A method for promoting zinc and arsenic removal from electric furnace flue dust and improving indium leaching rate

Abstract

The application provides a method for promoting zinc and arsenic removal of electric furnace flue dust and improving indium leaching rate, comprising the following steps: mixing electric furnace flue dust and dilute sulfuric acid solution, heating and stirring in a water bath kettle to carry out low-acid leaching, filtering after a period of reaction to obtain low-acid leaching solution and low-acid leaching residue; mixing the low-acid leaching residue with concentrated sulfuric acid to carry out sulfation roasting to obtain roasting residue; dissolving the roasting residue with a sulfuric acid solution, adding sodium chloride, heating and stirring in a water bath to carry out chlorine addition leaching to obtain leaching residue and leaching solution. The method uses electric furnace flue dust rich in tin, lead, zinc and arsenic as raw material, removes zinc and arsenic through one-stage low-acid leaching, and realizes efficient recovery of indium through sulfation roasting and chlorine addition acid leaching of the leaching residue. The method can significantly improve the indium recovery rate of complex electric furnace flue dust rich in tin, lead, zinc and arsenic, and realizes efficient recovery and reuse of low-grade indium.

Description

Technical Field

[0001] This invention belongs to the field of indium recycling technology, specifically relating to a technical method for improving the indium leaching rate in low-grade indium-containing electric furnace dust, and particularly to a method for promoting the removal of zinc and arsenic from electric furnace dust and improving the indium leaching rate. Background Technology

[0002] Indium, as an important rare and dispersed metal, has a wide range of applications in industry due to its unique physical and chemical properties. Among them, 70% of indium is used in the optoelectronic field to prepare ITO thin film materials due to its excellent light transmittance and conductivity. In addition, indium also plays an important role in electronic semiconductors, solders and alloys, optical fiber communications, battery corrosion protection, and modern military applications.

[0003] Currently, domestic indium production mainly uses byproducts such as leaching residue, dust, and slag generated during the smelting of non-ferrous metals like zinc, lead, tin, copper, and antimony as raw materials. Indium is enriched and recovered through a process involving concentration, leaching, extraction and back-extraction, displacement, and refining. Among these processes, extraction and back-extraction, displacement, and refining are relatively mature, while leaching has become the core factor limiting the improvement of indium recovery rates. Common indium leaching processes at present include conventional acid leaching, oxidative leaching, oxygen pressure leaching, and concentrated sulfuric acid roasting-water leaching.

[0004] The process logic of the concentrated sulfuric acid roasting-water leaching method is as follows: the raw material is mixed with sulfuric acid and roasted at a specific temperature, followed by water leaching of the roasted product. The advantage of this process is that the sulfation roasting process converts the metal in the raw material into easily leached sulfate forms; simultaneously, for indium that is partially encapsulated by quartz and sulfides and difficult to leach effectively, concentrated sulfuric acid roasting can disrupt the structure of the encapsulating minerals, thereby enhancing the reactivity of the encapsulated indium. However, when processing flue dust with high tin content, the limitations of the concentrated sulfuric acid roasting-water leaching method become apparent, with the indium leaching rate only reaching about 50%. Related studies have shown that as the tin content in the flue dust increases, the proportion of non-stoichiometric solid solutions InO·xSnO2 and In2O·xSnO2 in the flue dust increases significantly; because these solid solutions are chemically stable, they are difficult to react with the leaching system, ultimately leading to a low indium leaching rate.

[0005] To address this issue, researchers have proposed a sodium chlorate-enhanced chlorination leaching process: utilizing the oxidizing properties of sodium chlorate, indium in the chemically stable InO·xSnO2 and In2O·xSnO2 solid solutions is converted into a higher valence state, facilitating subsequent leaching. Simultaneously, the chloride ions generated by the sodium chlorate reaction can further achieve chlorination leaching, a technique commonly used for treating sparingly soluble minerals and offering the advantage of high metal leaching rates. However, existing technologies still have significant limitations: they cannot simultaneously enhance the reactivity of indium encapsulated by minerals such as quartz, as well as indium within the InO·xSnO2 and In2O·xSnO2 solid solutions, thus failing to significantly improve the overall indium leaching rate.

[0006] Therefore, there is an urgent need to develop a leaching process that can take into account both forms of indium in order to break through the existing technical bottlenecks and achieve efficient recovery of indium from tin-rich electric furnace flue dust. Summary of the Invention

[0007] The technical problem to be solved by this invention is to overcome the shortcomings of the existing technology. Combining the advantages of sulfation roasting and chlorination leaching, this invention proposes a method to promote the removal of zinc and arsenic from electric furnace flue dust and improve the indium leaching rate. It aims to provide a method to improve the efficiency of removing zinc and arsenic from lead-tin smelting electric furnace flue dust and recovering low-grade indium. Through this method, the removal of zinc and arsenic from electric furnace flue dust and the effective recovery of low-grade indium can be achieved, thereby maximizing the utilization of resources.

[0008] To address the aforementioned technical problems, this invention provides a method for promoting zinc and arsenic removal from electric furnace flue dust and increasing indium leaching rate, the method comprising the following steps: S1. Mix electric furnace flue gas and dilute sulfuric acid solution, heat and stir in a water bath to carry out low-acid leaching, filter after reacting for a period of time to obtain low-acid leaching solution and low-acid leaching residue. S2. The low-acid leaching residue is mixed with concentrated sulfuric acid and then subjected to sulfation roasting to obtain roasted residue. S3. Dissolve the roasted residue in sulfuric acid solution, then add sodium chloride and heat and stir in a water bath to carry out chlorination leaching, to obtain leaching residue and leaching solution.

[0009] Furthermore, in the above method, S1 further includes: sending the low-acid leaching solution to the acid leaching and zinc removal process of the flue gas furnace to achieve zinc enrichment and recovery and arsenic removal.

[0010] In the above method, the concentration of the dilute sulfuric acid solution in S1 is 80 g / L to 150 g / L, the leaching liquid-to-solid ratio is 3.0 to 6.0:1, the leaching temperature is 50℃ to 90℃, and the leaching time is 1 h to 4 h.

[0011] In the above method, further, the mass ratio of concentrated sulfuric acid to low-acid leaching residue in S2 is 0.5 to 0.7:1.

[0012] In the above method, the sulfation roasting temperature in S2 is 200℃~300℃, and the roasting time is 1 h~2 h.

[0013] In the above method, further, in step S3, the liquid-to-solid ratio of sodium chloride leaching is 3–6:1, the concentration of sulfuric acid solution is 0 g / L–100 g / L, and Cl is introduced in the form of NaCl. - The concentration is 30 g / L.

[0014] Furthermore, in the above method, the chlorination leaching temperature in S3 is 50℃~90℃, and the leaching time is 2 h~8 h.

[0015] Further, in step S3, the leaching solution is enriched and recovered to obtain indium product, and the indium-free solution is returned to step S1 for low-acid leaching.

[0016] Further, in step S3, the leaching residue is smelted in an electric furnace to obtain a lead-tin alloy.

[0017] Compared with the prior art, the advantages of the present invention are as follows: (1) This invention provides a method for promoting the removal of zinc and arsenic from electric furnace dust and increasing the indium leaching rate. The method employs sulfation roasting and chloric acid leaching processes to improve the indium leaching rate from tin-rich electric furnace dust. The sulfation roasting process, with the participation of oxygen, can both destroy the mineral structure of the indium and promote the conversion of low-valence InO / In2O·xSnO2 solid solution into easily soluble high-valence indium. Simultaneously, chloric acid leaching further promotes the leaching of sparingly soluble indium. This method has advantages such as high indium leaching rate, good operational safety, and fewer interfering impurities in subsequent extraction. It significantly increases the leaching rate of low-grade indium in complex tin-rich lead-zinc-arsenic dust to approximately 80%, effectively promoting the recycling of low-grade indium.

[0018] (2) The present invention provides a method for promoting the removal of zinc and arsenic from electric furnace flue dust and improving the indium leaching rate. The indium ion concentration in the leaching solution is high, while the concentration of impurity ions such as iron and tin is low, which is conducive to the subsequent enrichment and recovery process.

[0019] (3) The present invention provides a method for promoting the removal of zinc and arsenic from electric furnace flue dust and improving the indium leaching rate. The method is simple to operate and has good safety. It does not require the use of high-risk pressurized acid leaching process, nor does it require the addition of strong oxidizing reagents. Attached Figure Description

[0020] To make the objectives, technical solutions, and advantages of the embodiments of the present invention clearer, the technical solutions of the embodiments of the present invention will be clearly and completely described below with reference to the accompanying drawings.

[0021] Figure 1 The above are process flow diagrams for Embodiments 1 to 3 of the present invention. Detailed Implementation

[0022] The present invention will be further described below with reference to specific preferred embodiments, but this does not limit the scope of protection of the present invention.

[0023] The materials, reagents, and instruments used in the following examples are all commercially available. Unless otherwise specified, the experimental methods used in the following examples are conventional methods in the art.

[0024] Example 1 A method for promoting zinc and arsenic removal from electric furnace flue dust and improving indium leaching rate, the process is as follows: Figure 1 This includes the following steps: (1) Low-acid leaching of electric furnace dust: 40g of electric furnace dust and dilute sulfuric acid solution were mixed with a liquid-solid ratio of 5:1 and an initial H2SO4 concentration of 150 g / L. The mixture was heated and stirred in a 90℃ water bath for 4.0 h for low-acid leaching. After the reaction was completed, 183 mL of low-acid leaching solution was obtained by filtration. The low-acid leaching residue was dried at 60℃ to obtain 25.78 g.

[0025] The zinc concentration in the low-acid leaching solution reached 44.7 g / L, with a zinc leaching rate of 72.30% and an arsenic leaching rate of 71.43%. Lead, tin, indium, and other elements were almost not leached. The low-acid leaching solution was then sent to the acid leaching and zinc removal process in the flue gas furnace to achieve zinc enrichment and recovery, as well as arsenic removal.

[0026] (2) Sulphation roasting of low acid leaching residue: A certain amount of concentrated sulfuric acid is mixed evenly with low acid leaching residue (the solid-liquid ratio of concentrated sulfuric acid to low acid leaching residue is 0.7:1), and then roasted in a muffle furnace at 300℃ for 2.0 h. After cooling, 33 g of roasted residue is obtained.

[0027] (3) Chloric acid leaching of roasted residue: The roasted residue was dissolved in 100 g / L H2SO4 at a liquid-to-solid ratio of 4:1 (liquid / low-acid leaching residue), and 30 g / L Cl was added in the form of NaCl. - The reaction was then carried out at 90℃ for 4.0 h, and the leachate and leachate residue were obtained by filtration after the reaction was completed.

[0028] The indium concentration in the leachate was measured to be 878.8 mg / L, with an indium leaching rate of 80.30%. The indium content in the slag decreased to 0.13%. The tin concentration in the leachate was 9272 mg / L, with a tin leaching rate of 24.85%.

[0029] The leaching solution is enriched and recovered to obtain indium products. After indium removal, the solution is returned to step S1 for low-acid leaching. The leaching residue is smelted in an electric furnace to obtain a lead-tin alloy.

[0030] Example 2 A method for promoting zinc and arsenic removal from electric furnace flue dust and improving indium leaching rate, the process is as follows: Figure 1 This includes the following steps: (1) Low-acid leaching of electric furnace dust: 40g of electric furnace dust and dilute sulfuric acid solution were mixed with a liquid-solid ratio of 5:1 and an initial H2SO4 concentration of 150 g / L. The mixture was heated and stirred in a 90℃ water bath for 4.0 h for low-acid leaching. After the reaction was completed, 183 mL of low-acid leaching solution was obtained by filtration. The low-acid leaching residue was dried at 60℃ to obtain 25.78 g.

[0031] The zinc concentration in the low-acid leaching solution reached 44.7 g / L, with a zinc leaching rate of 72.30% and an arsenic leaching rate of 71.43%. Lead, tin, indium, and other elements were almost not leached. The low-acid leaching solution was then sent to the acid leaching and zinc removal process in the flue gas furnace to achieve zinc enrichment and recovery, as well as arsenic removal.

[0032] (2) Sulphation roasting of low acid leaching residue: A certain amount of concentrated sulfuric acid is mixed evenly with low acid leaching residue (the solid-liquid ratio of concentrated sulfuric acid to low acid leaching residue is 0.7:1), and then roasted in a muffle furnace at 300℃ for 2.0 h. After cooling, 33 g of roasted residue is obtained.

[0033] (3) Chloric acid leaching of roasted residue: The roasted residue was dissolved in H2SO4 with a concentration of 35 g / L at a liquid-to-solid ratio of 4:1 (liquid / low-acid leaching residue), and 30 g / L Cl was added in the form of NaCl. - The reaction was then carried out at 90℃ for 4.0 h, and the leachate and leachate residue were obtained by filtration after the reaction was completed.

[0034] The indium concentration in the leachate was measured to be 805.2 mg / L, with an indium leaching rate of 82.90%. The indium content in the slag decreased to 0.122%. The tin concentration in the leachate was 2170.8 mg / L, with a tin leaching rate of 7.78%.

[0035] The leaching solution is enriched and recovered to obtain indium products. After indium removal, the solution is returned to step S1 for low-acid leaching. The leaching residue is smelted in an electric furnace to obtain a lead-tin alloy.

[0036] Example 3 A method for promoting zinc and arsenic removal from electric furnace flue dust and improving indium leaching rate, the process is as follows: Figure 1 This includes the following steps: (1) Low-acid leaching of electric furnace dust: 40g of electric furnace dust and dilute sulfuric acid solution were mixed with a liquid-solid ratio of 5:1 and an initial H2SO4 concentration of 150 g / L. The mixture was heated and stirred in a 90℃ water bath for 4.0 h for low-acid leaching. After the reaction was completed, 183 mL of low-acid leaching solution was obtained by filtration. The low-acid leaching residue was dried at 60℃ to obtain 25.78 g.

[0037] The zinc concentration in the low-acid leaching solution reached 44.7 g / L, with a zinc leaching rate of 72.30% and an arsenic leaching rate of 71.43%. Lead, tin, indium, and other elements were almost not leached. The low-acid leaching solution was then sent to the acid leaching and zinc removal process in the flue gas furnace to achieve zinc enrichment and recovery, as well as arsenic removal.

[0038] (2) Sulphation roasting of low acid leaching residue: A certain amount of concentrated sulfuric acid is mixed evenly with low acid leaching residue (the solid-liquid ratio of concentrated sulfuric acid to low acid leaching residue is 0.7:1), and then roasted in a muffle furnace at 300℃ for 2.0 h. After cooling, 33 g of roasted residue is obtained.

[0039] (3) Chloric acid leaching of roasted residue: Dissolve the roasted residue in water at a liquid-to-solid ratio of 4:1 (liquid / low-acid leaching residue), and add 30 g / L Cl in the form of NaCl. - The reaction was then carried out at 90℃ for 4.0 h, and the leachate and leachate residue were obtained by filtration after the reaction was completed.

[0040] The indium concentration in the leachate was measured to be 834 mg / L, with an indium leaching rate of 75.39%. The indium content in the slag decreased to 0.159%. The tin concentration in the leachate was 101 mg / L, with a tin leaching rate of 0.3%.

[0041] The leaching solution is enriched and recovered to obtain indium products. After indium removal, the solution is returned to step S1 for low-acid leaching. The leaching residue is smelted in an electric furnace to obtain a lead-tin alloy.

[0042] Based on the above examples 1 to 3, it can be analyzed that as the concentration of H2SO4 in step (3) decreases, the tin leaching rate decreases, proving that appropriately reducing the sulfuric acid concentration can simultaneously achieve a high indium leaching rate and a low tin leaching rate, thereby reducing the interference of tin impurities on the subsequent enrichment and recovery of indium.

[0043] Example 4 A method for promoting zinc and arsenic removal from electric furnace flue dust and improving indium leaching rate includes the following steps: (1) Low-acid leaching of electric furnace dust: 40g of electric furnace dust and dilute sulfuric acid solution were mixed with a liquid-solid ratio of 5:1 and an initial H2SO4 concentration of 150 g / L. The mixture was heated and stirred in a 90℃ water bath for 4.0 h for low-acid leaching. After the reaction was completed, 183 mL of low-acid leaching solution was obtained by filtration. The low-acid leaching residue was dried at 60℃ to obtain 25.78 g.

[0044] The zinc concentration in the low-acid leaching solution reached 44.7 g / L, with a zinc leaching rate of 72.30% and an arsenic leaching rate of 71.43%. Lead, tin, indium, and other elements were almost not leached. The low-acid leaching solution was then sent to the acid leaching and zinc removal process in the flue gas furnace to achieve zinc enrichment and recovery, as well as arsenic removal.

[0045] (2) Direct chlorination of low-acid leaching residue: Dissolve the low-acid leaching residue in 100 g / L H2SO4 at a liquid-to-solid ratio of 4:1 (liquid / low-acid leaching residue), and add 30 g / L Cl as NaCl. - The reaction was then carried out at 90℃ for 4.0 h, and the leachate and leachate residue were obtained by filtration after the reaction was completed.

[0046] The tin concentration in the leachate was measured to be 5584 mg / L, with a tin leaching rate of 17.82%, while the indium concentration reached 570.6 mg / L. The indium content in the slag decreased to 0.22%, and the indium leaching rate was 62.41%.

[0047] The leaching solution is enriched and recovered to obtain indium products. After indium removal, the solution is returned to step S1 for low-acid leaching. The leaching residue is smelted in an electric furnace to obtain a lead-tin alloy.

[0048] Comparing Example 1 and Example 4, it can be seen that the indium leaching rate of Example 4 is significantly lower than that of Example 1 (80.30%) after sulfuric acid roasting and chloric acid leaching.

[0049] Example 5 A method for promoting zinc and arsenic removal from electric furnace flue dust and improving indium leaching rate includes the following steps: (1) Low-acid leaching of electric furnace dust: 40g of electric furnace dust and dilute sulfuric acid solution were mixed with a liquid-solid ratio of 5:1 and an initial H2SO4 concentration of 150 g / L. The mixture was heated and stirred in a 90℃ water bath for 4.0 h for low-acid leaching. After the reaction was completed, 183 mL of low-acid leaching solution was obtained by filtration. The low-acid leaching residue was dried at 60℃ to obtain 25.78 g.

[0050] The zinc concentration in the low-acid leaching solution reached 44.7 g / L, with a zinc leaching rate of 72.30% and an arsenic leaching rate of 71.43%. Lead, tin, indium, and other elements were almost not leached. The low-acid leaching solution was then sent to the acid leaching and zinc removal process in the flue gas furnace to achieve zinc enrichment and recovery, as well as arsenic removal.

[0051] (2) Sulphation roasting of low acid leaching residue: A certain amount of concentrated sulfuric acid is mixed evenly with low acid leaching residue (the solid-liquid ratio of concentrated sulfuric acid to low acid leaching residue is 0.7:1), and then roasted in a muffle furnace at 300℃ for 2.0 h. After cooling, 33 g of roasted residue is obtained.

[0052] (3) Acid leaching of roasted residue: The roasted residue was dissolved in 100 g / L H2SO4 at a liquid-solid ratio of 4:1 (liquid / low acid leaching residue), and then reacted at a reaction temperature of 90℃ for 4.0 h. After the reaction was completed, the leachate and leaching residue were obtained by filtration.

[0053] The tin concentration in the leachate was measured to be 40.7 mg / L, with a tin leaching rate of 0.14%, while the indium concentration in the leachate reached 465.6 mg / L, the indium content in the slag was 0.36%, and the indium leaching rate was 40.98%.

[0054] The leaching solution is enriched and recovered to obtain indium products. After indium removal, the solution is returned to step S1 for low-acid leaching. The leaching residue is smelted in an electric furnace to obtain a lead-tin alloy.

[0055] Comparing Example 1 and Example 5, Example 5 showed a significantly lower indium leaching rate than Example 1 (80.30%) achieved by sulfuric acid roasting followed by chloric acid leaching.

[0056] The above description is merely a preferred embodiment of the present invention and is not intended to limit the present invention in any way. Although the present invention has been disclosed above with reference to preferred embodiments, it is not intended to limit the present invention. Any person skilled in the art can make many possible variations and modifications to the technical solutions of the present invention using the methods and techniques disclosed above, or modify them into equivalent embodiments with equivalent changes, without departing from the spirit and technical essence of the present invention. Therefore, any simple modifications, equivalent substitutions, equivalent changes, and modifications made to the above embodiments based on the technical essence of the present invention without departing from the content of the technical solutions of the present invention shall still fall within the protection scope of the technical solutions of the present invention.

Claims

1. A method for promoting the removal of zinc and arsenic from electric furnace flue dust and improving the indium leaching rate, characterized in that, The method includes the following steps: S1. Mix electric furnace flue gas and dilute sulfuric acid solution, heat and stir in a water bath to carry out low acid leaching, filter after reacting for a period of time to obtain low acid leaching solution and low acid leaching residue. S2. The low-acid leaching residue is mixed with concentrated sulfuric acid and then subjected to sulfation roasting to obtain roasted residue; S3. Dissolve the roasted residue in sulfuric acid solution, then add sodium chloride and heat and stir in a water bath to perform chlorination leaching, to obtain leaching residue and leaching solution.

2. The method according to claim 1, characterized in that, The S1 further includes: sending the low-acid leaching solution to the acid leaching and zinc removal process of the flue gas furnace to achieve zinc enrichment and recovery and arsenic removal.

3. The method according to claim 1, characterized in that, The concentration of the dilute sulfuric acid solution in S1 is 80 g / L to 150 g / L, the leaching liquid-to-solid ratio is 3.0 to 6.0:1, the leaching temperature is 50℃ to 90℃, and the leaching time is 1 h to 4 h.

4. The method according to claim 1, characterized in that, The mass ratio of concentrated sulfuric acid to low-acid leaching residue in S2 is 0.5 to 0.7:

1.

5. The method according to claim 1, characterized in that, The sulfation roasting in S2 is carried out at a temperature of 200℃ to 300℃ for 1 h to 2 h.

6. The method according to claim 1, characterized in that, In step S3, the liquid-to-solid ratio of sodium chloride leaching is 3–6:1, the concentration of sulfuric acid solution is 0 g / L–100 g / L, and Cl is introduced in the form of sodium chloride. - The concentration is 30 g / L.

7. The method according to claim 1, characterized in that, The chlorination leaching in S3 is carried out at a temperature of 50℃ to 90℃ for a time of 2 h to 8 h.

8. The method according to any one of claims 1 to 7, characterized in that, In step S3, the leaching solution is enriched and recovered to obtain indium products. After indium removal, the solution is returned to step S1 for low-acid leaching.

9. The method according to any one of claims 1 to 7, characterized in that, In step S3, the leaching residue is smelted in an electric furnace to obtain a lead-tin alloy.

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