A method for selective leaching and recycling treatment of multi-metal hazardous waste

Abstract

The application discloses a kind of multi-metal hazardous waste selective leaching recovery processing methods, belong to hazardous waste resource processing technical field.The method is mixed after multi-metal hazardous waste, sodium carbonate and sulfur powder is carried out mechanical force chemistry ball milling, and sulfurized activated powder is obtained;Sulfurized activated powder is mixed with water glass and pore-forming agent and is formed, and porous sulfurized calcined body is obtained in microwave field;Porous sulfurized calcined body is placed in the middle chamber of diaphragm electrolytic cell, and anode chamber is input into chlorine, so that copper and cadmium enter solution in the form of chlorine complex, and copper and cadmium containing leaching solution and decopper cadmium calcined body are obtained;Ammonium sulfate and ammonia are added to decopper cadmium calcined body and are subjected to reducing complex leaching, so that zinc and nickel enter solution in the form of ammonia complex, and zinc and nickel containing leaching solution and lead residue are obtained;Copper and cadmium containing leaching solution is subjected to electrowinning decoppering and cadmium is recovered by sodium carbonate precipitation, and zinc and nickel containing leaching solution is subjected to solvent extraction to recover zinc and nickel respectively, and lead residue is transferred into pyrometallurgical smelting process.

Description

Technical Field

[0001] This invention relates to the field of hazardous waste resource utilization technology, specifically a selective leaching and recycling method for multi-metal hazardous waste. Background Technology

[0002] Heavy metals such as copper, cadmium, zinc, nickel, and lead in polymetallic hazardous waste exist in complex forms and have high levels of toxic components. Conventional treatment methods include pyrometallurgical smelting, wet acid leaching, ammonia leaching, and chlorination leaching. Pyrometallurgical smelting has high energy consumption, and the metals in the generated fumes and slag are dispersed, making secondary pollution control difficult. Wet acid leaching can dissolve most heavy metals in acidic media, but copper, cadmium, zinc, and nickel often enter the leachate simultaneously, resulting in long subsequent extraction, precipitation, or electrowinning processes and high reagent consumption. The stability of lead in the acid leaching residue is significantly affected by leaching conditions. Ammonia leaching is selective for zinc and nickel, but copper and cadmium may also partially dissolve in ammoniacal systems, leading to a mixed composition of the leachate. Chlorination leaching can utilize chloride ions to form chloride complexes with copper and cadmium, but conventional chlorination leaching lacks sufficient control over the oxidation and release process of sulfide materials, and zinc and nickel are easily co-soluble. Current processes typically involve first sulfiding and roasting the material to convert the metal into sulfides before stepwise leaching. However, conventional roasting results in a dense, low-porosity roasted body, making it difficult for the leaching agent to penetrate the particles. The limited metal sulfide reaction interface also leads to slow leaching rates. While mechanochemical ball milling can lower the sulfidation reaction temperature, subsequent conventional heating can cause powder agglomeration and sintering, failing to preserve the activated structure. Microwave heating offers internal heating, but its improvement on leaching performance is limited if the material does not form suitable sulfidation precursors or introduce a pore-forming structure. Existing technologies lack effective control over redox conditions in the selective separation of copper-cadmium (Cd) and zinc-nickel-lead (ZN-L) processes, making it difficult to remove Cd and ZN in the same process before leaching ZN and ZN while simultaneously stabilizing lead in the slag. There is a need in the treatment of polymetallic hazardous waste to obtain highly active porous sulfided roasted bodies at lower temperatures and within shorter processes, as well as to achieve stepwise selective leaching of Cd and ZN. Summary of the Invention

[0003] This invention provides a selective leaching and recycling method for multi-metal hazardous waste, which utilizes mechanochemical ball milling sulfidation, microwave field roasting to create pores, chlorination leaching in the intermediate chamber of a diaphragm electrolytic cell, and ammonia-based reduction complexation leaching to achieve stepwise selective recovery of copper, cadmium, zinc, nickel, and lead.

[0004] To achieve the above objectives, the present invention provides the following technical solution: The present invention provides a selective leaching and recycling method for polymetallic hazardous waste, comprising the following steps: mixing polymetallic hazardous waste, sodium carbonate, and sulfur powder, and then subjecting the mixture to mechanochemical ball milling to obtain sulfidated activated powder; mixing the sulfidated activated powder with water glass and a pore-forming agent to form a precipitated body, and then calcining it in a microwave field to obtain a porous sulfidated calcined body; placing the porous sulfidated calcined body in the intermediate chamber of an electrically powered diaphragm electrolytic cell, adding sodium chloride solution to the anode chamber and sodium hydroxide solution to the cathode chamber, and introducing chlorine gas into the anode chamber. Copper and cadmium in the intermediate chamber slurry are introduced into the solution as chloride complexes to obtain a copper-cadmium leaching solution and a copper-cadmium-removed roasted body. Ammonium sulfate and ammonia are added to the copper-cadmium-removed roasted body for reducing complexation leaching, allowing zinc and nickel to enter the solution as ammonia complexes to obtain a zinc-nickel leaching solution and lead slag. The copper-cadmium leaching solution is subjected to electrowinning to remove copper, yielding cathode copper and copper removal solution. Sodium carbonate is added to the copper removal solution to precipitate and recover cadmium. The zinc-nickel leaching solution is subjected to solvent extraction to obtain a zinc-containing back-extraction solution and a nickel-containing back-extraction solution, respectively. The lead slag is used as a lead recovery raw material and transferred to the pyrometallurgical smelting process.

[0005] Preferably, the multi-metallic hazardous waste is dried and crushed into powder with a particle size of less than 0.5 mm. Sodium carbonate and sulfur powder are mixed at a mass ratio of 3:1 to 5:1 to prepare a sulfiding agent. The powder and sulfiding agent are fed into a planetary ball mill, and zirconia balls and anhydrous ethanol are added. The milling is carried out under an argon atmosphere at a speed of 300 to 500 rpm for 2 to 4 hours. After the milling is completed, the anhydrous ethanol is recovered by vacuum distillation to obtain the sulfidated activated powder. Mechanochemical activation allows the copper, cadmium, zinc, nickel and other metal components in the hazardous waste to fully contact with sulfur, thereby improving the efficiency of subsequent sulfidation conversion.

[0006] Preferably, the sulfur activation powder, water glass, pore-forming agent, and binder are mixed in a mass ratio of 100:8:15:5, and an appropriate amount of water is added and stirred evenly. The mixture is then pressed into a cylindrical blank. The cylindrical blank is dried at 60°C to 80°C for 6 to 8 hours, placed in a microwave calcining furnace, and calcined in a nitrogen atmosphere using a segmented heating method. The first segment temperature is 200°C to 300°C, and the second segment temperature is 450°C to 550°C. After calcination, the blank is cooled to room temperature with the furnace to obtain a porous sulfur-cured calcined body. The porous structure facilitates the subsequent penetration and mass transfer of the leachate inside the calcined body, and microwave calcination can improve the uniformity and porosity of the sulfur-cured calcined body.

[0007] Preferably, a porous baffle is installed at the bottom of the intermediate chamber of the diaphragm electrolytic cell, and the porous sulfurized calcined body is placed on the porous baffle. An anion exchange membrane is installed between the anode chamber and the intermediate chamber, and a cation exchange membrane is installed between the cathode chamber and the intermediate chamber. Sodium chloride solution is added to the anode chamber, and sodium hydroxide solution is added to the cathode chamber. Chlorine gas is introduced into the anode chamber, and the chlorine water generated in the anode chamber permeates into the intermediate chamber through the anion exchange membrane. Sodium chloride and hydrochloric acid are added to the intermediate chamber to adjust the chloride ion concentration in the slurry of the intermediate chamber. Under stirring conditions, copper and cadmium enter the solution in the form of chloride complexes. The slurry of the intermediate chamber is filtered to obtain a copper-cadmium leaching solution and a copper-cadmium-removed calcined body. Through membrane separation and chlorination, the selective leaching rate of copper and cadmium is improved, while most of zinc, nickel, and lead are retained in the calcined body, reducing the burden of subsequent leaching.

[0008] In a further preferred embodiment, chlorine gas is pumped into a storage tank via a compressor and then introduced into the anode chamber through a microporous aeration pipe from the bottom of the storage tank. The sodium chloride solution in the anode chamber absorbs the chlorine gas to generate chlorine water. The chlorine gas flow rate is controlled to maintain the redox potential of the slurry in the intermediate chamber between 0.2 V and 0.6 V. Chlorine molecules and hypochlorous acid in the chlorine water permeate through the anion exchange membrane into the intermediate chamber. The sulfides in the intermediate chamber are oxidized, releasing copper and cadmium ions. These copper and cadmium ions combine with chloride ions to form chloride complexes that enter the solution. Controlling the redox potential prevents the excessive oxidation and dissolution of zinc and nickel.

[0009] Preferably, the reducing complex leaching process is as follows: The copper-cadmium-free calcined body is added to a reducing complex leaching tank, ammonium sulfate is added and stirred to dissolve, and ammonia is added to adjust the slurry pH to 8-10. Sodium sulfite and iron powder are added in three batches while stirring to maintain a reducing atmosphere in the slurry, allowing zinc and nickel to react with ammonia to form ammonia complexes that enter the solution, while lead remains in the slag as solid lead sulfate. After leaching, solid-liquid separation is performed to obtain a zinc-nickel leaching solution and lead slag. The reducing atmosphere inhibits lead sulfate dissolution, ensuring that lead remains stably in the slag.

[0010] Further preferred embodiment involves the following steps: First, sodium sulfite and iron powder are added at 2% and 1% of the mass of the copper-cadmium-removed roasted body, respectively. After stirring for 20 minutes, the dissolved oxygen concentration in the slurry is measured. Second, sodium sulfite and iron powder are added at 1% and 0.5% of the mass of the copper-cadmium-removed roasted body, respectively, and stirring for 30 minutes. Third, sodium sulfite and iron powder are added at 0.5% and 0.3% of the mass of the copper-cadmium-removed roasted body, respectively, and stirring for 1 hour, reducing the dissolved oxygen concentration in the slurry to below 0.1 mg / L. Maintaining a continuous reducing environment by adding the reducing agent in stages is beneficial for the complexation leaching of zinc and nickel.

[0011] Preferably, the electrowinning process for removing copper from copper-cadmium leaching solution is as follows: using a titanium plate as the cathode and a plate-shaped lead-silver alloy as the anode, the copper-cadmium leaching solution is injected into an electrowinning tank for electrowinning, causing copper to be deposited at the cathode, yielding cathode copper and copper-removed solution; the copper-removed solution is transferred to a precipitation reactor, heated to 50-70 degrees Celsius, and sodium carbonate solution is slowly added under stirring, causing cadmium to precipitate as cadmium carbonate; the cadmium carbonate precipitate is separated by filtration, and the filtrate is returned to the intermediate chamber of the diaphragm electrolytic cell for reuse as leaching solution. This step achieves the sequential recovery of copper and cadmium, while the reuse of filtrate reduces wastewater discharge.

[0012] In a further preferred embodiment, the pH of the copper removal solution is adjusted to 9 to 11 using sodium hydroxide solution, and sodium carbonate solution is added at a rate controlled at 5 to 10 ml per minute. During the addition of sodium carbonate solution, carbon dioxide gas is simultaneously introduced into the bottom of the precipitation reactor, causing the carbon dioxide to react with cadmium ions in the copper removal solution to form cadmium carbonate precipitate. After precipitation, the mixture is allowed to stand for 2 hours, and the cadmium carbonate precipitate is washed with water until the pH of the washing solution is less than 8. The solution is then filtered and dried at 105 degrees Celsius to obtain the cadmium carbonate product. This treatment method can improve the completeness of cadmium precipitation and the purity of cadmium carbonate.

[0013] Preferably, the solvent extraction process for the zinc-nickel leaching solution is as follows: the zinc-nickel leaching solution is mixed with P204 extractant at a 1:1 ratio to extract zinc, yielding a zinc-containing organic phase and a nickel-containing raffinate; the zinc-containing organic phase is back-extracted with sulfuric acid to obtain a zinc-containing back-extract; the nickel-containing raffinate is mixed with LIX84 extractant to extract nickel, yielding a nickel-containing organic phase; the nickel-containing organic phase is back-extracted with hydrochloric acid to obtain a nickel-containing back-extract. By preferentially extracting zinc with P204 and extracting nickel with LIX84, efficient separation of zinc and nickel is achieved.

[0014] Lead slag is used as a raw material for lead recovery and is then transferred to the pyrometallurgical smelting process, enabling the recovery and utilization of lead. This method selectively leaches and separates copper, cadmium, zinc, nickel, and lead in sequence according to their metallic properties, avoiding interference between complex components in multi-metal hazardous waste. The recovery route is clear, and the leaching selectivity is good, which is conducive to improving the recovery rate of each valuable metal and reducing the difficulty of subsequent smelting and processing.

[0015] The technical effects and advantages provided by the present invention in the above technical solution are as follows: Sodium carbonate and sulfur powder are used as sulfiding agents to carry out solid-phase reactions with multi-metal hazardous waste under mechanochemical ball milling conditions. Mechanical force induces particle refinement, lattice distortion, and surface activation, allowing metal components such as copper, cadmium, zinc, and nickel to combine with sulfur at relatively low temperatures to form sulfides, resulting in sulfidated activated powder. After mixing and molding the sulfidated activated powder with water glass and a pore-forming agent, it is calcined in a segmented manner in a microwave field. The selective heating of the material by microwaves further completes the sulfidation reaction. The pore-forming agent decomposes and volatilizes, forming interconnected pores within the calcined body, while water glass participates in forming a porous framework with a certain strength. The interconnected pores within the porous sulfidated calcined body allow the leaching agent to penetrate deep into the particles, increasing the contact area between the metal sulfides and the leaching agent, accelerating the formation rate of copper-cadmium chloride complexes and zinc-nickel amine complexes, reducing the loss of metal components due to high-temperature calcination, shortening the leaching time, and lowering the temperature requirements of the calcination process.

[0016] The porous sulfide-cured body is placed in the intermediate chamber of a diaphragm electrolytic cell. In the anode chamber, sodium chloride solution absorbs chlorine gas to form chlorine water. Chlorine molecules and hypochlorous acid enter the intermediate chamber through anion exchange membrane. By controlling the redox potential, copper and cadmium sulfides are preferentially oxidized. The released copper and cadmium ions form chloride complexes with chloride ions and enter the solution. Zinc and nickel do not enter the solution at this stage, resulting in a copper-cadmium leaching solution and a copper-cadmium-removed calcined body. The copper-cadmium-removed calcined body undergoes reducing complexation leaching in an ammonium sulfate and ammonia system. Sodium sulfite and iron powder maintain the reducing atmosphere. Zinc and nickel form ammonia complexes and enter the solution, while lead remains in the slag as solid lead sulfate. The sequential separation of copper and cadmium from zinc, nickel, and lead avoids the complexity of subsequent separation processes and high reagent consumption caused by simultaneous leaching of multiple metals. The lead slag can be directly transferred to the pyrometallurgical smelting process. Copper, cadmium, zinc, and nickel in the leaching solution can be subsequently recovered as cathode copper, cadmium carbonate, zinc-containing back-extraction solution, and nickel-containing back-extraction solution, respectively. Attached Figure Description

[0017] To more clearly illustrate the technical solutions in the embodiments of this application or the prior art, the drawings used in the embodiments will be briefly introduced below. Obviously, the drawings described below are only some embodiments recorded in this invention. For those skilled in the art, other drawings can be obtained based on these drawings.

[0018] Figure 1 This is a process flow diagram of the selective leaching and recycling method for polymetallic hazardous waste; Figure 2 This is a flowchart of the reducing complexation leaching process for copper-cadmium-free calcined bodies; Figure 3 This is a process flow chart for copper removal and cadmium precipitation using copper-cadmium leaching solutions. Figure 4 This is a particle size distribution diagram of sulfurized activated powder under ball milling processes at different rotation speeds; Figure 5This is a graph showing the change in copper leaching rate under different redox potentials. Detailed Implementation

[0019] 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. Obviously, the described embodiments are only some embodiments of the present invention, not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0020] See Figure 1 This invention provides a selective leaching and recycling method for polymetallic hazardous waste. The method involves mixing polymetallic hazardous waste, sodium carbonate, and sulfur powder, followed by mechanochemical ball milling to obtain sulfidated activated powder. The sulfidated activated powder is then mixed with water glass and a pore-forming agent, molded, and calcined in a microwave field to obtain a porous sulfidated calcined body. This porous sulfidated calcined body is placed in the intermediate chamber of an electrically powered diaphragm electrolytic cell. Sodium chloride solution is added to the anode chamber, and sodium hydroxide solution is added to the cathode chamber. Chlorine gas is introduced into the anode chamber, causing copper and cadmium in the slurry in the intermediate chamber to dissolve in chlorine. The zinc and nickel complexes are introduced into the solution to obtain a copper-cadmium leaching solution and a copper-cadmium-removed roasted body. Ammonium sulfate and ammonia are added to the copper-cadmium-removed roasted body for reducing complex leaching, allowing zinc and nickel to enter the solution as ammonia complexes, resulting in a zinc-nickel leaching solution and lead slag. The copper-cadmium leaching solution is subjected to electrowinning to remove copper, resulting in cathode copper and copper removal solution. Sodium carbonate is added to the copper removal solution to precipitate and recover cadmium. The zinc-nickel leaching solution is subjected to solvent extraction to obtain a zinc-containing back-extraction solution and a nickel-containing back-extraction solution, respectively. The lead slag is used as a lead recovery raw material and transferred to the pyrometallurgical smelting process.

[0021] In practice, the polymetallic hazardous waste is dried in a drying device. The dried waste is then fed into a crusher for further crushing. The crushed material is sieved to obtain powder with a particle size of less than 0.5 mm. A vulcanizing agent is prepared by mixing sodium carbonate and sulfur powder at a mass ratio of 3:1 to 5:1, ensuring the sodium carbonate and sulfur powder are thoroughly mixed.

[0022] The powder and vulcanizing agent are fed into a planetary ball mill, into which zirconia balls and anhydrous ethanol are added. The grinding chamber of the planetary ball mill is replaced with an argon atmosphere, and ball milling is performed at a speed of 300 to 500 rpm for 2 to 4 hours under argon atmosphere. The powder and vulcanizing agent undergo a mechanochemical activation reaction under the impact and friction of the zirconia balls. After ball milling, the ball milling slurry is transferred to a vacuum distillation unit to recover the anhydrous ethanol, yielding the vulcanized activated powder.

[0023] In some embodiments, the vulcanized activated powder, water glass, pore-forming agent, and binder are mixed in a mass ratio of 100:8:15:5. Water glass serves as a high-temperature binder, the pore-forming agent forms pores during calcination, and the binder improves the molding strength of the cylindrical preform. A suitable amount of water is added and stirred until homogeneous to obtain a molding slurry. The molding slurry is pressed into a cylindrical preform. The cylindrical preform is dried at 60°C to 80°C for 6 to 8 hours to evaporate the moisture. The dried cylindrical preform is placed in a microwave calcination furnace, and nitrogen gas is introduced into the furnace. Calcination is performed in a nitrogen atmosphere using a segmented heating method. The first segment temperature is 200°C to 300°C, and the second segment temperature is 450°C to 550°C. The pore-forming agent is removed at the first segment temperature, and the vulcanized activated powder is sintered and maintains its porous structure at the second segment temperature. After calcination, the material is cooled to room temperature in the furnace to obtain a porous sulfide calcined body.

[0024] See Figure 4 In the figure, the horizontal axis represents the particle size of the sulfurized activated powder (unit: micrometers), and the vertical axis represents the probability density. The figure shows the particle size distribution curves of the sulfurized activated powder under different ball milling speeds, corresponding to 300 rpm (solid blue line), 400 rpm (dashed orange line), and 500 rpm (dotted green line), respectively.

[0025] It can be seen that the powder produced by ball milling at 300 rpm has a relatively wide particle size distribution, mainly concentrated between 40 and 90 micrometers, with a peak probability density of approximately 0.05, indicating a coarser particle size characteristic. At 400 rpm, the peak particle size distribution occurs at approximately 40 micrometers, with a peak probability density of approximately 0.06, showing a more concentrated distribution and a smaller overall particle size compared to 300 rpm. At 500 rpm, the particle size further decreases, with the peak particle size concentrated in the 20 to 30 micrometer range, and the peak probability density reaching approximately 0.09, exhibiting a finer particle size distribution characteristic.

[0026] In practice, a porous baffle is installed at the bottom of the intermediate chamber of the diaphragm electrolyzer, and the porous sulfurized calcined body is placed on the porous baffle. An anion exchange membrane is installed between the anode chamber and the intermediate chamber, and a cation exchange membrane is installed between the cathode chamber and the intermediate chamber. Sodium chloride solution is added to the anode chamber, and sodium hydroxide solution is added to the cathode chamber.

[0027] Chlorine gas generated from an external chlorine source is pumped into a storage tank via a compressor, and then introduced into the anode chamber from the bottom of the storage tank through a microporous aeration pipe. The sodium chloride solution in the anode chamber absorbs the chlorine gas to produce chlorine water, which contains chlorine molecules and hypochlorous acid.

[0028] The chlorine gas flow rate is controlled to maintain the redox potential of the slurry in the intermediate chamber between 0.2 V and 0.6 V. Chlorine molecules and hypochlorous acid in the chlorine water pass through the anion exchange membrane into the intermediate chamber. Sulfides in the intermediate chamber are oxidized, releasing copper and cadmium ions. These copper and cadmium ions combine with chloride ions to form chloride complexes that enter the solution.

[0029] In some embodiments, sodium chloride and hydrochloric acid are added to the intermediate chamber to adjust the chloride ion concentration in the slurry, and copper and cadmium are introduced into the solution as chloride complexes under stirring conditions. The copper and cadmium chloride complexes in the intermediate chamber slurry dissolve in the liquid phase.

[0030] The slurry in the intermediate chamber is filtered to obtain a copper-cadmium-containing leachate and a copper-cadmium-removed calcined body. The copper-cadmium-containing leachate enters the subsequent electrowinning copper removal process, while the copper-cadmium-removed calcined body enters the reducing complexation leaching process.

[0031] See Figure 5 The curves in the figure illustrate the trend of copper leaching rate with leaching time under different redox potentials in the intermediate chamber of the diaphragm electrolyzer in the above embodiment. The horizontal axis represents leaching time in minutes, ranging from 0 to 120 minutes; the vertical axis represents copper leaching rate in percentage, ranging from 0% to 100%. The legend indicates three different conditions with redox potentials of 0.2 V (solid line), 0.4 V (dashed line), and 0.6 V (dotted line).

[0032] As shown in the figure, the copper leaching rate under all three potential conditions increases with leaching time and gradually stabilizes. Specifically, at a redox potential of 0.6 V, the copper leaching rate increases the fastest, reaching a stable value of over 85% after approximately 60 minutes, indicating high copper leaching efficiency. At a redox potential of 0.4 V, the copper leaching rate is second, approaching a stable value of 80% after approximately 80 minutes. At a redox potential of 0.2 V, the copper leaching rate is relatively low, reaching only about 60% after 120 minutes, and the rate of increase is slow.

[0033] In specific implementation, please refer to Figure 2 The copper-cadmium-free calcined body is added to a reducing complexing leaching tank. Ammonium sulfate is added to the reducing complexing leaching tank, and the stirring device is turned on to dissolve the ammonium sulfate in the slurry. Ammonia water is added to the reducing complexing leaching tank to adjust the pH of the slurry to 8 to 10.

[0034] Sodium sulfite and iron powder were added in three batches while stirring to maintain a reducing atmosphere in the slurry. The first addition of sodium sulfite was 2% of the mass of the copper-cadmium-removed roasted body, and the addition of iron powder was 1% of the mass of the copper-cadmium-removed roasted body. The dissolved oxygen concentration in the slurry was measured after stirring for 20 minutes. The second addition of sodium sulfite and iron powder was 1% of the mass of the copper-cadmium-removed roasted body, and the addition of iron powder was 0.5% of the mass of the copper-cadmium-removed roasted body. Stirring was continued for 30 minutes. The third addition of sodium sulfite and iron powder was 0.5% of the mass of the copper-cadmium-removed roasted body, and the addition of iron powder was 0.3% of the mass of the copper-cadmium-removed roasted body. Stirring was continued for 1 hour to reduce the dissolved oxygen concentration in the slurry to below 0.1 mg / L.

[0035] In some embodiments, sodium sulfite and iron powder are in full contact with the slurry in a reducing complexing leaching tank to maintain a reducing atmosphere in the slurry. Zinc and nickel react with ammonia to form ammonium complexes that enter the solution, while lead remains in the slag as solid lead sulfate.

[0036] After leaching, the slurry in the reducing complex leaching tank undergoes solid-liquid separation to obtain zinc-nickel leaching solution and lead slag. The zinc-nickel leaching solution enters the solvent extraction process, while the lead slag is transferred to the pyrometallurgical smelting process as a lead recovery raw material.

[0037] In specific implementation, please refer to Figure 3 A titanium plate is used as the cathode, and a plate-shaped lead-silver alloy is used as the anode. The titanium plate and the plate-shaped lead-silver alloy are installed in an electrodeposition tank. A copper-cadmium leaching solution is injected into the electrodeposition tank, and electroposition is performed by applying an electric current. During the electroposition process, copper ions in the copper-cadmium leaching solution gain electrons and are deposited on the titanium plate cathode to form cathode copper. Electroposition continues until the copper concentration in the copper-cadmium leaching solution decreases to a preset range, at which point the electroposition is stopped, the cathode copper is removed, and the copper removal solution is collected.

[0038] Transfer the copper removal solution to a precipitation reactor, and heat the precipitation reactor to 50°C to 70°C. While stirring, add sodium hydroxide solution to the precipitation reactor to adjust the pH of the copper removal solution to 9 to 11. After adjusting the pH, add sodium carbonate solution to the precipitation reactor while continuously stirring, controlling the addition rate at 5 to 10 ml per minute.

[0039] During the addition of sodium carbonate solution, carbon dioxide gas is simultaneously introduced into the bottom of the precipitation reactor. The carbon dioxide gas enters the copper removal solution in the form of bubbles, and reacts with cadmium ions in the solution to form cadmium carbonate precipitate. The simultaneous introduction of carbon dioxide gas and the addition of sodium carbonate solution allow the cadmium carbonate precipitate to gradually form.

[0040] In some embodiments, the carbon dioxide gas flow is stopped after the sodium carbonate solution is added. The slurry in the precipitation reactor is allowed to stand for 2 hours to allow the cadmium carbonate precipitate to settle completely. After standing, the supernatant from the precipitation reactor is drained, and the cadmium carbonate precipitate is washed with water until the pH of the washing solution is less than 8. The washed cadmium carbonate precipitate is filtered and dried at 105 degrees Celsius to obtain the cadmium carbonate product. The filtrate discharged from the precipitation reactor is returned to the intermediate chamber of the diaphragm electrolyzer for reuse as leachate.

[0041] In the specific implementation, the zinc-nickel leaching solution and P204 extractant are added to the extraction equipment at a 1:1 ratio. The stirring device is started to ensure thorough mixing of the zinc-nickel leaching solution and P204 extractant. During the mixing process, the P204 extractant reacts with the zinc ions in the zinc-nickel leaching solution, and the zinc ions enter the organic phase of the P204 extractant, while the nickel ions remain in the aqueous phase. Stirring is stopped, and the mixture is allowed to stand for phase separation, yielding a zinc-containing organic phase and a nickel-containing raffinate. The zinc-containing organic phase and the nickel-containing raffinate are then separated.

[0042] In some embodiments, a zinc-containing organic phase is added to a back-extraction device, and sulfuric acid is added to the back-extraction device. The stirring device is started to ensure sufficient contact between the zinc-containing organic phase and the sulfuric acid. The sulfuric acid reacts with the zinc ions in the zinc-containing organic phase in a back-extraction reaction, and the zinc ions move from the zinc-containing organic phase into the aqueous sulfuric acid phase. Stirring is stopped, and the mixture is allowed to stand for phase separation to obtain a zinc-containing back-extraction solution.

[0043] Optionally, the nickel-containing raffinate is added to an extraction device, and LIX84 extractant is added to the extraction device. The stirring device is started to ensure thorough mixing of the nickel-containing raffinate and the LIX84 extractant. During mixing, the LIX84 extractant reacts with the nickel ions in the nickel-containing raffinate, and the nickel ions enter the organic phase of the LIX84 extractant. Stirring is stopped, and the mixture is allowed to stand for phase separation, yielding a nickel-containing organic phase and a raffinate aqueous phase. The nickel-containing organic phase and the raffinate aqueous phase are then separated.

[0044] In some embodiments, a nickel-containing organic phase is added to a back-extraction device, and hydrochloric acid is added to the back-extraction device. The stirring device is started to ensure sufficient contact between the nickel-containing organic phase and the hydrochloric acid. The hydrochloric acid reacts with the nickel ions in the nickel-containing organic phase, and the nickel ions move from the nickel-containing organic phase into the hydrochloric acid aqueous phase. Stirring is stopped, and the mixture is allowed to stand for phase separation to obtain a nickel-containing back-extraction solution.

[0045] The above description is merely a specific embodiment of this application, but the scope of protection of this application is not limited thereto. Any changes or substitutions that can be easily conceived by those skilled in the art within the scope of the technology disclosed in this application should be included within the scope of protection of this application.

Claims

1. A method for selective leaching and recycling of polymetallic hazardous waste, characterized in that, The method includes: Multimetallic hazardous waste, sodium carbonate and sulfur powder are mixed and then subjected to mechanochemical ball milling to obtain sulfurized activated powder. The sulfurized activated powder was mixed with water glass and a pore-forming agent and shaped into a porous sulfurized calcined body by calcination in a microwave field. The porous sulfurized calcined body is placed in the intermediate chamber of an electrically powered diaphragm electrolytic cell. Sodium chloride solution is added to the anode chamber and sodium hydroxide solution is added to the cathode chamber. Chlorine gas is introduced into the anode chamber so that copper and cadmium in the slurry in the intermediate chamber enter the solution in the form of chlorine complexes, thus obtaining a copper-cadmium leaching solution and a copper-cadmium-free calcined body. Ammonium sulfate and ammonia were added to the copper- and cadmium-free roasted body for reducing complexation leaching, so that zinc and nickel entered the solution in the form of ammonia complexes, resulting in zinc- and nickel-containing leachate and lead slag. Copper-cadmium leaching solution was subjected to electrowinning to remove copper, yielding cathode copper and copper-removed solution. Sodium carbonate was added to the copper-removed solution to precipitate and recover cadmium. Solvent extraction was performed on the zinc-nickel leaching solution to obtain zinc-containing back-extraction solution and nickel-containing back-extraction solution, respectively. Lead slag is used as a raw material for lead recycling and transferred to the pyrometallurgical smelting process.

2. The selective leaching and recycling method for polymetallic hazardous waste according to claim 1, characterized in that, The process of mixing multi-metal hazardous waste, sodium carbonate, and sulfur powder, followed by mechanochemical ball milling, yields a sulfurized activated powder, comprising: First, the polymetallic hazardous waste is dried and crushed into powder with a particle size of less than 0.5 mm. Then, sodium carbonate and sulfur powder are mixed in a mass ratio of 3:1 to 5:1 to prepare a vulcanizing agent. The powder and sulfiding agent are fed into a planetary ball mill, and zirconia balls and anhydrous ethanol are added. The mixture is ball-milled for 2 to 4 hours at a speed of 300 to 500 rpm under an argon atmosphere. After ball milling, anhydrous ethanol was recovered by vacuum distillation to obtain sulfidated activated powder.

3. The selective leaching and recycling method for polymetallic hazardous waste according to claim 1, characterized in that, The process involves mixing the sulfurized activated powder with water glass and a pore-forming agent, molding the mixture, and calcining it in a microwave field to obtain a porous sulfurized calcined body, comprising: The sulfurized activated powder, water glass, pore-forming agent and binder are mixed in a mass ratio of 100:8:15:5, an appropriate amount of water is added and stirred evenly, and then pressed into a cylindrical blank. The cylindrical blank is dried at 60°C to 80°C for 6 to 8 hours, and then placed in a microwave baking oven and baked in a nitrogen atmosphere by segmented heating. The first segment temperature is 200°C to 300°C, and the second segment temperature is 450°C to 550°C. After calcination, the material is cooled to room temperature in the furnace to obtain a porous sulfide calcined body.

4. The selective leaching and recycling method for polymetallic hazardous waste according to claim 1, characterized in that, The process involves placing a porous sulfidated calcined body in the intermediate chamber of an electrically powered diaphragm electrolytic cell. Sodium chloride solution is added to the anode chamber, and sodium hydroxide solution is added to the cathode chamber. Chlorine gas is introduced into the anode chamber, causing copper and cadmium in the slurry in the intermediate chamber to enter the solution as chloride complexes, resulting in a copper-cadmium leaching solution and a copper-cadmium-removed calcined body. A porous partition is installed at the bottom of the intermediate chamber of the diaphragm electrolyzer, and a porous sulfurized calcined body is placed on the porous partition. An anion exchange membrane is installed between the anode chamber and the intermediate chamber, and a cation exchange membrane is installed between the cathode chamber and the intermediate chamber. Chlorine gas is introduced into the anode chamber, causing the chlorine water generated in the anode chamber to permeate into the intermediate chamber through the anion exchange membrane; Sodium chloride and hydrochloric acid are added to the intermediate chamber to adjust the chloride ion concentration in the slurry in the intermediate chamber. Under stirring conditions, copper and cadmium are introduced into the solution in the form of chloride complexes. The slurry in the intermediate chamber was filtered to obtain a copper-cadmium leaching solution and a copper-cadmium-free calcined body.

5. The selective leaching and recycling method for polymetallic hazardous waste according to claim 1, characterized in that, The process involves adding ammonium sulfate and ammonia to the copper- and cadmium-free calcined body for reducing complex leaching, allowing zinc and nickel to enter the solution as ammonia complexes, resulting in a zinc- and nickel-containing leachate and lead slag. Add the copper- and cadmium-free calcined body to a reducing complexing leaching tank, first add ammonium sulfate and stir to dissolve, then add ammonia water to adjust the pH of the slurry to 8 to 10; Sodium sulfite and iron powder are added in three batches while stirring to maintain a reducing atmosphere in the slurry. This allows zinc and nickel to react with ammonia to form ammonia complexes that enter the solution, while lead remains in the slag as solid lead sulfate. After leaching, solid-liquid separation is performed to obtain zinc-nickel leaching solution and lead slag.

6. The selective leaching and recycling method for polymetallic hazardous waste according to claim 1, characterized in that, The process of electrowinning to remove copper from copper-containing cadmium leaching solution to obtain cathode copper and copper-removed solution, and adding sodium carbonate to the copper-removed solution to precipitate and recover cadmium includes: Using a titanium plate as the cathode and a plate-shaped lead-silver alloy as the anode, a copper-cadmium leaching solution is injected into an electrodeposition cell for electrodeposition, causing copper to be deposited at the cathode, resulting in cathode copper and copper removal solution. The copper removal solution is transferred to a precipitation reactor and heated to 50 to 70 degrees Celsius. Sodium carbonate solution is slowly added while stirring to precipitate cadmium as cadmium carbonate. Cadmium carbonate precipitate is separated by filtration, and the filtrate is returned to the intermediate chamber of the diaphragm electrolyzer as leachate for reuse.

7. The selective leaching and recycling method for polymetallic hazardous waste according to claim 1, characterized in that, The solvent extraction of the zinc-nickel leaching solution yields a zinc-containing back-extraction solution and a nickel-containing back-extraction solution, respectively, comprising: The zinc-nickel leaching solution was mixed with P204 extractant at a ratio of 1:1 to extract zinc, resulting in a zinc-containing organic phase and a nickel-containing raffinate. The zinc-containing organic phase was back-extracted with sulfuric acid to obtain a zinc-containing back-extract. The nickel-containing raffinate was mixed with LIX84 extractant to extract nickel, resulting in a nickel-containing organic phase. The nickel-containing organic phase was then back-extracted with hydrochloric acid to obtain a nickel-containing back-extract.

8. The selective leaching and recycling method for polymetallic hazardous waste according to claim 4, characterized in that, The process of introducing chlorine gas into the anode chamber, allowing the chlorine water generated in the anode chamber to permeate into the intermediate chamber through the anion exchange membrane, includes: Chlorine gas generated by an external chlorine source is pumped into a storage tank via a compression pump, and then introduced into the anode chamber from the bottom of the storage tank through a microporous aeration pipe. The sodium chloride solution in the anode chamber absorbs the chlorine gas to generate chlorine water. The flow rate of chlorine gas is controlled to maintain the redox potential of the slurry in the intermediate chamber between 0.2 volts and 0.6 volts, allowing chlorine molecules and hypochlorous acid in the chlorine water to pass through the anion exchange membrane into the intermediate chamber. The sulfides in the intermediate chamber are oxidized, releasing copper and cadmium ions. The copper and cadmium ions combine with chloride ions to form chloride complexes that enter the solution.

9. The selective leaching and recycling method for polymetallic hazardous waste according to claim 5, characterized in that, The process involves adding sodium sulfite and iron powder in three batches under stirring to maintain a reducing atmosphere in the slurry. This allows zinc and nickel to react with ammonia to form an ammonia complex, which then enters the solution. Lead remains in the slag as solid lead sulfate. The process includes: Sodium sulfite and iron powder were added for the first time, at amounts of 2% and 1% of the mass of the copper-cadmium-de-copper calcined body, respectively. After stirring for 20 minutes, the dissolved oxygen concentration in the slurry was measured. Sodium sulfite and iron powder were added a second time, at amounts of 1% and 0.5% of the mass of the copper-cadmium-removed calcined body, respectively, and stirred for 30 minutes. Sodium sulfite and iron powder were added for the third time, at amounts of 0.5% and 0.3% of the mass of the copper-cadmium-de-copper calcined body, respectively. The mixture was stirred for 1 hour to reduce the dissolved oxygen concentration in the slurry to below 0.1 mg / L.

10. The selective leaching and recycling method for polymetallic hazardous waste according to claim 6, characterized in that, The process of transferring the copper removal solution to a precipitation reactor, heating it to 50°C to 70°C, and slowly adding sodium carbonate solution while stirring, to precipitate cadmium as cadmium carbonate, includes: First, adjust the pH of the copper removal solution to 9 to 11 with sodium hydroxide solution, then add sodium carbonate solution at a rate of 5 to 10 ml per minute. During the addition of sodium carbonate solution, carbon dioxide gas is simultaneously introduced into the bottom of the precipitation reactor, so that the carbon dioxide reacts with cadmium ions in the copper removal solution to form cadmium carbonate precipitate. After precipitation, the cadmium carbonate precipitate was allowed to stand for 2 hours. The precipitate was then washed with water until the pH of the washing solution was less than 8. After filtration, the precipitate was dried at 105 degrees Celsius to obtain the cadmium carbonate product.

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