid leaching, and the acid leaching time is 3h. After acid leaching, filter to separate filtrate A and filter residue B. Dry and crush filter residue B and then smelt and recover crude tin. Add potassium permanganate to the slurry during acid leaching. The amount of potassium permanganate added is 1% of the weight of the slurry. Stir the slurry at a speed of 200r / min during acid leaching.

[0033] (4) Take the filtrate A from step (3) and add sodium sulfite. Stir and mix for 10 min at a speed of 50 r / min. Then add calcium carbonate to adjust the pH of filtrate A to 4.5 and stir for 45 min. Filter to obtain filtrate C and filter residue D. Take the filter residue D and add it to the acid solution from step (3) until the pH of the solution is 1.5. Continue stirring for 10 min and filter to obtain filtrate E. Take the filtrate E and add zinc powder to replace and recover crude indium. At the same time, collect the liquid phase after replacement to obtain the replacement liquid.

[0034] (5) Take the filtrate C obtained in step (4) and the replacement liquid and mix them evenly to obtain a mixture. Add potassium permanganate to the mixture and control the pH value of the mixture to 5.0. Stir for 20 minutes and then add zinc powder for 10 minutes of reaction. During the reaction, the pH value is controlled at 5.0. After the reaction is completed, filter the liquid phase. After evaporation, crystallization, drying and pulverization, zinc sulfate is obtained. Example 3:

[0035] A method for recovering tin and indium from lead ore refining waste includes the following steps:

[0036] (1) Take lead ore refining waste residue, add limestone and crushed coal and mix evenly, then send it into a side-blown furnace. Then, oxygen-enriched air with an oxygen content of 50% is introduced and heated to 680℃ for smelting for 1.5h to obtain crude lead liquid and slag. After casting the crude lead liquid into ingots, it is sent to the electrolysis process for refining. At the same time, the slag is collected and rapidly cooled at a rate of 12℃ / min. The weight ratio of lead ore refining waste residue, limestone and crushed coal is 100:8:28.

[0037] (2) Take the slag that has been rapidly cooled in step (1), add 18% sulfuric acid solution and sodium dodecyl sulfate, mix them evenly, and then send them to a grinder for grinding to obtain slag powder with a particle size of 90 mesh. The weight ratio of slag, sulfuric acid solution and sodium dodecyl sulfate is 100:2.6:1.8.

[0038] (3) Take the slag powder obtained in step (2), add acid solution to mix and obtain slurry for acid leaching. The liquid-solid ratio of acid solution to slag powder is 18L:1kg. The acid solution is a sulfuric acid solution with a concentration of 80g / L. The acid leaching pressure is 0.8MPa, the acid leaching temperature is 75℃, the pH value of the slurry is controlled at 1.8 during acid leaching, and the acid leaching time is 2.5h. After acid leaching, filter to separate filtrate A and filter residue B. Dry and crush filter residue B and then smelt and recover crude tin. Add potassium permanganate to the slurry during acid leaching. The amount of potassium permanganate added is 2.5% of the weight of the slurry. Stir the slurry at a speed of 250r / min during acid leaching.

[0039] (4) Take the filtrate A from step (3) and add sodium sulfite. Stir and mix at a speed of 70 r / min for 15 min. Then add calcium carbonate to adjust the pH of filtrate A to 4.8 and stir for 55 min. Filter to obtain filtrate C and filter residue D. Take the filter residue D and add it to the acid solution from step (3) until the pH of the solution is 1.8. Continue stirring for 20 min and filter to obtain filtrate E. Take the filtrate E and add zinc powder to replace and recover crude indium. At the same time, collect the liquid phase after replacement to obtain the replacement liquid.

[0040] (5) Take the filtrate C obtained in step (4) and the replacement liquid and mix them evenly to obtain a mixture. Add potassium permanganate to the mixture and control the pH value of the mixture to 5.3. Stir for 250 min and then add zinc powder for 15 min of reaction. During the reaction, the pH value is controlled at 5.3. After the reaction is completed, filter the liquid phase. After evaporation, crystallization, drying and pulverization, zinc sulfate is prepared. Example 4:

[0041] A method for recovering tin and indium from lead ore refining waste includes the following steps:

[0042] (1) Take lead ore refining waste residue, add limestone and crushed coal and mix evenly, then send it into a side-blown furnace. Then, oxygen-enriched air with an oxygen content of 50% is introduced and heated to 700℃ for smelting for 1 hour to obtain crude lead liquid and slag. After casting the crude lead liquid into ingots, it is sent to the electrolysis process for refining. At the same time, the slag is collected and rapidly cooled at a rate of 20℃ / min. The weight ratio of lead ore refining waste residue, limestone and crushed coal is 100:10:30.

[0043] (2) Take the slag that has been rapidly cooled in step (1), add a 20% sulfuric acid solution and sodium dodecyl sulfate, mix them evenly, and then send them to a grinder for grinding to obtain slag powder with a particle size of 100 mesh. The weight ratio of slag, sulfuric acid solution and sodium dodecyl sulfate is 100:3:2.

[0044] (3) Take the slag powder obtained in step (2), add acid solution to mix and obtain slurry for acid leaching. The liquid-solid ratio of acid solution and slag powder is 15L:1kg. The acid solution is a sulfuric acid solution with a concentration of 100g / L. The acid leaching pressure is 1.0MPa, the acid leaching temperature is 80℃, the pH value of the slurry is controlled at 2.0 during acid leaching, and the acid leaching time is 1h. After acid leaching, filter to separate filtrate A and filter residue B. Dry and crush filter residue B and then smelt and recover crude tin. Add potassium permanganate to the slurry during acid leaching. The amount of potassium permanganate added is 3% of the weight of the slurry. Stir the slurry at a speed of 300r / min during acid leaching.

[0045] (4) Take the filtrate A from step (3) and add sodium sulfite. Stir and mix at a speed of 80 r / min for 10-20 min. Then add calcium carbonate to adjust the pH of filtrate A to 5.5 and stir for 60 min. Filter to obtain filtrate C and filter residue D. Take the filter residue D and add it to the acid solution from step (3) until the pH of the solution is 2.0. Continue stirring for 30 min and filter to obtain filtrate E. Take the filtrate E and add zinc powder to replace and recover crude indium. At the same time, collect the liquid phase after replacement to obtain the replacement liquid.

[0046] (5) Take the filtrate C obtained in step (4) and the replacement liquid and mix them evenly to obtain a mixture. Add potassium permanganate to the mixture and control the pH value of the mixture to 5.5. Stir for 30 min and then add zinc powder for 10-15 min of reaction treatment. During the reaction treatment, the pH value is controlled at 5.5. After the reaction treatment is completed, filter the liquid phase. After evaporation, crystallization, drying and pulverization, zinc sulfate is prepared.

[0047] Comparative Example 1: The method for recovering tin and indium from lead ore refining waste in this comparative example differs from the method in Example 1 only in that, in step (1), the slag obtained after smelting is naturally cooled at room temperature.

[0048] Comparative Example 2: The method for recovering tin and indium from lead ore refining waste in this comparative example differs from the method in Example 1 only in that sodium dodecyl sulfate is not added when grinding the slag in step (2).

[0049] Comparative Example 3: The method for recovering tin and indium from lead ore refining waste in this comparative example differs from the method in Example 1 only in that potassium permanganate is not added during the acid leaching process in step (3).

[0050] Comparative Example 4: The method for recovering tin and indium from lead ore refining waste in this comparative example differs from the method in Example 1 only in that sodium sulfite is not added to filtrate A for treatment in step (4).

[0051] Experimental Example: The same batch of lead ore refining slag was used for the experiment. The main components of the lead ore refining slag are shown in Table 1. The lead ore refining slag was evenly divided into 24 portions, with 3 portions forming a group, for a total of 8 groups of samples to be processed. Then, a group of samples to be processed was randomly selected according to the methods in Examples 1-4 and Comparative Examples 1-4, and the recovery rates of tin and indium were detected. The specific results are shown in Table 2.

[0052] Table 1. Main components of lead ore refining slag

[0053] Chemical composition: Pb, Sb, Ag, Cu, Bi, Sn content (%) 27.85, 9.12, 0.10, 1.04, 0.26, 0.68; Chemical composition: Zn, Fe, SiO2, CaO, S, In content (%) 7.59, 13.34, 8.07, 3.95, 0.30, 0.06

[0054] Table 2 Recovery rates (%) of tin and indium obtained by different methods

[0055] Example 1 Example 2 Example 3 Example 4 Tin recovery rate (%) 93.1 92.4 92.8 92.1 Indium recovery rate (%) 91.2 90.2 90.6 90.5 Comparative Example 1 Comparative Example 2 Comparative Example 3 Comparative Example 4 Tin recovery rate (%) 88.2 90.5 88.4 92.3 Indium recovery rate (%) 85.3 89.5 88.2 83.5

[0056] As can be seen from the data in Table 2, the present invention adopts a combination of pyrometallurgical and hydrometallurgical methods. By rapidly cooling the slag and adding agents such as sodium dodecyl sulfate, potassium permanganate, and sodium sulfite, tin and indium in lead ore refining waste can be effectively recovered.

[0057] Furthermore, it should be understood that although this specification describes embodiments, not every embodiment contains only one independent technical solution. This narrative style is merely for clarity. Those skilled in the art should consider the specification as a whole, and the technical solutions in each embodiment can also be appropriately combined to form other embodiments that can be understood by those skilled in the art.

Claims

1. A method for recovering tin and indium from lead ore refining waste, characterized in that: Includes the following steps: (1) Take lead ore refining waste residue, add limestone and crushed coal and mix evenly, then send it into a side blow furnace. Then, oxygen-enriched air with an oxygen content of 50% is introduced and heated to 600-700℃ for 1-2 hours to obtain crude lead liquid and slag. After casting the crude lead liquid into ingots, it is sent to the electrolysis process for refining. At the same time, the slag is collected and rapidly cooled. (2) Take the slag that has been rapidly cooled in step (1), add sulfuric acid solution and sodium dodecyl sulfate, mix them evenly, and then send them to a grinder for grinding to obtain slag powder. The weight ratio of the slag, sulfuric acid solution and sodium dodecyl sulfate is 100:(2~3):(1~2). (3) Take the slag powder obtained in step (2), add acid solution to mix and obtain slurry for acid leaching. The liquid-solid ratio of the acid solution and slag powder is (10-15) L:1kg. The acid solution is a sulfuric acid solution with a concentration of 50-100g / L. The acid leaching pressure is 0.5-1.0MPa, the acid leaching temperature is 60-80℃, the pH value of the slurry is controlled at 1.5-2.0 during acid leaching, and the acid leaching time is 1-3h. After acid leaching, filter to separate filtrate A and filter residue B. Dry and crush filter residue B and then smelt and recover crude tin. (4) Take the filtrate A from step (3) and add sodium sulfite and stir for 10-20 min. Then add calcium carbonate to adjust the pH of filtrate A to 4.5-5.5 and stir for 45-60 min. Filter to obtain filtrate C and filter residue D. Take the filter residue D and add the acid solution mentioned in step (3) to mix until the pH of the solution is 1.5-2.

0. Continue stirring for 10-30 min and filter to obtain filtrate E. Take the filtrate E and add zinc powder to replace and recover crude indium. At the same time, collect the liquid phase after replacement to obtain the replacement liquid. (5) Take the filtrate C obtained in step (4) and the replacement liquid, mix them evenly, and use them to prepare zinc sulfate.

2. The method for recovering tin and indium from lead ore refining waste according to claim 1, characterized in that: In step (1), the weight ratio of lead ore refining waste, limestone and crushed coal is 100:(5-10):(20-30).

3. The method for recovering tin and indium from lead ore refining waste according to claim 1, characterized in that: In step (1), the slag is rapidly cooled at a rate of 10-20℃ / min.

4. The method for recovering tin and indium from lead ore refining waste according to claim 1, characterized in that: In step (2), the concentration of the sulfuric acid solution is 10-20%.

5. The method for recovering tin and indium from lead ore refining waste according to claim 1, characterized in that: In step (2), the particle size of the slag powder is 80-100 mesh.

6. The method for recovering tin and indium from lead ore refining waste according to claim 1, characterized in that: In step (3), potassium permanganate is added to the slurry during acid leaching, and the amount of potassium permanganate added is 1 to 3% of the weight of the slurry.

7. The method for recovering tin and indium from lead ore refining waste according to claim 1, characterized in that: In step (3), the slurry is stirred at a speed of 200-300 r / min during acid leaching.

8. The method for recovering tin and indium from lead ore refining waste according to claim 1, characterized in that: In step (4), take the filtrate A from step (3) and add sodium sulfite. Stir and mix for 10 to 20 minutes at a speed of 50 to 80 r / min.

9. The method for recovering tin and indium from lead ore refining waste according to claim 1, characterized in that: In step (5), the filtrate C obtained in step (4) and the replacement liquid are mixed evenly to obtain a mixed solution. Potassium permanganate is added to the mixed solution and the pH value of the mixed solution is controlled to be 5.0-5.

5. After stirring for 20-30 minutes, zinc powder is added and reacted for 10-15 minutes. During the reaction, the pH value is controlled to be 5.0-5.

5. After the reaction is completed, the solution is filtered. The filtered liquid phase is concentrated by evaporation, crystallization, drying and pulverization to prepare zinc sulfate.