A method for selectively extracting and recovering tin from tin-containing residues

Abstract

The present disclosure provides a method for selectively extracting and recovering tin from tin-containing residue, the method comprising: ball milling the tin-containing residue, drying I, to obtain pretreated tin-containing residue; slurryizing the pretreated tin-containing residue with water to obtain slurryized tin-containing residue; mixing the slurryized tin-containing residue with a leaching agent solution, leaching, to obtain a tin-containing leaching solution and a leaching residue; dropping a solution containing a sulfidizing agent into the tin-containing leaching solution, precipitating, to obtain a purified solution and a purified residue; adding a sulfuric acid solution to the purified solution, reacting, to obtain a stannic acid precipitate and a mother liquor containing a complexing agent; washing the stannic acid precipitate, and then drying II, calcining the stannic acid precipitate, to obtain tin dioxide. In the method of the present disclosure, the tin leaching rate is greater than or equal to 95%, the complexing agent can be recycled, no high-salinity wastewater is generated, the process is short, the operation is simple, the cost is low, and the method is suitable for efficient recovery of tin from high-impurity and low-grade tin-containing residue.

Description

Technical Field

[0001] This disclosure relates to the fields of hydrometallurgy and secondary resource recovery technology, and more specifically, to a method for selectively extracting and recovering tin from tin-containing slag. Background Technology

[0002] Copper anode slime is an important byproduct of copper smelting electrolytic refining, rich in various valuable metals and rare elements, possessing extremely high resource recovery value. In the process of recovering precious metals such as gold, silver, platinum, and palladium from copper anode slime, pretreatment methods such as hydrochloric acid leaching are first required to dissolve large amounts of base metals such as copper, antimony, bismuth, and lead, thereby enriching the precious metals and creating conditions for subsequent extraction. The purified slag produced after the pretreatment liquid is purified and deposited still contains elements such as copper, arsenic, antimony, bismuth, and lead, as well as a small amount of tin. This purified slag is usually further processed to prepare sodium antimonate and bismuth concentrate. However, in this process, some tin is lost with the wastewater and waste residue, while some enters the products such as sodium antimonate as impurities, resulting in a waste of tin resources. Tin, as an important industrial metal, is widely used in many fields such as electronics, machinery, and chemicals, with stable market demand and high value.

[0003] Therefore, there is an urgent need to develop an efficient and specialized tin recovery technology for the high-impurity, low-grade tin-containing purification slag generated during the extraction of precious metals. This technology is of great practical significance for improving the comprehensive utilization efficiency of copper anode mud resources and reducing resource waste.

[0004] In existing technologies, some processes employ high-temperature smelting to separate tin through high-temperature volatilization. However, this method is energy-intensive and easily leads to tin forming a difficult-to-separate mixture with other impurity metals, resulting in high difficulty and cost for subsequent purification. Another process uses a multi-step acid-alkali leaching combined process, which can achieve tin leaching to a certain extent, but requires multiple steps of impurity removal, conversion, and leaching, making the process lengthy, complex, and difficult to treat the high-salt, highly polluting wastewater, posing significant environmental pressure. The vacuum oxidation volatilization method requires the construction of specialized equipment such as vacuum refining furnaces and sealed circulation systems, resulting in high equipment costs. Furthermore, the volatilization products are mostly oxidized mixtures of various impurity elements, requiring further separation and purification.

[0005] Prior art 1 (application number: 202310847698.X, application date: 2023.07.11) discloses a method for recovering tin from silver-separated slag by rotary kiln roasting and wet leaching, specifically including the following steps in sequence: S1) Mix copper anode mud, silver-separated slag, and additives in a certain proportion to form spherical material, mix the spherical material and carbonaceous material in a certain proportion, and then roast in a rotary kiln to obtain roasted slag; S2) Grind the roasted slag obtained in S1) into powder, add it to an alkali metal salt solution containing a leaching agent, and perform tin leaching to obtain tin leaching solution and tin leaching slag; S3) Add an oxidant and a calcium-containing reagent to the tin leaching solution obtained in S2) to precipitate tin to obtain tin-precipitated liquid and tin-precipitated slag, the tin-precipitated liquid is sent for antimony recovery, and the tin-precipitated slag is dried to obtain tin concentrate. This method employs rotary kiln roasting to transform tin in the raw materials through pyrometallurgical conversion, followed by leaching of tin through a complex system of leaching agent and alkali metal solution. The reaction system is complex.

[0006] Therefore, the problems of long process flow and high difficulty in impurity separation in the selective recovery of tin from tin-containing slag in the existing technology have become technical problems that urgently need to be solved in this field. Summary of the Invention

[0007] In view of this, this disclosure provides a method for selectively extracting and recovering tin from tin-containing slag with high impurities and low grade, in order to solve the problems of long process flow and high difficulty in impurity separation in the above-mentioned selective recovery of tin from tin-containing slag.

[0008] One aspect of this disclosure provides a method for selectively extracting and recovering tin from tin-containing slag, the method comprising:

[0009] The tin-containing slag is ball-milled and dried to obtain pretreated tin-containing slag; the pretreated tin-containing slag is slurried with water to obtain slurried tin-containing slag; the slurried tin-containing slag is mixed with a leaching agent solution, and water is added to control the liquid-solid ratio at 10:1 mL / g-15:1 mL / g to obtain a mixed solution; the pH value of the mixed solution is controlled to be stable at 3.5-7.0, and leaching is performed to obtain a tin-containing leaching solution and leaching slag; the tin-containing slag includes copper, arsenic, antimony, bismuth, lead, and tin, wherein the tin content is 2%-5%; the concentration of the leaching agent in the leaching agent solution is 40 g / L-60 g / L. The leaching agent comprises an amine carboxyl complexing agent selected from one or more of diethylenetriaminepentaacetic acid and hypozonyltriacetic acid; the amount of the leaching agent added is 1.2 to 2.0 times the theoretical amount of total tin in the pretreated tin-containing slag, in terms of molar amount; the leaching temperature is 30℃-70℃, and the leaching time is 1h-3h; controlling the pH value of the mixture to be stable at 3.5-7.0 includes: adding sulfuric acid with a concentration of 10wt%-15wt% or ammonia water with a concentration of 5wt%-10wt% to the mixture to control the pH value of the mixture to be stable at 3.5-7.0;

[0010] A solution containing a sulfiding agent is added dropwise to the tin-containing leaching solution to obtain a reaction solution, followed by precipitation to obtain a purified solution and a purified residue. The sulfiding agent is selected from NaHS and Na2S. The mass concentration of the sulfiding agent in the solution is 0.5wt%-2wt%. The precipitation temperature is 40℃-50℃, and the precipitation time is 1h-2h. The amount of sulfiding agent added is 1.2-1.5 times the copper content in the tin-containing leaching solution. During the precipitation process, sulfuric acid with a concentration of 10wt%-15wt% or ammonia water with a concentration of 5wt%-10wt% is added to the reaction solution to control the pH value of the reaction solution to 3-5.

[0011] A 10wt%-15wt% sulfuric acid solution is added to the purification solution to obtain an acidified purification solution with a pH of 1-2. The reaction proceeds to obtain stannic acid precipitate and a mother liquor containing a complexing agent. The reaction temperature is 50℃-80℃, and the reaction time is 40min-60min. A 5wt%-10wt% ammonia solution is added to the mother liquor containing the complexing agent to adjust its pH to 3.5-7.0, thereby regenerating the complexing agent.

[0012] The stannic acid precipitate is washed with water until the solution pH value is 7-8, and then the stannic acid precipitate is subjected to drying II and calcination to obtain tin dioxide. The drying II temperature is 80℃-100℃ and the drying II time is 6h-10h; the calcination temperature is 600℃-700℃ and the calcination time is 2h-3h.

[0013] Optionally, the liquid-to-solid ratio is independently selected from any value among 10:1 mL / g, 12:1 mL / g, 15:1 mL / g, or any range between any two of the above points.

[0014] Optionally, the pH value of the mixture is independently selected from any value among 3.5, 5, 6.5, 7.0 or any range between any two of the above points.

[0015] Optionally, the concentration of the leachate is independently selected from any value among 40 g / L, 45 g / L, 50 g / L, and 60 g / L, or any range between any two of the above.

[0016] Optionally, the amount of leaching agent added is independently selected from any value among 1.2 times, 1.4 times, 1.6 times, 1.8 times, and 2.0 times the theoretical total tin content, or any range between any two of the above points.

[0017] Optionally, the amount of the sulfiding agent added is independently selected from any value among 1.2 times, 1.3 times, 1.4 times, and 1.5 times the copper content in the tin-containing leaching solution, or any range between any two of the above points.

[0018] Optionally, the pH value of the reaction solution is independently selected from any value among 3, 4, and 5, or any range between any two of the above points.

[0019] Optionally, the pH value of the acidification and purification solution is independently selected from any value among 1, 1.5, and 2, or any range between any two of the above points.

[0020] Optionally, the temperature of the drying process I is 80℃-100℃, and the drying time is 2h-5h.

[0021] Optionally, the leaching is carried out under stirring conditions, wherein the stirring speed is 200 r / min-350 r / min.

[0022] Optionally, the precipitation is carried out under stirring conditions, and the stirring speed is 200 r / min-250 r / min.

[0023] Optionally, the reaction is carried out under stirring conditions, wherein the stirring speed is 100 r / min-200 r / min.

[0024] Compared with existing technologies, the method for selectively extracting and recovering tin from tin-containing slag provided in this disclosure achieves at least the following beneficial effects:

[0025] First, this disclosure achieves efficient and selective leaching of tin and solidification and separation of impurity elements through a two-step method of "amine-carboxyl complexation leaching-sulfidation impurity removal". At the same time, the leaching reagent can be recycled, significantly reducing the process cost and meeting the needs of green metallurgy.

[0026] Secondly, in the method disclosed herein, amine carboxyl complexing agents such as diethylenetriaminepentaacetic acid (DTPA) and nitrotriacetic acid (NTA) can form complexes with tin ions, and the complexation stability constant is much higher than that with impurity ions such as antimony, bismuth, and lead. Tin can be selectively leached in a weakly acidic to neutral system, with a tin leaching rate of greater than or equal to 95%.

[0027] Third, in the method disclosed herein, the leaching process can effectively suppress the dissolution of impurities. The leaching rates of impurities such as antimony, bismuth, and lead are all less than 1%, and only trace amounts of copper are leached (2%-10%). Subsequent sulfidation to remove copper can remove copper to below 5 mg / L (less than 0.005 g / L), which greatly reduces the difficulty of impurity purification and shortens the purification process.

[0028] Fourth, in the method disclosed herein, the amine carboxyl complexing agent can be recycled through acidification-complex breaking-ammonia water neutralization and pH adjustment, which significantly reduces the reagent consumption cost of the leaching process, generates no high-salt wastewater throughout the process, and the leaching residue can be used for further recovery of valuable metals, which is in line with the concept of circular economy.

[0029] Of course, any product implementing this disclosure does not necessarily need to achieve all of the technical effects described above at the same time.

[0030] Other features and advantages of the invention will become clear from the following detailed description of exemplary embodiments of the invention with reference to the accompanying drawings. Attached Figure Description

[0031] The accompanying drawings, which are incorporated in and form part of this specification, illustrate embodiments of the invention and, together with their description, serve to explain the principles of the invention.

[0032] Figure 1 This is a flowchart of a method for selectively extracting and recovering tin from tin-containing slag;

[0033] Figure 2 This is a flowchart of the method for selectively extracting and recovering tin from tin-containing dross according to Embodiment 1 of this disclosure;

[0034] Figure 3 This is a flowchart of the method for selectively extracting and recovering tin from tin-containing slag according to Embodiment 2 of this disclosure;

[0035] Figure 4 This is a flowchart of the method for selectively extracting and recovering tin from tin-containing slag according to Embodiment 2 of this disclosure. Detailed Implementation

[0036] Various exemplary embodiments of the present invention will now be described in detail with reference to the accompanying drawings. It should be noted that, unless otherwise specifically stated, the relative arrangement, numerical expressions, and values ​​of the components and steps set forth in these embodiments do not limit the scope of the invention.

[0037] The following description of at least one exemplary embodiment is merely illustrative and is in no way intended to limit the invention or its application or use.

[0038] Techniques, methods, and equipment known to those skilled in the art may not be discussed in detail, but where appropriate, such techniques, methods, and equipment should be considered part of the specification.

[0039] In all the examples shown and discussed herein, any specific values ​​should be interpreted as merely exemplary and not as limitations. Therefore, other examples of exemplary embodiments may have different values.

[0040] It should be noted that similar labels and letters in the following figures indicate similar items; therefore, once an item is defined in one figure, it does not need to be discussed further in subsequent figures.

[0041] In existing technologies, for the high-impurity, low-grade tin-containing purification slag generated during the extraction of precious metals, some processes employ high-temperature smelting to separate tin through high-temperature volatilization. However, this method is energy-intensive and easily leads to the formation of difficult-to-separate mixtures of tin and other impurity metals, resulting in high difficulty and cost for subsequent purification. Other processes use a multi-step acid-alkali leaching combined process, which can achieve tin leaching to a certain extent, but requires multiple steps of impurity removal, conversion, and leaching, making the process lengthy, complex, and generating high-salt, highly polluting wastewater that is difficult to treat, posing a significant environmental burden. Vacuum oxidation volatilization requires specialized equipment such as vacuum refining furnaces and sealed circulation systems, resulting in high equipment costs, and the volatilization products are mostly oxidized mixtures of various impurity elements, requiring further separation and purification. To address the shortcomings of these existing technologies, this disclosure proposes a method for selectively extracting and recovering tin from high-impurity, low-grade tin-containing slag containing impurities such as copper, arsenic, antimony, bismuth, and lead. This method aims to solve the problems of long process flows, high difficulty in impurity separation, significant environmental burden, and high costs associated with existing processes.

[0042] This disclosure provides a method for selectively extracting and recovering tin from high-impurity, low-grade tin-containing slag, such as... Figure 1 As shown, it includes the following steps:

[0043] S100. Selective Leaching of Tin: High-impurity, low-grade tin-containing slag is ball-milled to break up some of the uneven, blocky materials in the slag, resulting in uniform particle size (specifically, uniform granules or powder). The slag is then dried at 100℃ for 2-5 hours to obtain pretreated tin-containing slag. A leaching agent is dissolved in water to prepare a solution of 40 g / L-60 g / L, and stirred to ensure uniform dispersion. The leaching agent is selected from one or a mixture of diethylenetriaminepentaacetic acid (DTPA) and nitric acid triacetic acid (NTA). The amount of leaching agent added is 1.2-2.0 times (in moles) of the theoretical total tin content in the pretreated tin-containing slag. The pretreated tin-containing slag is slurried with water and stirred until homogeneous. The prepared leaching agent solution is added to the slurried tin-containing slag, and water is added again to control the liquid-solid ratio at 10-15:1 (mL / g). The pH value of the mixture is tested. If the pH value of the mixture is not within the range of 3.5-7.0, an appropriate amount of 10% dilute sulfuric acid or 5% dilute ammonia solution is added as a pH adjuster to stabilize the pH value of the solution at 3.5-7.0. The leaching is carried out at 30-70℃ and a stirring speed of 200-350 r / min for 1-3 hours. After the reaction is completed, solid-liquid separation is performed to obtain tin-containing leaching solution and leaching residue. The leaching residue is enriched with impurities such as antimony, bismuth, lead, and arsenic, which can be further processed and recycled.

[0044] S200, Deep Copper Removal: During the tin leaching process, a small amount of copper is leached into the liquid phase. The tin-containing leaching solution is transferred to a copper removal reaction tank, and 0.5wt%-2wt% of dilute NaHS (or Na2S) solution is slowly added dropwise. The amount of NaHS (or Na2S) added is 1.2-1.5 times the copper content in the leaching solution (in molars). The system temperature is maintained at 40-50℃ and the pH at 3-5. The stirring speed is controlled at 200r / min-250r / min, and the reaction is carried out for 1-2 hours. The solid and liquid are separated to obtain purified solution and purified residue. It should be noted that because the concentration of copper in the liquid phase is not high, the addition rate of NaHS or Na2S needs to be slowed down. During the reaction, the pH is monitored in real time using a pH meter. When the pH is too low, 5% dilute ammonia is slowly added dropwise to adjust it. When the pH is too high, 10% dilute sulfuric acid is added dropwise to adjust it, controlling the pH between 3 and 5.

[0045] S300, Acidification and Complex Breaking and Stannic Acid Precipitation: Slowly add 10wt%-15wt% dilute sulfuric acid solution to the purification solution to adjust the pH of the system to 1-2, so that the tin complex is completely dissociated and stannic acid precipitate is generated, and the amine carboxyl complexing agent is converted into a free state and enters the solution; control the temperature at 50~80℃, stir the reaction at 100~200r / min for 40-60min, and separate the solid and liquid to obtain stannic acid precipitate and mother liquor containing complexing agent.

[0046] S400, Preparation of Tin Dioxide by Stannic Acid Calcination: The stannic acid precipitate is washed with deionized water until the solution pH is neutral to remove the complexing agent and impurities adsorbed on the surface. Then, the precipitate is dried in an oven at 80-100℃ for 6-10 hours and then transferred to a calcination furnace. The calcination process is controlled at 600-700℃ and calcined at a constant temperature for 2-3 hours. After crushing and sieving, the tin dioxide product is obtained.

[0047] S500: Add 5%-10% dilute ammonia to the mother liquor containing the complexing agent to adjust the pH value to 3.5-7.0, thereby regenerating the complexing agent. The regenerated complexing agent is then returned to step S100 for recycling.

[0048] In the entire tin recovery process, during the leaching process, the tin dioxide in the raw material first dissociates into Sn under the action of a complexing agent. 4+ Sn 4+ It immediately forms a soluble and stable chelate complex with the active complexing agent anion, and the reaction promotes the continuous leaching of tin compounds. The main chemical reactions that occur are (taking DTPA as an example):

[0049] SnO2+DTPA 5- +4H + =SnDTPA - +2H2O;

[0050] The copper hydroxide in the raw material slightly dissociates into free Cu at this acidity. 2+ A small amount of Cu 2+ It is leached by complexing with complexing agents:

[0051] Cu(OH)2+DTPA 5- +2H + =CuDTPA 3- +2H2O;

[0052] The arsenic, antimony, bismuth, lead and other substances in the raw materials remain in the slag phase because their complexation constants are much lower than those of tin and they are difficult to dissociate into free ions at this pH.

[0053] The main chemical reactions that occur during the deep copper removal process are:

[0054] CuDTPA 3- +S 2- =CuS↓+DTPA 5- ;

[0055] During the acidification, complex breakdown, and precipitation process, DTPA is protonated under strong acid conditions, and the main chemical reaction that leads to the regeneration of molecular H5DTPA is as follows:

[0056] SnDTPA - +5H + =Sn 4++H5DTPA;

[0057] Sn 4+ +3H₂O=H₂SnO₃↓+4H + .

[0058] Example 1

[0059] Reference Figure 2 This embodiment employs a method for selectively extracting and recovering tin from high-impurity, low-grade tin-containing slag, comprising:

[0060] S101. Selective leaching of tin: Take 1 kg of high-impurity, low-grade tin-containing slag, which contains 10.65% Cu, 15.24% Sb, 9.72% As, 3.88% Sn, 2.06% Pb, 8.49% Bi, and 38.16% water. After ball milling, dry at 100℃ for 2 h to obtain pretreated tin-containing slag.

[0061] Weigh 95.2g of diethylenetriaminepentaacetic acid (DTPA) as the leaching agent (1.2 times the theoretical amount), add 2L of water to dissolve and stir evenly to prepare a 45g / L solution.

[0062] The pretreated tin-containing slag was slurried with 1L of water and stirred until homogeneous. The prepared leaching agent solution was then added to the slurried tin-containing slag, followed by the addition of 2.7L of water to control the liquid-to-solid ratio at 10:1 (mL / g). At this point, the solution pH was 5.8. The pH was adjusted to 3.5 using 10wt% dilute sulfuric acid. The leaching temperature was controlled at 30℃ and the stirring speed at 200r / min. After leaching for 1 hour, solid-liquid separation was performed to obtain tin-containing leaching solution and leaching slag. The tin leaching rate was tested to be 95.16%, and the leaching rates of impurities such as antimony, bismuth, lead, and arsenic were all less than 1%. The copper leaching rate was 5.25%.

[0063] S201. Deep Copper Removal: The tin-containing leaching solution is transferred to a copper removal reaction tank. A 0.5wt% dilute NaHS solution is slowly added dropwise at a volume equivalent to 1.2 times the copper content in the leaching solution. The system temperature is maintained at 40℃ and the pH at 3. During the reaction, the pH is monitored in real time using a pH meter. If the pH is too low, 5% dilute ammonia is slowly added dropwise to adjust it; if the pH is too high, 10% dilute sulfuric acid is added dropwise to adjust it, controlling the pH at 3. The stirring speed during the reaction is 200 r / min. After 1 hour of reaction, solid-liquid separation is performed to obtain a purified solution. Analysis shows that the copper content in the purified solution is less than 0.005 g / L.

[0064] S301, Acidification and Complex Breaking and Stannic Acid Precipitation: Slowly add 10wt% dilute sulfuric acid solution to the purified solution, adjust the pH of the system to 1, control the temperature at 50℃, stir the reaction for 40 min, and separate the solid and liquid to obtain stannic acid precipitate and mother liquor containing complexing agent.

[0065] S401. Preparation of tin dioxide by calcination of stannic acid: The stannic acid precipitate is washed with deionized water until the pH is neutral, dried in an oven at 100°C for 6 hours, and then transferred to a calcination furnace and calcined at a constant temperature of 600°C for 2 hours. After crushing and sieving, tin dioxide product is obtained. The tin recovery rate is 94.21% from low-grade tin-containing slag to tin dioxide.

[0066] S501. Add 5wt% dilute ammonia to the mother liquor containing the complexing agent to adjust the pH to 3.5 to regenerate DTPA. The regenerated DTPA can be returned to the S100 leaching process for use.

[0067] Example 2

[0068] Reference Figure 3 This embodiment employs a method for selectively extracting and recovering tin from high-impurity, low-grade tin-containing slag, comprising:

[0069] S102. Selective leaching of tin: Take 1 kg of high-impurity, low-grade tin-containing slag, which contains 13.15% Cu, 14.16% Sb, 7.82% As, 2.98% Sn, 4.06% Pb, 6.57% Bi, and 41.02% water. After ball milling, dry at 100℃ for 3.5 h to obtain pretreated tin-containing slag.

[0070] Weigh 47.3g of NTA as the leaching agent (1.6 times the theoretical amount), add 790mL of water to dissolve and stir evenly to prepare a 60g / L solution.

[0071] The pretreated tin-containing slag was slurried with 2L of water and stirred until homogeneous. The prepared leaching agent solution was then added to the slurried tin-containing slag, and 3L of water was added to control the liquid-solid ratio at 12:1 (mL / g). At this point, the pH of the solution was 3.3. The pH of the solution was adjusted to 5 with 5wt% dilute ammonia. The leaching temperature was controlled at 50℃ and the stirring speed at 300r / min. After leaching for 2 hours, solid-liquid separation was achieved. The tin leaching rate reached 97.05%, and the leaching rates of impurities such as antimony, bismuth, lead, and arsenic were all less than 1%. The copper leaching rate was 7.46%.

[0072] S202. Deep Copper Removal: Add 1.25wt% NaHS dilute solution dropwise to the tin-containing leaching solution at 1.3 times the copper content, maintaining the system temperature at 45℃ and pH at 4. During the reaction, monitor the pH in real time using a pH meter. If the pH is too low, slowly add 5% dilute ammonia to adjust it; if the pH is too high, add 10% dilute sulfuric acid to adjust it, controlling the pH to 3. The stirring speed during the reaction is 200 r / min. After 1.5 h of reaction, solid and liquid are separated to obtain the purified solution. Analysis shows that the copper content in the purified solution is less than 0.005 g / L.

[0073] S302, Acidification and Complex Breaking and Stannic Acid Precipitation: Add 10wt% dilute sulfuric acid to the purified solution to adjust the pH to 1.5, control the temperature at 65℃, stir the reaction for 50 min, and separate the solid and liquid to obtain stannic acid precipitate and mother liquor containing complexing agent.

[0074] S402, Preparation of tin dioxide by calcination of stannic acid: The stannic acid precipitate is washed to neutral, dried in an oven at 100°C for 8 hours, and then calcined at a constant temperature of 650°C for 2.5 hours. After crushing and sieving, tin dioxide product is obtained. The tin recovery rate is 96.08% from low-grade tin-containing slag to tin dioxide.

[0075] S502: The mother liquor containing the complexing agent is adjusted to pH 5 with 5% dilute ammonia water and then regenerated to NTA. The regenerated NTA can be returned to the S100 leaching process for use.

[0076] Example 3

[0077] Reference Figure 4 This embodiment employs a method for selectively extracting and recovering tin from high-impurity, low-grade tin-containing slag, comprising:

[0078] S103. Selective leaching of tin: Take 1 kg of high-impurity, low-grade tin-containing slag, the composition of which is the same as in Example 1, ball mill it, and dry it at 100℃ for 5 h to obtain pretreated tin-containing slag. Weigh 158.62 g (2.0 equivalents) of diethylenetriaminepentaacetic acid (DTPA), add 4 L of water to dissolve and stir evenly to prepare a 40 g / L solution.

[0079] The pretreated tin-containing slag was slurried with 1L of water and stirred until homogeneous. The prepared leaching agent solution was then added to the slurried tin-containing slag, followed by the addition of 850mL of water to control the liquid-to-solid ratio at 15:1 (mL / g). At this point, the solution pH was 4.2. The pH was adjusted to 6.5 with 5wt% dilute ammonia. The leaching temperature was controlled at 70℃ and the stirring speed at 350r / min. After leaching for 3 hours, solid-liquid separation was achieved. The tin leaching rate was 96.1%, while the leaching rates of impurities such as antimony, bismuth, lead, and arsenic were all less than 1%. The copper leaching rate was 9.55%.

[0080] S203, Deep Copper Removal: Add 2wt% Na₂S dilute solution dropwise to the tin-containing leaching solution at 1.5 times the copper content, maintaining a temperature of 50℃ and a pH of 5. During the reaction, monitor the pH in real time using a pH meter. If the pH is too low, slowly add 5% dilute ammonia to adjust it; if the pH is too high, add 10% dilute sulfuric acid to adjust it, controlling the pH to 3. The stirring speed during the reaction is 250 r / min. After 2 hours of reaction, solid-liquid separation is performed to obtain the purified solution. Analysis shows that the copper content in the purified solution is less than 0.005 g / L.

[0081] S303, Acidification and Complex Breaking and Stannic Acid Precipitation: Add 10wt% dilute sulfuric acid to the purified solution to adjust the pH to 2, control the temperature at 80℃, stir the reaction for 60 min, and separate the solid and liquid to obtain stannic acid precipitate and mother liquor containing complexing agent.

[0082] S403, Preparation of tin dioxide by calcination of stannic acid: The stannic acid precipitate is washed until neutral, dried in an oven at 80°C for 6 hours, and then calcined at a constant temperature of 700°C for 3 hours. After crushing and sieving, tin dioxide product is obtained. The tin recovery rate is 95.13% from low-grade tin-containing slag to tin dioxide.

[0083] S503: The mother liquor containing the complexing agent is adjusted to pH 7.0 with 5wt% dilute ammonia water to regenerate DTPA. The regenerated DTPA can be returned to the S100 leaching process for use.

[0084] As can be seen from the above embodiments, the method for selectively extracting and recovering tin from high-impurity, low-grade tin-containing slag provided in this disclosure achieves at least the following beneficial effects:

[0085] 1. This disclosure achieves efficient and selective leaching of tin and solidification and separation of impurity elements through a two-step method of "amine-carboxyl complexation leaching - sulfidation impurity removal". At the same time, the leaching reagent can be recycled, significantly reducing the process cost and meeting the requirements of green metallurgy.

[0086] 2. In the method disclosed herein, amine carboxyl complexing agents such as DTPA and NTA can form complexes with tin ions, and the complexation stability constant is much higher than that with impurity ions such as antimony, bismuth, and lead. This allows for selective leaching of tin in weakly acidic to neutral systems, with a tin leaching rate greater than or equal to 95%. For example, in Example 1, the tin leaching rate was 95.16%; in Example 2, the tin leaching rate reached 97.05%; and in Example 3, the tin leaching rate was 96.1%.

[0087] 3. In the method disclosed herein, the leaching process can effectively suppress the dissolution of impurities. The leaching rates of impurities such as antimony, bismuth, and lead are all less than 1%, and only trace amounts of copper are leached (2%-10%). Subsequent sulfidation to remove copper can remove copper to below 5 mg / L (less than 0.005 g / L), which greatly reduces the difficulty of impurity purification and shortens the purification process.

[0088] 4. In the method disclosed herein, the amine carboxyl complexing agent can be recycled through acidification-complex breaking-ammonia water neutralization and pH adjustment, which significantly reduces the reagent consumption cost of the leaching process. No high-salt wastewater is generated throughout the process, and the leaching residue can be used for further recovery of valuable metals, which is in line with the concept of circular economy.

[0089] While specific embodiments of the invention have been described in detail by way of examples, those skilled in the art should understand that the examples are for illustrative purposes only and not intended to limit the scope of the invention. Those skilled in the art should understand that modifications can be made to the above embodiments without departing from the scope and spirit of the invention. The scope of the invention is defined by the appended claims.

Claims

1. A method for selectively extracting and recovering tin from tin-containing slag, characterized in that, The method includes: The tin-containing slag is ball-milled and dried to obtain pretreated tin-containing slag; the pretreated tin-containing slag is slurried with water to obtain slurried tin-containing slag; the slurried tin-containing slag is mixed with a leaching agent solution, and water is added to control the liquid-solid ratio at 10:1 mL / g-15:1 mL / g to obtain a mixed solution; the pH value of the mixed solution is controlled to be stable at 3.5-7.0, and leaching is performed to obtain a tin-containing leaching solution and leaching slag; the tin-containing slag includes copper, arsenic, antimony, bismuth, lead, and tin, wherein the tin content is 2%-5%; the concentration of the leaching agent in the leaching agent solution is 40g / L-60g / L. L, the leaching agent comprises an amine-carboxyl complexing agent selected from one or more of diethylenetriaminepentaacetic acid and hypozonyltriacetic acid; the amount of the leaching agent added is 1.2-2.0 times the theoretical amount of total tin in the pretreated tin-containing slag, in terms of molar amount; the leaching temperature is 30℃-70℃, and the leaching time is 1h-3h; controlling the pH value of the mixture to be stable at 3.5-7.0 includes: adding sulfuric acid with a concentration of 10wt%-15wt% or ammonia water with a concentration of 5wt%-10wt% to the mixture to control the pH value of the mixture to be stable at 3.5-7.0; A solution containing a sulfiding agent is added dropwise to the tin-containing leaching solution to obtain a reaction solution, followed by precipitation to obtain a purified solution and a purified residue. The sulfiding agent is selected from NaHS and Na2S. The mass concentration of the sulfiding agent in the solution is 0.5wt%-2wt%. The precipitation temperature is 40℃-50℃, and the precipitation time is 1h-2h. The amount of sulfiding agent added is 1.2-1.5 times the copper content in the tin-containing leaching solution. During the precipitation process, sulfuric acid with a concentration of 10wt%-15wt% or ammonia water with a concentration of 5wt%-10wt% is added to the reaction solution to control the pH value of the reaction solution to 3-5. A 10wt%-15wt% sulfuric acid solution is added to the purification solution to obtain an acidified purification solution with a pH of 1-2. The reaction proceeds to obtain stannic acid precipitate and a mother liquor containing a complexing agent. The reaction temperature is 50℃-80℃, and the reaction time is 40min-60min. A 5wt%-10wt% ammonia solution is added to the mother liquor containing the complexing agent to adjust its pH to 3.5-7.0, thereby regenerating the complexing agent. The stannic acid precipitate is washed with water until the solution pH is 7-8, and then dried and calcined to obtain tin dioxide. The drying temperature is 80℃-100℃ and the drying time is 6h-10h; the calcination temperature is 600℃-700℃ and the calcination time is 2h-3h.

2. The method according to claim 1, characterized in that, The temperature of the drying process I is 80℃-100℃, and the drying time is 2h-5h.

3. The method according to claim 1, characterized in that, The leaching is carried out under stirring conditions, and the stirring speed is 200 r / min-350 r / min.

4. The method according to claim 1, characterized in that, The precipitation is carried out under stirring conditions, and the stirring speed is 200 r / min-250 r / min.

5. The method according to claim 1, characterized in that, The reaction is carried out under stirring conditions at a speed of 100 r / min to 200 r / min.

Images