WO2024040910A1 - 一种从废旧锂离子电池中回收有价金属的方法 - Google Patents

一种从废旧锂离子电池中回收有价金属的方法 Download PDF

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WO2024040910A1
WO2024040910A1 PCT/CN2023/079351 CN2023079351W WO2024040910A1 WO 2024040910 A1 WO2024040910 A1 WO 2024040910A1 CN 2023079351 W CN2023079351 W CN 2023079351W WO 2024040910 A1 WO2024040910 A1 WO 2024040910A1
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lithium
sulfate
battery
waste
pyrolysis
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PCT/CN2023/079351
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English (en)
French (fr)
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周启
郑宇�
石泉清
刘勇奇
巩勤学
李长东
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广东邦普循环科技有限公司
湖南邦普循环科技有限公司
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Publication of WO2024040910A1 publication Critical patent/WO2024040910A1/zh

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    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • C22B7/006—Wet processes
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B01—PHYSICAL OR CHEMICAL PROCESSES OR APPARATUS IN GENERAL
    • B01D—SEPARATION
    • B01D53/00—Separation of gases or vapours; Recovering vapours of volatile solvents from gases; Chemical or biological purification of waste gases, e.g. engine exhaust gases, smoke, fumes, flue gases, aerosols
    • B01D53/34—Chemical or biological purification of waste gases
    • B01D53/46—Removing components of defined structure
    • B01D53/54—Nitrogen compounds
    • B01D53/58—Ammonia
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B01—PHYSICAL OR CHEMICAL PROCESSES OR APPARATUS IN GENERAL
    • B01D—SEPARATION
    • B01D53/00—Separation of gases or vapours; Recovering vapours of volatile solvents from gases; Chemical or biological purification of waste gases, e.g. engine exhaust gases, smoke, fumes, flue gases, aerosols
    • B01D53/34—Chemical or biological purification of waste gases
    • B01D53/74—General processes for purification of waste gases; Apparatus or devices specially adapted therefor
    • B01D53/77—Liquid phase processes
    • B01D53/78—Liquid phase processes with gas-liquid contact
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/02—Roasting processes
    • C22B1/06—Sulfating roasting
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/02—Obtaining nickel or cobalt by dry processes
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/04—Obtaining nickel or cobalt by wet processes
    • C22B23/0407—Leaching processes
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B26/00—Obtaining alkali, alkaline earth metals or magnesium
    • C22B26/10—Obtaining alkali metals
    • C22B26/12—Obtaining lithium
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/26—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds
    • C22B3/38—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds containing phosphorus
    • C22B3/384—Pentavalent phosphorus oxyacids, esters thereof
    • C22B3/3844—Phosphonic acid, e.g. H2P(O)(OH)2
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/26—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds
    • C22B3/38—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds containing phosphorus
    • C22B3/384—Pentavalent phosphorus oxyacids, esters thereof
    • C22B3/3846—Phosphoric acid, e.g. (O)P(OH)3
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B47/00—Obtaining manganese
    • H—ELECTRICITY
    • H01—ELECTRIC ELEMENTS
    • H01M—PROCESSES OR MEANS, e.g. BATTERIES, FOR THE DIRECT CONVERSION OF CHEMICAL ENERGY INTO ELECTRICAL ENERGY
    • H01M10/00—Secondary cells; Manufacture thereof
    • H01M10/54—Reclaiming serviceable parts of waste accumulators
    • Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02W—CLIMATE CHANGE MITIGATION TECHNOLOGIES RELATED TO WASTEWATER TREATMENT OR WASTE MANAGEMENT
    • Y02W30/00—Technologies for solid waste management
    • Y02W30/50—Reuse, recycling or recovery technologies
    • Y02W30/84—Recycling of batteries or fuel cells

Definitions

  • the invention belongs to the field of waste material recycling, and specifically relates to a method for recovering valuable metals from waste lithium-ion batteries.
  • the recovery of valuable metals from waste lithium-ion batteries mostly involves dissolving the metal ions through a wet recovery process of reducing acid leaching, and then recovering the metal ions using methods such as precipitation, extraction, crystallization and resynthesis.
  • the wet recovery route is The process is mature and adaptable, but there are problems such as long process, low metal recovery rate and high cost of auxiliary materials.
  • the industry also uses pyrometallurgy to recover valuable metals.
  • the pyrometallurgical process is to put the waste battery materials into a high-temperature pyrolysis furnace under high-temperature conditions, and control the reaction in different temperature ranges.
  • the waste lithium-ion battery cathode materials will undergo a series of physical and chemical changes when heated, and then according to the differences in the properties of the resulting products Realize the separation of valuable metal elements.
  • the pyrolysis recycling method is the simplest process for processing used batteries. It has a large battery processing capacity and a high degree of process automation. However, the pyrolysis treatment process has difficulties in temperature control of the pyrolysis furnace, high safety risks, and impurities in the black powder produced. There are problems such as high content, high alloy content in black powder, and difficulty in post-processing.
  • CN106505270A provides a method for recovering valuable metals from used lithium-ion battery cathode materials, including salt solution discharge; dismantling and separating the cathode sheets.
  • the positive electrode sheet is broken and separated from the positive electrode material and aluminum foil; the positive electrode material is mixed with the roasting agent ammonium sulfate and/or ammonium bisulfate and roasted at low temperature; the roasted material is immersed in water to separate the carbon and leachate; add a precipitant to the leachate, and use NH 3
  • the flue gas adjusts the pH, precipitates other metals except Li, and separates solids and liquids; uses flue gas containing NH 3 to adjust the pH of the filtrate, adds ammonium carbonate or ammonium bicarbonate, or blows in CO 2 gas to precipitate lithium to obtain lithium carbonate products. This method requires manual disassembly to obtain positive electrode sheets.
  • CN113764758A provides a method for pyrolysis and temperature control of waste lithium battery recycling.
  • the method includes the following steps: mixing waste battery raw materials with solid medium quartz sand, cinder or gravel, performing anaerobic pyrolysis, screening, and crushing. After sorting, the battery powder is obtained; the pyrolysis temperature can be controlled at 350-550°C.
  • This method introduces difficult-to-treat impurities such as quartz stone, which increases the processing capacity of subsequent wet recovery.
  • the amount of waste residue that needs to be processed increases sharply, increasing the processing cost and processing difficulty, and there is also the problem of low recovery rate of valuable metals.
  • one of the purposes of the present invention is to provide a method for recovering valuable metals from used lithium-ion batteries.
  • the method has a high degree of automation, the process is easy to control, and the recovery amount per batch is high. It is large, produces less waste residue and waste liquid, and has low material cost for recycling auxiliary materials.
  • the second object of the present invention is to provide a method for recovering valuable metals from used lithium-ion batteries for application in battery recycling.
  • a first aspect of the present invention provides a method for recovering valuable metals from used lithium-ion batteries, including the following steps:
  • S1 Discharge the used lithium-ion battery in a sulfate-containing discharge solution, then mix it with sulfate and/or sulfide and pyrolyze and roast it. Spray the sulfate-containing solution during the pyrolysis and roasting process to recover the battery black powder.
  • the waste lithium-ion batteries are first discharged through a discharge solution containing sulfate, and then the waste lithium-ion batteries and sulfates are crushed and mixed, and then sent to the pyrolysis furnace for pyrolysis roasting treatment.
  • Adding a sulfate-containing solution can take away the heat released during the pyrolysis process of used lithium-ion batteries through water evaporation and heat absorption, maintain the relative stability of the pyrolysis temperature in the pyrolysis furnace, and avoid excessive temperature rise in the pyrolysis furnace during the pyrolysis process.
  • Spraying the sulfate-containing solution in step S1 of the present invention can, on the one hand, take away the heat in the pyrolysis furnace through the evaporation of water and the decomposition of sulfate, preventing the temperature in the pyrolysis furnace from being too high, thereby reducing battery black powder
  • the content of alloys or metal elements and the content of other impurities can, on the one hand, take away the heat in the pyrolysis furnace through the evaporation of water and the decomposition of sulfate, preventing the temperature in the pyrolysis furnace from being too high, thereby reducing battery black powder.
  • spraying sulfate solution can supplement the sulfate during the pyrolysis process of used lithium-ion batteries, so that the nickel-cobalt-manganese lithium in the used lithium-ion batteries can be fully converted into lithium nickel-cobalt-manganese sulfate.
  • the sulfate-containing discharge solution further includes sodium chloride.
  • the waste lithium-ion battery is at least one of a ternary lithium-ion battery containing nickel, cobalt and manganese, a lithium cobalt oxide lithium-ion battery, a lithium manganate lithium-ion battery, and a lithium nickel-oxide lithium-ion battery.
  • the sulfate is at least one of ammonium sulfate, ammonium bisulfate, sodium bisulfate, sodium sulfate, potassium bisulfate, potassium sulfate, Na 2 S 2 O 3 and Na 2 S 2 O 7 ; further preferably Preferably, the sulfate is ammonium sulfate or ammonium bisulfate; further preferably, the sulfate is ammonium sulfate.
  • Ni and Co can be directly converted into corresponding sulfates under sulfate roasting conditions, while the process of converting Mn element into sulfates under sulfate roasting conditions is as follows:
  • the sulfide is at least one of CuS, MnS, NiS, Na 2 S, and sulfur S8.
  • the concentration of the sulfate solution in step S1 is 100-750g/L; further preferably, the concentration of the sulfate solution in step S1 is 300-500g/L.
  • the discharge time is 1-6h; further preferably, the discharge time is 2-4h.
  • the step of mixing with sulfate and/or sulfide in step S1 is: mixing and pulverizing waste lithium ion batteries with sulfate and/or sulfide to prepare waste lithium ion battery with sulfate and/or sulfide. or a mixture of sulfides.
  • the temperature of the crushing step is 50-250°C; further preferably, the temperature of the crushing step is 60-200°C.
  • the crushing time is 0.5-6h; further preferably, the crushing time is 1-6h.
  • the step of extracting lithium to obtain lithium carbonate and nickel cobalt manganese sulfate is: extract the battery black powder using the leachate after solvent leaching to obtain a nickel cobalt manganese sulfate solution and a lithium sulfate solution, and then remove the lithium sulfate solution. After impurity, it is concentrated and lithium is precipitated to obtain lithium carbonate and nickel cobalt manganese sulfate.
  • the extraction agent is at least one of P507 extraction agent and P204 extraction agent.
  • the impurity removal step is: adding calcium oxide to the lithium sulfate solution, adjusting the pH to 8-9, and the temperature to 60-90°C, and then adding carbonate and filtering.
  • the function of the carbonate added in this step is to remove calcium impurities in the lithium sulfate solution.
  • the material ratio of carbonate and calcium oxide is 1:1.
  • the concentration step is: concentrating the lithium sulfate solution until the lithium ion content is 8-15g/L.
  • the pH of the lithium precipitation step is 10.5-12.5.
  • the temperature of the lithium precipitation step is 30-90°C; further preferably, the temperature of the lithium precipitation step is 40-80°C; still further preferably, the temperature of the lithium precipitation step is 60-80°C. .
  • the time of the lithium precipitation step is 0.5-3 hours; further preferably, the time of the lithium precipitation step is 1-3 hours.
  • the sulfate-containing solution in step S1 is a sulfate-containing solution obtained by passing the waste gas generated in the pyrolysis roasting step into the mixed solution of the waste liquid generated in the lithium precipitation step and sulfuric acid; or, recycling A sulfate-containing solution is obtained from the waste liquid generated in the discharge step.
  • the mixture of waste liquid and sulfuric acid produced in the lithium precipitation step can be used to absorb the ammonia-containing waste gas in the waste gas generated by the pyrolysis furnace. Through this step, the waste liquid of lithium precipitation can be recycled and the sulfate can be recycled.
  • the NH3- containing waste gas discharged from the pyrolysis furnace is directly introduced into the waste liquid produced by the lithium precipitation step of adding sulfuric acid to acidify, so as to recover the ammonia in the tail gas generated by the pyrolysis roasting step to generate an ammonium sulfate solution, and then absorb the lithium precipitation from the tail gas
  • the waste liquid is sprayed into the pyrolysis furnace for recycling, which reduces the consumption of ammonium sulfate auxiliary materials and reduces production costs; the ammonium sulfate solution can also be crystallized and precipitated to recover the ammonium sulfate, and then put into step S1 to crush and mix with used lithium-ion batteries. .
  • the ratio of the amount of sulfuric acid to the amount of sulfate and/or sulfide in step S1 is (0.5-2.5):1; further preferably, the amount of sulfuric acid to the amount of sulfate in step S1
  • the amount ratio of the amount of sulfate and/or sulfide is (1-2.5):1; further preferably, the amount ratio of the amount of sulfuric acid to the amount of sulfate and/or sulfide in step S1 is (1-2):1.
  • the temperature of the waste liquid produced in the lithium precipitation step is 60-80°C.
  • the spray flow rate of the sulfate-containing solution is 0.5-3.5m 3 /h; further preferably, the spray flow rate of the sulfate-containing solution is 1-3 m 3 /h.
  • the mass ratio of the sulfate in the sulfate and/or sulfide to the used lithium ion battery is (0.6-1.2):1.
  • the pyrolysis roasting temperature is 350-600°C; further preferably, the pyrolysis roasting temperature is 350-500°C; still further preferably, the pyrolysis roasting temperature is 400-500°C.
  • the pyrolysis roasting time is 30-150 min; further preferably, the pyrolysis roasting time is 40-120 min.
  • the pyrolysis roasting is performed in the presence of oxygen-containing gas.
  • the oxygen-containing gas is air or oxygen.
  • the step of recovering battery black powder is: crushing and screening the pyrolysis roasted product to obtain battery black powder.
  • the crushing time is 30-150 min; further preferably, the crushing time is 40-120 min.
  • copper and aluminum slag is also obtained after the crushing and screening.
  • the leaching time is 50-200 min; further preferably, in step S2, the leaching time is 60-150 min.
  • the leaching temperature is 30-90°C; further preferably, in step S2, the leaching temperature is 40-80°C.
  • the step of using a solvent to leach metal ions from battery black powder results in a leachate and leach residue.
  • the solid-liquid ratio of the solvent to battery black powder is 2-10L/kg.
  • the solvent is water.
  • the leaching residue is transferred to a wet acid leaching reduction system for treatment.
  • a second aspect of the present invention is to provide the application of the above method for recovering valuable metals from used lithium ion batteries in battery recycling.
  • the method of the present invention uses a sulfate solution to discharge used lithium-ion batteries, so that the sulfate and the positive electrode material in the used lithium-ion batteries are fully infiltrated and contacted, which can promote the active discharge of the used lithium-ion batteries. sulfation of valence metals and improve the pyrolysis roasting effect.
  • the method in the present invention sprays a sulfate-containing solution during the pyrolysis roasting process.
  • the heat generated by the pyrolysis of used lithium-ion batteries can be taken away through water evaporation and pyrolysis of sulfate, thereby achieving controlled pyrolysis.
  • Roasting temperature to avoid excessive pyrolysis roasting temperature, which may lead to the production of a large amount of metal alloys and impurities.
  • the method of the present invention mixes the waste liquid produced by lithium precipitation with sulfuric acid and then uses it to absorb the ammonia-containing waste gas produced during the pyrolysis roasting process, thereby realizing the recovery treatment of the lithium precipitation waste liquid and pyrolysis waste gas, and at the same time recycling
  • the treated discharge solution and lithium precipitation waste liquid are sprayed during the pyrolysis roasting process.
  • it can reduce the pyrolysis roasting temperature and achieve precise control of the pyrolysis roasting furnace temperature; on the other hand, it can supplement sulfuric acid for the pyrolysis of waste lithium-ion batteries.
  • the salt component improves the sulfation of valuable metals in used lithium-ion batteries, reduces the impurity content, greatly reduces the amount of slag of battery black powder after water leaching, and can reduce the amount of processing required for acid leaching reduction wet recycling. , reduce the amount of auxiliary materials used in subsequent process processes and reduce process costs.
  • the method in the present invention directly discharges the entire waste lithium-ion battery material and then pyrolyzes and roasts it to recover valuable metals. There is no need to separate the positive electrode material from the waste lithium-ion battery material. It has a high degree of automation, reduces labor costs, and has a simple recycling process. , easy to operate, saving process costs. At the same time, the discharge waste liquid, lithium precipitation waste liquid, and pyrolysis roasting waste gas generated in this method can be recycled and used. The amount of waste residue generated is small, the post-processing process is simple, and it is more in line with green recycling. Process requirements.
  • harmful organic substances such as the electrolyte in the used lithium-ion battery that enter the sulfate solution can be thermally decomposed by the pyrolysis furnace, which can reduce the amount of discharge waste liquid containing electrolyte.
  • the method in the present invention can achieve selective extraction of nickel and cobalt, reduce sulfation of manganese, and achieve selective extraction of valuable metals by adjusting the mass ratio of sulfate and waste lithium-ion batteries, combined with the adjustment of furnace temperature.
  • Figure 1 is an SEM image of the battery black powder obtained by the method in Example 3.
  • Figure 2 is an SEM image of the leaching residue obtained by the method in Example 3.
  • Figure 3 is a schematic flow chart of the method for recovering valuable metals from used lithium ion batteries in Examples 1-4.
  • the testing method for the valuable metal content in the battery black powder in Examples 1-4 and Comparative Examples 1-2 of the present invention is:
  • the testing method for the main metal ion concentration in the leachate in Examples 1-4 and Comparative Examples 1-2 of the present invention is: after dilution, use an ICP inductively coupled plasma mass spectrometer to test the concentration of each metal ion.
  • the method of recovering valuable metals from used lithium-ion batteries in this example includes the following steps:
  • step (2) Send the broken and mixed battery materials in step (1) into a pyrolysis furnace for pyrolysis and roasting in an air atmosphere of 350°C.
  • the roasting time is 120 minutes.
  • the recovered liquid discharge solution is sprayed Into the pyrolysis furnace, the spray flow rate is 0.5m 3 /h, and the ammonia-containing waste gas generated by roasting is collected;
  • step (3) Crush the pyrolyzed and roasted material in step (2) in a crusher for 40 minutes, and screen to obtain battery black powder and copper and aluminum slag; take a small amount of battery black powder and measure the valuable metal content in the battery black powder , the measurement results are shown in Table 1.
  • step (3) Mix the battery black powder obtained in step (3) with water at a solid-liquid ratio of 1kg:10L, stir at 30°C for 60 minutes, and then separate the solid and liquid to obtain leaching residue and leachate.
  • the obtained 320.5kg leaching residue Transfer to wet acid leaching reduction system for treatment. Take a small amount of the leachate and measure the concentration of the main ions in the leachate. The results are shown in Table 2.
  • the leaching rate of the main ions in the leachate is estimated based on the theoretical content composition of the used battery cathode material (accounting for 40% of the total battery mass).
  • the main ions in the leachate are The estimated leaching rates of ions are shown in Table 3.
  • the main ion concentration and ion leaching rate in the leach solution are respectively: Li + concentration is 3.56g/L, leaching rate is 95%; Ni 2+ concentration is 10.35g/L, leaching rate is 72%; The Co 2+ concentration is 4.06g/L, and the leaching rate is 70%; the Mn 2+ concentration is 2.61g/L, and the leaching rate is 30%;
  • step (2) Pour the ammonia-containing waste gas generated by pyrolysis in the pyrolysis furnace in step (2) into the lithium precipitation waste liquid, and add 525L concentrated sulfuric acid to the lithium precipitation waste liquid to absorb the ammonia-containing waste gas generated by the pyrolysis furnace, and then Spray the lithium precipitation waste liquid into the pyrolysis furnace at a flow rate of 2m 3 /h, and increase the roasting circulation volume of ammonium sulfate in the pyrolysis furnace.
  • the waste gas and liquid produced in all steps can be effectively recycled and reused.
  • the ammonia-containing waste gas generated by pyrolysis in the pyrolysis furnace in step (2) can be passed into the lithium precipitation waste liquid in step (8), and 525L of concentrated sulfuric acid is added to the lithium precipitation waste liquid to fully absorb the ammonia-containing waste gas.
  • the lithium precipitation waste liquid that has absorbed ammonia-containing waste gas can be put back into the pyrolysis furnace (spray into the pyrolysis furnace at a flow rate of 2m3 /h) to promote the roasting circulation of ammonium sulfate in the pyrolysis furnace in the previous step and increase the Metal recycling efficiency.
  • the method of recovering valuable metals from used lithium-ion batteries includes the following steps:
  • step (2) Send the crushed and mixed battery materials in step (1) into a pyrolysis furnace for pyrolysis and roasting in an air atmosphere at 400°C.
  • the roasting time is 100 minutes.
  • the recovered liquid discharge solution is sprayed into In the pyrolysis furnace, the spray flow rate is 2m 3 /h, and the ammonia-containing waste gas generated by roasting is collected;
  • step (3) Crush the pyrolyzed roasted material in step (2) for 100 minutes in a crusher, and screen to obtain battery black powder and copper and aluminum slag; take a small amount of battery black powder and measure the valuable metal content in the battery black powder , the measurement results are shown in Table 1.
  • step (3) Mix the battery black powder obtained in step (3) with water at a solid-liquid ratio of 1kg:10L, stir at 50°C for 120 minutes, and then separate the solid and liquid to obtain leaching residue and leachate.
  • the obtained 294.2kg leaching residue Transfer to wet acid leaching reduction system for treatment. Take a small amount of the leachate and measure the concentration of the main ions in the leachate. The results are shown in Table 2.
  • the leaching rate of the main ions in the leachate is estimated based on the theoretical content composition of the used battery cathode material (accounting for 40% of the total battery mass).
  • the main ions in the leachate are The estimated ion leaching rate is shown in Table 3.
  • the main ion concentration and ion leaching rate in the leachate are respectively: Li + concentration is 3.62g/L, leaching rate is 98%; Ni 2+ The concentration is 12.22g/L, and the leaching rate is 85%; the Co 2+ concentration is 4.76g/L, and the leaching rate is 82%; the Mn 2+ concentration is 4.35g/L, and the leaching rate is 50%;
  • step (2) Pour the ammonia-containing waste gas generated by pyrolysis in the pyrolysis furnace in step (2) into the lithium precipitation waste liquid, and add 525L concentrated sulfuric acid to the lithium precipitation waste liquid to absorb the ammonia-containing waste gas generated by the pyrolysis furnace, and then Spray the lithium precipitation waste liquid into the pyrolysis furnace at a flow rate of 3m 3 /h, and increase the roasting circulation volume of ammonium sulfate in the pyrolysis furnace.
  • the waste gas and liquid produced in all steps can be effectively recycled and reused.
  • the ammonia-containing waste gas generated by pyrolysis in the pyrolysis furnace in step (2) can be passed into the lithium precipitation waste liquid in step (8), and 525L of concentrated sulfuric acid is added to the lithium precipitation waste liquid to fully absorb the ammonia-containing waste gas.
  • the lithium precipitation waste liquid that has absorbed ammonia-containing waste gas can be put back into the pyrolysis furnace (spray into the pyrolysis furnace at a flow rate of 3m3 /h) to promote the roasting circulation of ammonium sulfate in the pyrolysis furnace in the previous step and increase the Metal recycling efficiency.
  • the method of recovering valuable metals from used lithium-ion batteries includes the following steps:
  • step (2) Send the broken and mixed battery materials in step (1) into a pyrolysis furnace for pyrolysis and roasting in an air atmosphere at 500°C.
  • the roasting time is 80 minutes.
  • the recovered liquid discharge solution is sprayed into In the pyrolysis furnace, the spray flow rate is 3.5m 3 /h, and the ammonia-containing waste gas generated by roasting is collected;
  • step (3) Crush the pyrolyzed roasted material in step (2) in a crusher for 40 minutes, and screen to obtain battery black powder and copper and aluminum slag; take a small amount of battery black powder and measure the valuable metal content in the battery black powder.
  • the measurement results are shown in Table 1.
  • the main ion concentration and ion leaching rate in the leachate are respectively: Li + concentration is 3.64g/L, leaching rate is 98%; Ni 2+ concentration is 12.94g/L, leaching rate is 90% ; Co 2+ concentration is 5.30g/L, leaching rate is 91%; Mn 2+ concentration is 6.96g/L, leaching rate is 85%;
  • step (2) Pass the ammonia-containing waste gas generated by pyrolysis in the pyrolysis furnace in step (2) into the lithium precipitation waste liquid, and add 626L concentrated sulfuric acid to the lithium precipitation waste liquid to absorb the ammonia-containing waste gas generated by the pyrolysis furnace, and then Spray the lithium precipitation waste liquid into the pyrolysis furnace at a flow rate of 3.5m 3 /h, and increase the roasting circulation volume of ammonium sulfate in the pyrolysis furnace.
  • the waste gas and liquid produced in all steps can be effectively recycled and reused.
  • the ammonia-containing waste gas generated by pyrolysis in the pyrolysis furnace in step (2) can be passed into the lithium precipitation waste liquid in step (8), and concentrated sulfuric acid is added to the lithium precipitation waste liquid to fully absorb the ammonia-containing waste gas.
  • the lithium precipitation waste liquid that has absorbed ammonia-containing waste gas can be put back into the pyrolysis furnace (spray into the pyrolysis furnace at a flow rate of 3.5m3 /h) to promote the roasting circulation of ammonium sulfate in the pyrolysis furnace in the previous step. Improve metal recycling efficiency.
  • the SEM images of the battery black powder obtained in step (3) and the leached slag obtained in step (4) were tested respectively.
  • the SEM image of the battery black powder is shown in Figure 1
  • the SEM image of the leached slag is shown in Figure 2. Comparing Figure 1 and Figure 2, it can be seen that there are a large number of caves where metal ions are leached out on the leaching residue.
  • the method of recovering valuable metals from used lithium-ion batteries includes the following steps:
  • step (2) Send the broken and mixed battery materials in step (1) into a pyrolysis furnace for pyrolysis and roasting in an air atmosphere at 600°C.
  • the roasting time is 60 minutes.
  • the recovered liquid discharge solution is sprayed into In the pyrolysis furnace, the spray flow rate is 3.5m 3 /h, and the ammonia-containing waste gas generated by roasting is collected;
  • step (3) Crush the pyrolyzed roasted material in step (2) in a crusher for 40 minutes, and screen to obtain battery black powder and copper and aluminum slag; take a small amount of battery black powder and measure the valuable metal content in the battery black powder.
  • the measurement results are shown in Table 1.
  • step (3) Mix the battery black powder in step (3) with water at a solid-liquid ratio of 1kg:10L, stir at 90°C for 150 minutes, and then separate the solid and liquid to obtain leaching residue and leachate.
  • the obtained 235.3kg leaching residue Transfer to wet acid leaching reduction system for treatment. Take a small amount of the leachate and measure the concentration of the main ions in the leachate. The results are shown in Table 2.
  • the leaching rate of the main ions in the leachate is estimated based on the theoretical content composition of the used battery cathode material (accounting for 40% of the total battery mass).
  • the main ions in the leachate are The estimated ion leaching rate is shown in Table 3.
  • the main ion concentration and ion leaching rate in the leachate are respectively: Li + concentration is 3.45g/L, leaching rate is 95%; Ni 2+ concentration is 13.35g/L, leaching rate is 80% ; Co 2+ concentration is 5.30g/L, leaching rate is 73%; Mn 2+ concentration is 6.84g/L, leaching rate is 70%;
  • step (2) Pass the ammonia-containing waste gas generated by pyrolysis in the pyrolysis furnace in step (2) into the lithium precipitation waste liquid, and add 626L concentrated sulfuric acid to the lithium precipitation waste liquid to absorb the ammonia-containing waste gas generated by the pyrolysis furnace, and then Spray the lithium precipitation waste liquid into the pyrolysis furnace at a flow rate of 3.5m 3 /h, and increase the roasting circulation volume of ammonium sulfate in the pyrolysis furnace.
  • the waste gas and liquid produced in all steps can be effectively recycled and reused.
  • the ammonia-containing waste gas generated by pyrolysis in the pyrolysis furnace in step (2) can be passed into the lithium precipitation waste liquid in step (8), and concentrated sulfuric acid is added to the lithium precipitation waste liquid to fully absorb the ammonia-containing waste gas.
  • the lithium precipitation waste liquid that has absorbed ammonia-containing waste gas can be put back into the pyrolysis furnace (spray into the pyrolysis furnace at a flow rate of 3.5m3 /h) to promote the roasting circulation of ammonium sulfate in the pyrolysis furnace in the previous step. Improve metal recycling efficiency.
  • the method of recovering valuable metals from used lithium-ion batteries includes the following steps:
  • step (2) Send the broken and mixed battery materials in step (1) into a pyrolysis furnace for pyrolysis in an air atmosphere.
  • the roasting time is 120 minutes; because there are no corresponding temperature control measures during the pyrolysis process, the furnace temperature rises to above 750°C. .
  • step (3) Crush the pyrolyzed material in step (2) in a crusher for 40 minutes, and screen to obtain battery black powder and copper and aluminum slag; measure the valuable metal content in the battery black powder, and record the measurement results in the table 1 in.
  • the main ion concentration and ion leaching rate in the leachate are respectively: Li + concentration is 0.18g/L, leaching rate is 4.8%; Ni 2+ concentration is 0.33g/ L, the leaching rate is 2.3%; the Co 2+ concentration is 0.05g/L, the leaching rate is 0.8%; the Mn 2+ concentration is 0.01g/L, the leaching rate is 0.2%.
  • the method of recovering valuable metals from used lithium-ion batteries includes the following steps:
  • step (2) Send the broken and mixed battery materials in step (1) into a pyrolysis furnace for pyrolysis and roasting in an air atmosphere of 500°C.
  • the roasting time is 120 minutes; the ammonium sulfate solution is not sprayed and discharged for temperature control.
  • the pyrolysis furnace The temperature will reach 700°C.
  • step (3) Crush the pyrolyzed roasted material in step (2) for 60 minutes in a crusher, and screen to obtain battery black powder and copper and aluminum slag; measure the valuable metal content in the battery black powder, and record the measurement results in in FIG. 1.
  • the main ion concentration and ion leaching rate in the leachate are respectively: Li + concentration is 3.37g/L, leaching rate is 90%; Ni 2+ concentration is 10.11g/ L, the leaching rate is 71%; the Co 2+ concentration is 3.86g/L, the leaching rate is 68%; the Mn 2+ concentration is 5.39g/L, the leaching rate is 62%;
  • Example 1-4 can control the temperature of the pyrolysis furnace by spraying the sulfate solution in the pyrolysis furnace, thereby reducing Ni, The content of Co and Mn elements can avoid the deterioration of waste ternary lithium-ion battery materials due to excessive temperature in the pyrolysis furnace. Ni, Co and Mn elements generate more alloy materials.
  • the pyrolysis roasting temperature used in Example 4 is 600°C.
  • Example 4 The content of nickel, cobalt and manganese elements and alloys in the battery black powder are significantly higher than those in Examples 1-3, which further shows that the calcination temperature of 350-500°C can reduce the content of nickel, cobalt and manganese elements and alloys in the battery black powder.
  • the method in Examples 1-4 discharges the waste ternary lithium-ion battery in an ammonium sulfate solution, so that the sulfate can fully infiltrate the waste ternary lithium-ion battery material, and at the same time Combined with spraying sulfate solution in the pyrolysis furnace to control the temperature of the pyrolysis furnace, the sulfation of Li, Ni, Co, and Mn elements can be significantly improved, and the recovery rate of Li, Ni, Co, and Mn can be increased.
  • the method in Examples 1-4 can further improve the sulfation of Li, Ni, Co, and Mn elements by spraying the sulfate solution in the pyrolysis furnace to adjust the temperature of the pyrolysis furnace.
  • the method in Examples 1-4 discharges the waste ternary lithium-ion battery in an ammonium sulfate solution, so that the sulfate can fully infiltrate the waste ternary lithium-ion battery material, and at the same time Combined with spraying sulfate solution in the pyrolysis furnace to control the furnace temperature, the leaching rate of Li, Ni, Co, and Mn ions can be significantly improved, and the recovery rate of Li, Ni, Co, and Mn can be greatly improved. It greatly reduces the content of leaching residue, reduces the processing volume of leaching residue and the Slag treatment costs. Compared with Comparative Example 2, the method in Examples 1-4 can further improve the leaching rate of Li, Ni, Co, and Mn ions by spraying the sulfate solution in the pyrolysis furnace to adjust the temperature of the pyrolysis furnace.
  • Example 3 controlled the pyrolysis furnace temperature at 500°C by spraying ammonium sulfate solution, sulfate and waste ternary lithium-ion battery materials The mass ratio is controlled at 1:1, which can further improve the selective sulfate conversion of nickel, cobalt and manganese elements.
  • Example 1 controls the temperature of the pyrolysis furnace at 350°C by spraying ammonium sulfate solution, and the mass ratio of sulfate and waste ternary lithium-ion battery materials is controlled at 1:1, which can make Mn
  • the converted sulfate is reduced, and nickel and cobalt ions are mainly selectively leached.

Abstract

本发明公开了一种从废旧锂离子电池中回收有价金属的方法，该方法包括以下步骤：S1：将废旧锂离子电池在含硫酸盐的放电溶液中放电，然后与硫酸盐和/或硫化物混合并热解焙烧，在热解焙烧过程中喷淋含硫酸盐的溶液，回收电池黑粉；S2：使用溶剂浸取电池黑粉中的金属离子，萃取沉锂得到碳酸锂和硫酸镍钴锰。本发明中的方法使用硫酸盐溶液对废旧锂离子电池进行放电处理，可以促进废旧锂离子电池中的有价金属的硫酸盐化，并提高热解焙烧效果。本发明中的方法通过在热解焙烧过程中喷淋含硫酸盐的溶液，实现控制热解焙烧温度，避免热解焙烧温度过高而导致产生大量的金属合金和杂质。

Description

一种从废旧锂离子电池中回收有价金属的方法 技术领域
本发明属于废旧材料回收领域，具体涉及一种从废旧锂离子电池中回收有价金属的方法。
背景技术
目前，从废旧锂离子电池回收有价金属，大多是通过还原酸浸的湿法回收工艺将金属离子溶解出来，再用沉淀、萃取、结晶和再合成等方法回收金属离子，湿法回收路线的工艺成熟、适应性好，但存在流程较长、金属回收率偏低和辅料用料成本较高等问题。除了采用溶解法进行有价金属回收，行业中还使用火法冶金来回收有价金属。火法冶金工艺是在高温条件下，将电池废旧材料投入高温热解炉内，控制反应在不同的温度段内，废旧锂离子电池正极材料受热发生一系列物理化学变化，再根据所得产物性质差异实现有价金属元素的分离。火法回收方法是一种最简单的处理废旧电池的工艺，电池处理量大、工艺自动化程度高，但火法处理过程中存在热解炉控温困难、安全风险大、产出黑粉中杂质含量高、黑粉中合金含量高、后处理困难等问题。
目前比较常用的回收有价金属的技术是火法湿法联合回收工艺，CN106505270A提供了一种从废旧锂离子电池正极材料中回收有价金属的方法，包括盐溶液放电；拆解分离出正极片；正极片破碎分离正极材料和铝箔；正极材料与焙烧剂硫酸铵和/或硫酸氢铵混合低温焙烧；焙烧料水浸，分离得到碳和浸出液；向浸出液中加入沉淀剂，并使用含NH3烟气调节pH，沉淀除Li以外的其他金属，固液分离；使用含NH3烟气调节滤液的pH，加入碳酸铵或碳酸氢铵或者鼓入CO2气体，沉锂，得到碳酸锂产品。该方法需要人工拆解得到正极片，废电池处理量小，人工成本高，不利于进行大规模化生产，后续的水浸出渣中残留的正极材料未得到有效的回收利用，造成资源的浪费。CN113764758A提供了一种废旧锂电池回收热解控温的方法，该方法包括以下步骤：将废旧电池原料和固体介质石英砂、煤渣或碎石料混合，进行无氧热解，筛分，破碎，分选，即得电池粉；热解的温度可控制在350-550℃。该方法引入石英石等难处理的杂质，增加后续湿法回收的处理量，需处理的废渣量急剧增加，增加了处理成本和处理难度，而且存在有价金属回收率偏低的问题。
因此，亟需开发一种新的从废旧锂离子电池回收有价金属的方法，以克服传统方法中湿法回收工艺路线长，辅料用量大，金属回收率低，电池预处理段自动化程度低、火法回收预 处理中热解炉温度难以控制等电池回收处理领域亟待解决的普遍性难题。
发明内容
为了克服上述现有技术存在的问题，本发明的目的之一在于提供一种从废旧锂离子电池中回收有价金属的方法，该方法的自动化程度高，工艺过程易于控制、每批次回收量大、产生的废渣和废液量少，回收辅料用料成本低。
本发明的目的之二在于提供一种从废旧锂离子电池中回收有价金属的方法在电池回收中的应用。
为了实现上述目的，本发明所采取的技术方案是：
本发明的第一个方面提供了一种从废旧锂离子电池中回收有价金属的方法，包括以下步骤：
S1：将废旧锂离子电池在含硫酸盐的放电溶液中放电，然后与硫酸盐和/或硫化物混合并热解焙烧，在热解焙烧过程中喷淋含硫酸盐的溶液，回收电池黑粉；
S2：使用溶剂浸取电池黑粉中的金属离子，萃取沉锂得到碳酸锂和硫酸镍钴锰。
本发明中，废旧锂离子电池先通过含硫酸盐的放电溶液进行放电处理，再将废旧锂离子电池与硫酸盐破碎混合，送入热解炉进行热解焙烧处理，通过向热解炉内喷入含硫酸盐的溶液，可以通过水蒸发吸热带走废旧锂离子电池热解过程中放出的热量，保持热解炉内热解温度的相对稳定，避免热解炉在热解过程中温度大幅度升高而无法实现控温，进而减少了因热解炉内的温度过高使废旧锂离子电池中的Ni、Co和Mn元素生成较多的合金料，不利于电池黑粉的后续处理。本发明步骤S1中喷淋含硫酸盐的溶液，一方面可以通过水的蒸发和硫酸盐的分解过程带走热解炉中的热量，避免热解炉内的温度过高，从而降低电池黑粉中的合金或金属单质的含量和其他杂质的含量。另一个方面，喷淋硫酸盐溶液可以实现对废旧锂离子电池热解过程补充硫酸盐，使废旧锂离子电池中的镍钴锰锂充分转化成硫酸镍钴锰锂。
优选地，所述含硫酸盐的放电溶液中还包括氯化钠。
优选地，所述废旧锂离子电池为含镍钴锰的三元锂离子电池、钴酸锂锂离子电池、锰酸锂锂离子电池、镍酸锂锂离子电池中的至少一种。
优选地，所述硫酸盐为硫酸铵、硫酸氢铵、硫酸氢钠、硫酸钠、硫酸氢钾、硫酸钾、Na2S2O3、Na2S2O7中的至少一种；进一步优选地，所述硫酸盐为硫酸铵或硫酸氢铵；再进一步优选地，所述硫酸盐为硫酸铵。
当硫酸盐为硫酸铵时，硫酸铵热分解产生SO2可以使废旧锂离子电池热力学不稳定，生成相应的硫酸盐，具体反应式如下：
(NH4)2SO4→NH3+(NH4)HSO4
2(NH4)HSO4→2NH3+2SO2+2H2O+O2
2LiNixCoyMn1-x-yO2+2SO2→Li2SO4+2xNiSO4+2yCoSO4+2(1-x-y)MnSO4
2LiNixCoyMn1-x-yO2+2(NH4)HSO4→Li2SO4+2xNiSO4+2yCoSO4+
2(1-x-y)MnSO4+2H2O+2NH3+O2
Ni和Co在硫酸盐焙烧条件下可以直接转换成对应的硫酸盐，而Mn元素在硫酸盐焙烧条件下转化成硫酸盐的过程如下：
4LiMn2O4+10(NH4)2SO4→2(NH4)2Mn2(SO4)3+4Li(NH4)SO4+4MnO2+12NH3+6H2O+O2
5MnO2+5(NH4)2SO4→(NH4)2Mn2(SO4)3+Mn3O4+(NH4)3H(SO4)2+5NH3+2H2O+2O2
LiNixCoyMnzO2→MnO2→Mn2O3→MnSO4→Na2Mn(SO4)2
优选地，所述硫化物为CuS、MnS、NiS、Na2S、硫磺S8中的至少一种。
优选地，所述步骤S1中的硫酸盐溶液的浓度为100-750g/L；进一步优选地，所述步骤S1中的硫酸盐溶液的浓度为300-500g/L。
优选地，所述放电时间为1-6h；进一步优选地，所述放电时间为2-4h。
优选地，所述步骤S1中的与硫酸盐和/或硫化物混合的步骤为：将废旧锂离子电池与硫酸盐和/或硫化物混合并粉碎，制得废旧锂离子电池与硫酸盐和/或硫化物的混合物。
优选地，所述粉碎步骤的温度为50-250℃；进一步优选地，所述粉碎步骤的温度为60-200℃。
优选地，所述粉碎时间为0.5-6h；进一步优选地，所述粉碎时间为1-6h。
优选地，所述萃取沉锂得到碳酸锂和硫酸镍钴锰的步骤为：将电池黑粉使用溶剂浸取后的浸出液萃取，得到硫酸镍钴锰溶液和硫酸锂溶液，然后对硫酸锂溶液除杂后浓缩，再沉锂，得到碳酸锂和硫酸镍钴锰。
优选地，所述萃取步骤中，萃取剂为P507萃取剂、P204萃取剂中的至少一种。
优选地，所述除杂步骤为：在硫酸锂溶液中加入氧化钙，调节pH为8-9，温度为60-90℃，然后再加入碳酸盐并过滤。此步骤中加入的碳酸盐的作用是除去硫酸锂溶液中的钙杂质。
优选地，所述碳酸盐和氧化钙的物质的量比为1:1。
优选地，所述浓缩步骤为：将硫酸锂溶液浓缩至锂离子含量为8-15g/L。
优选地，所述沉锂步骤的pH为10.5-12.5。
优选地，所述沉锂步骤的温度为30-90℃；进一步优选地，所述沉锂步骤的温度为40-80℃；再进一步优选地，所述沉锂步骤的温度为60-80℃。
优选地，所述沉锂步骤的时间为0.5-3h；进一步优选地，所述沉锂步骤的时间为1-3h。
优选地，所述步骤S1中的含硫酸盐的溶液为：将热解焙烧步骤产生的废气通入沉锂步骤产生的废液和硫酸的混合液中得到的含硫酸盐的溶液；或，回收所述放电步骤产生的废液得到的含硫酸盐的溶液。沉锂步骤产生的废液和硫酸的混合液可以用于吸收热解炉产生的废气中的含氨废气，通过该步骤可以实现沉锂废液的回收利用，实现硫酸盐的循环使用。热解炉排出的含NH3废气直接导入加入硫酸酸化的沉锂步骤产生的废液中，实现回收热解焙烧步骤产生的尾气中的氨气生成硫酸铵溶液，再将吸收尾气后的沉锂废液喷淋进入热解炉循环利用，减少硫酸铵辅料的消耗量，可以降低生产成本；也可以将硫酸铵溶液进行结晶沉淀后回收硫酸铵，再投入步骤S1中与废旧锂离子电池破碎混合。
优选地，所述硫酸的用量与步骤S1中硫酸盐和/或硫化物的用量的物质的量比为(0.5-2.5)：1；进一步优选地，所述硫酸的用量与步骤S1中硫酸盐和/或硫化物的用量的物质的量比为(1-2.5)：1；再进一步优选地，所述硫酸的用量与步骤S1中硫酸盐和/或硫化物的用量的物质的量比为(1-2)：1。
优选地，所述沉锂步骤产生的废液的温度为60-80℃。
优选地，所述含硫酸盐的溶液的喷淋流量为0.5-3.5m3/h；进一步优选地，所述含硫酸盐的溶液的喷淋流量为1-3m3/h。
优选地，所述硫酸盐和/或硫化物中的硫酸根与废旧锂离子电池的质量比为(0.6-1.2)：1。
优选地，所述热解焙烧温度为350-600℃；进一步优选地，所述热解焙烧温度为350-500℃；再进一步优选地，所述热解焙烧温度为400-500℃。
优选地，所述热解焙烧时间为30-150min；进一步优选地，所述热解焙烧时间为40-120min。
优选地，所述热解焙烧是在含氧气体存在下进行。
优选地，所述含氧气体为空气或氧气。
优选地，所述回收电池黑粉的步骤为：将热解焙烧产物破碎筛分，得到电池黑粉。
优选地，所述破碎时间为30-150min；进一步优选地，所述破碎时间为40-120min。
优选地，所述破碎筛分后还得到了铜铝渣。
优选地，所述步骤S2中，浸取时间为50-200min；进一步优选地，所述步骤S2中，浸取时间为60-150min。
优选地，所述步骤S2中，浸取温度为30-90℃；进一步优选地，所述步骤S2中，浸取温度为40-80℃。
优选地，所述使用溶剂浸取电池黑粉中的金属离子步骤得到了浸出液和浸出渣。
优选地，所述溶剂与电池黑粉的固液比为2-10L/kg。
优选地，所述溶剂为水。
优选地，所述浸出渣转入湿法酸浸还原体系进行处理。
本发明的第二个方面在于提供上述从废旧锂离子电池中回收有价金属的方法在电池回收中的应用。
本发明的有益效果是：本发明中的方法使用硫酸盐溶液对废旧锂离子电池进行放电处理，使硫酸盐与废旧锂离子电池中的正极物质充分浸润接触，可以促进废旧锂离子电池中的有价金属的硫酸盐化，并提高热解焙烧效果。另外，本发明中的方法通过在热解焙烧过程中喷淋含硫酸盐的溶液，一方面可以通过水蒸发和硫酸盐的热解带走废旧锂离子电池热解产生的热量，实现控制热解焙烧温度，避免热解焙烧温度过高而导致产生大量的金属合金和杂质。
此外，本发明中的方法通过将沉锂产生的废液与硫酸混合后用于吸收热解焙烧过程中产生的含氨废气，实现了沉锂废液和热解废气的回收处理，同时将回收处理后的放电溶液和沉锂废液在热解焙烧过程中喷淋，一方面可以降低热解焙烧温度，实现精确控制热解焙烧炉温；另一方面可以为废旧锂离子电池热解补充硫酸盐成分，提高了废旧锂离子电池中有价金属的硫酸盐化，减少了杂质含量，使电池黑粉在水浸后的渣量大幅减少，可降低需要进行酸浸还原湿法回收的处理量，减少后续工艺处理流程的辅料用量，降低工艺成本。
本发明中的方法直接将整个废旧锂离子电池材料进行放电处理后热解焙烧回收有价金属，无需从废旧锂离子电池材料中分离出正极材料，自动化程度高，减少了人工成本，回收工艺简单，易操作，节约了工艺处理成本，同时该方法中产生的放电废液、沉锂废液、热解焙烧废气均可以回收使用，产生的废渣量少，后处理工艺简单，更符合绿色环保回收工艺的要求。在放电过程中，进入硫酸盐溶液中的废旧锂离子电池中的电解液等有害有机物质可以经过热解炉进行热分解，可以减少含电解液的放电废液的处理量。
本发明中的方法可以通过调节硫酸盐与废旧锂离子电池的质量比，结合炉温的调节，从而实现选择性提取镍钴，减少锰的硫酸盐化，进而实现选择性提取有价金属。
附图说明
图1为实施例3中方法所得电池黑粉的SEM图。
图2为实施例3中方法所得浸出渣的SEM图。
图3为实施例1-4中的从废旧锂离子电池中回收有价金属的方法的流程示意图。
具体实施方式
以下结合附图和实例对本发明的具体实施作进一步详细说明，但本发明的实施和保护不 限于此。需要指出的是，以下若有未特别详细说明之过程，均是本领域技术人员可参照现有技术实现或理解的。所用试剂或仪器未注明生产厂商者，视为可以通过市售购买得到的常规产品。
本发明实施例1-4和对比例1-2中的电池黑粉中的有价金属含量的测试方法为：
称取5g样品放入100mL烧杯中，加入适量去离子水，先使用自封袋包好的磁棒搅拌电池黑粉样品，来吸附具有磁性的镍钴锰单质和铁单质，再加入30mL的由浓盐酸和浓硝酸(浓盐酸和浓硝酸的体积比为3:1)组成的王水溶解磁性吸附物和剩余的电池黑粉，稀释后使用ICP电感耦合等离子体质谱仪测试各金属离子浓度。
本发明实施例1-4和对比例1-2中的浸出液中的主要金属离子浓度的测试方法为：稀释后使用ICP电感耦合等离子体质谱仪测试各金属离子浓度。
实施例1：
本例中的从废旧锂离子电池中回收有价金属的方法包括以下步骤：
(1)将1吨的废旧三元锂离子电池置于100g/L的硫酸氨的水溶液中进行放电处理，在硫酸氨溶液中放电4h，取出放电后的废旧三元锂离子电池，加入1000kg硫酸铵进行混合，在破碎机中混合破碎0.5h，得到固态的电池料；然后回收液态的放电溶液。
(2)将步骤(1)中破碎混合的电池料送入热解炉在350℃的空气气氛下进行热解焙烧，焙烧时间为120min，在焙烧过程中，将回收的液态的放电溶液喷淋到热解炉内，喷淋流量为0.5m3/h，收集焙烧产生的含氨废气；
(3)将步骤(2)中的热解焙烧后的物料在破碎机中破碎40min，筛分得到电池黑粉和铜铝渣；取少量电池黑粉，测量电池黑粉中的有价金属含量，测量结果如表1所示。
(4)将步骤(3)中所得电池黑粉与水以1kg：10L的固液比进行混合，在30℃下搅拌60min后固液分离，得到浸出渣和浸出液，将得到的320.5kg浸出渣转入湿法酸浸还原体系进行处理。取少量浸出液，测量浸出液中的主要离子浓度，结果如表2所示，根据废旧电池正极材料(占整体电池质量40％)的理论含量组成预估浸出液中的主要离子浸出率，浸出液中的主要离子预估浸出率如表3所示。在本实施例中，浸出液中的主要离子浓度及离子浸出率分别为：Li+浓度为3.56g/L，浸出率为95％；Ni2+浓度为10.35g/L，浸出率为72％；Co2+浓度为4.06g/L，浸出率为70％；Mn2+浓度为2.61g/L，浸出率为30％；
(5)先使用P204萃取剂对浸出液进行萃取，得到硫酸锰溶液，再使用P507萃取剂对除锰后的浸出液进行再次萃取，分别得到硫酸镍钴溶液和硫酸锂溶液；
(6)在硫酸锂溶液中加入6.5kg氧化钙调节pH到8，溶液温度为60℃；再加入12.3kg 的碳酸钠去除溶液中的钙杂质；
(7)对除杂后的硫酸锂溶液进行蒸发浓缩至Li+的浓度为10g/L，加入194.4kg碳酸钠在60℃下进行沉锂得到碳酸锂；回收沉锂废液；
将步骤(2)中的热解炉热解产生的含氨废气通入沉锂废液中，同时在沉锂废液中加入525L浓硫酸，用于吸收热解炉产生的含氨废气，再将沉锂废液以2m3/h流量喷淋进入热解炉，加大热解炉中硫酸铵的焙烧循环量。
在本实施例中，所有步骤中产生的废气废液均可以被有效回收再利用。如步骤(2)中的热解炉热解产生的含氨废气可以通入步骤(8)中的沉锂废液中，通过在沉锂废液中加入525L浓硫酸使其充分吸收含氨废气。吸收了含氨废气的沉锂废液可重新投入热解炉中(以2m3/h流量喷淋进入热解炉)，以促进前序步骤中热解炉中硫酸铵的焙烧循环量，提高金属回收效率。
实施例2：
本例中的从废旧锂离子电池中回收有价金属的方法，包括以下步骤：
(1)将1吨的废旧三元锂离子电池置于300g/L的硫酸氨的水溶液中进行放电处理，在硫酸铵溶液中放电3h，取出放电后的废旧三元锂离子电池，加入1000kg硫酸铵进行混合，在破碎机中混合破碎2h，得到固态的电池料；然后回收液态的放电溶液；
(2)将步骤(1)中破碎混合的电池料送入热解炉在400℃空气气氛下进行热解焙烧，焙烧时间为100min，在焙烧过程中，将回收的液态的放电溶液喷淋到热解炉内，喷淋流量为2m3/h，收集焙烧产生的含氨废气；
(3)将步骤(2)中的热解焙烧后的物料在破碎机中破碎100min，筛分得到电池黑粉和铜铝渣；取少量电池黑粉，测量电池黑粉中的有价金属含量，测量结果如表1所示。
(4)将步骤(3)中所得电池黑粉与水以1kg：10L的固液比进行混合，在50℃下搅拌120min后固液分离，得到浸出渣和浸出液，将得到的294.2kg浸出渣转入湿法酸浸还原体系进行处理。取少量浸出液，测量浸出液中的主要离子浓度，结果如表2所示，根据废旧电池正极材料(占整体电池质量40％)的理论含量组成预估浸出液中的主要离子浸出率，浸出液中的主要预估离子浸出率如表3所示，其中，在本实施例中，浸出液中的主要离子浓度及离子浸出率分别为：Li+浓度为3.62g/L，浸出率为98％；Ni2+浓度为12.22g/L，浸出率为85％；Co2+浓度为4.76g/L，浸出率为82％；Mn2+浓度为4.35g/L，浸出率为50％；
(5)使用P507萃取剂对浸出液进行萃取，分别得到硫酸镍钴锰溶液和硫酸锂溶液；
(6)在硫酸锂溶液中加入11.5kg氧化钙调节pH到8.5，溶液温度为70℃；再加入21.8kg的碳酸钠去除溶液中的钙杂质；
(7)对除杂后的硫酸锂溶液进行蒸发浓缩至Li+的浓度为12g/L，加入205.8kg碳酸钠在70℃下进行沉锂得到碳酸锂，回收沉锂废液；
将步骤(2)中的热解炉热解产生的含氨废气通入沉锂废液中，同时在沉锂废液中加入525L浓硫酸，用于吸收热解炉产生的含氨废气，再将沉锂废液以3m3/h流量喷淋进入热解炉，加大热解炉中硫酸铵的焙烧循环量。
在本实施例中，所有步骤中产生的废气废液均可以被有效回收再利用。如步骤(2)中的热解炉热解产生的含氨废气可以通入步骤(8)中的沉锂废液中，通过在沉锂废液中加入525L浓硫酸使其充分吸收含氨废气。吸收了含氨废气的沉锂废液可重新投入热解炉中(以3m3/h流量喷淋进入热解炉)，以促进前序步骤中热解炉中硫酸铵的焙烧循环量，提高金属回收效率。
实施例3：
本例中的从废旧锂离子电池中回收有价金属的方法，包括以下步骤：
(1)将1吨的废旧三元锂离子电池置于500g/L的硫酸氨的水溶液中进行放电处理，在硫酸铵溶液中放电2h，取出放电后的废旧三元锂离子电池，加入1000kg硫酸铵进行混合，在破碎机中混合破碎6h，得到固态的电池料；然后回收液态的放电溶液；
(2)将步骤(1)中破碎混合的电池料送入热解炉在500℃空气气氛下进行热解焙烧，焙烧时间为80min，在焙烧过程中，将回收的液态的放电溶液喷淋到热解炉内，喷淋流量为3.5m3/h，收集焙烧产生的含氨废气；
(3)将步骤(2)中的热解焙烧后物料在破碎机中破碎40min，筛分得到电池黑粉和铜铝渣；取少量电池黑粉，测量电池黑粉中的有价金属含量，测量结果如表1所示。
(4)将所得电池黑粉与水以1kg：10L的固液比进行混合，在90℃下搅拌150min后固液分离，得到浸出渣和浸出液，将得到的182.3kg浸出渣转入湿法酸浸还原体系进行处理。取少量浸出液，测量浸出液中的主要离子浓度，测量结果如表2所示，根据废旧电池正极材料(占整体电池质量40％)的理论含量组成预估浸出液中的主要离子浸出率，浸出液中的主要预估离子浸出率如表3所示。其中，本实施例中，浸出液中的主要离子浓度及离子浸出率分别为：Li+浓度为3.64g/L，浸出率为98％；Ni2+浓度为12.94g/L，浸出率为90％；Co2+浓度为5.30g/L，浸出率为91％；Mn2+浓度为6.96g/L，浸出率为85％；
(5)使用P507萃取剂对浸出液进行萃取，分别得到硫酸镍钴锰溶液和硫酸锂溶液；
(6)在硫酸锂溶液中加入16.25kg氧化钙调节pH到9，溶液温度为60℃；再加入30.75kg的碳酸钠去除溶液中的钙杂质；
(7)对除杂后的硫酸锂溶液进行蒸发浓缩至Li+的浓度为15g/L，加入208.5kg碳酸钠在 80℃下进行沉锂得到碳酸锂，回收沉锂废液；
将步骤(2)中的热解炉热解产生的含氨废气通入沉锂废液中，同时在沉锂废液中加入626L浓硫酸，用于吸收热解炉产生的含氨废气，再将沉锂废液以3.5m3/h流量喷淋进入热解炉，加大热解炉中硫酸铵的焙烧循环量。
在本实施例中，所有步骤中产生的废气废液均可以被有效回收再利用。如步骤(2)中的热解炉热解产生的含氨废气可以通入步骤(8)中的沉锂废液中，通过在沉锂废液中加入浓硫酸使其充分吸收含氨废气。吸收了含氨废气的沉锂废液可重新投入热解炉中(以3.5m3/h流量喷淋进入热解炉)，以促进前序步骤中热解炉中硫酸铵的焙烧循环量，提高金属回收效率。
分别测试步骤(3)中所得电池黑粉和步骤(4)中所得浸出渣的SEM图，其中，电池黑粉的SEM图如图1所示，浸出渣的SEM图如图2所示，通过对比图1和图2可知，浸出渣上存在大量金属离子溶出的溶洞。
实施例4：
本例中的从废旧锂离子电池中回收有价金属的方法，包括以下步骤：
(1)将1吨的废旧三元锂离子电池置于750g/L的硫酸氨的水溶液中进行放电处理，在硫酸铵溶液中放电2h，取出放电后的废旧三元锂离子电池，加入1200kg硫酸铵进行混合，在破碎机中混合破碎6h，得到固态的电池料；然后回收液态的放电溶液；
(2)将步骤(1)中破碎混合的电池料送入热解炉在600℃空气气氛下进行热解焙烧，焙烧时间为60min，在焙烧过程中，将回收的液态的放电溶液喷淋到热解炉内，喷淋流量为3.5m3/h，收集焙烧产生的含氨废气；
(3)将步骤(2)中的热解焙烧后物料在破碎机中破碎40min，筛分得到电池黑粉和铜铝渣；取少量电池黑粉，测量电池黑粉中的有价金属含量，测量结果如表1所示。
(4)将步骤(3)中的电池黑粉与水以1kg：10L的固液比进行混合，在90℃下搅拌150min后固液分离，得到浸出渣和浸出液，将得到的235.3kg浸出渣转入湿法酸浸还原体系进行处理。取少量浸出液，测量浸出液中的主要离子浓度，结果如表2所示，根据废旧电池正极材料(占整体电池质量40％)的理论含量组成预估浸出液中的主要离子浸出率，浸出液中的主要预估离子浸出率如表3所示。其中，本实施例中，浸出液中的主要离子浓度及离子浸出率分别为：Li+浓度为3.45g/L，浸出率为95％；Ni2+浓度为13.35g/L，浸出率为80％；Co2+浓度为5.30g/L，浸出率为73％；Mn2+浓度为6.84g/L，浸出率为70％；
(5)使用P507萃取剂对浸出液进行萃取，分别得到硫酸镍钴锰溶液和硫酸锂溶液；
(6)在硫酸锂溶液中加入16.25kg氧化钙调节pH到9，溶液温度为60℃；再加入30.75kg 的碳酸钠去除溶液中的钙杂质；
(7)对除杂后的硫酸锂溶液进行蒸发浓缩至Li+的浓度为15g/L，加入208.5kg碳酸钠在80℃下进行沉锂得到碳酸锂；
将步骤(2)中的热解炉热解产生的含氨废气通入沉锂废液中，同时在沉锂废液中加入626L浓硫酸，用于吸收热解炉产生的含氨废气，再将沉锂废液以3.5m3/h流量喷淋进入热解炉，加大热解炉中硫酸铵的焙烧循环量。
在本实施例中，所有步骤中产生的废气废液均可以被有效回收再利用。如步骤(2)中的热解炉热解产生的含氨废气可以通入步骤(8)中的沉锂废液中，通过在沉锂废液中加入浓硫酸使其充分吸收含氨废气。吸收了含氨废气的沉锂废液可重新投入热解炉中(以3.5m3/h流量喷淋进入热解炉)，以促进前序步骤中热解炉中硫酸铵的焙烧循环量，提高金属回收效率。
实施例1-4中的从废旧锂离子电池中回收有价金属的方法的流程示意图如图3所示。
对比例1：
本例中的从废旧锂离子电池中回收有价金属的方法，包括以下步骤：
(1)将1吨的废旧三元锂离子电池置于100g/L的氯化钠溶液中进行放电处理，在盐水中放电4h，放电后的废旧三元锂离子电池取出后在破碎机中混合破碎0.5h，得到电池料；
(2)将步骤(1)中破碎混合的电池料送入热解炉在空气气氛下进行热解，焙烧时间为120min；因热解过程中无相应温控措施，炉温上升至750℃以上。
(3)将步骤(2)中热解后的物料在破碎机中破碎40min，筛分得到电池黑粉和铜铝渣；测量电池黑粉中的有价金属含量，并将测量结果记录在表1中。
(4)将步骤(3)中所得电池黑粉与水以1kg：10L的固液比进行混合，在30℃下搅拌60min后固液分离，得到浸出渣和浸出液，将得到的786.5kg浸出渣转入湿法酸浸还原体系进行处理。测量浸出液中的主要离子浓度，并将浸出液中的主要离子浓度记录在下表2中，然后根据废旧电池正极材料(占整体电池质量40％)的理论含量组成预估浸出液中的主要离子浸出率，浸出液中的主要离子浸出率数据记录在下表3中，浸出液中的主要离子浓度及离子浸出率分别为：Li+浓度为0.18g/L，浸出率为4.8％；Ni2+浓度为0.33g/L，浸出率为2.3％；Co2+浓度为0.05g/L，浸出率为0.8％；Mn2+浓度为0.01g/L，浸出率为0.2％。
对比例2：
本例中的从废旧锂离子电池中回收有价金属的方法，包括以下步骤：
(1)将1吨的废旧三元锂离子电池置于100g/L的硫酸氨的水溶液中进行放电处理，在硫酸铵溶液中放电4h，放电后的废旧三元锂离子电池与1000kg硫酸铵进行混合，在破碎机 中混合破碎0.5h；
(2)将步骤(1)中破碎混合的电池料送入热解炉在500℃的空气气氛下进行热解焙烧，焙烧时间为120min；不喷淋放电硫酸铵溶液进行温控，热解炉温会达到700℃。
(3)将步骤(2)中的热解焙烧后物料在破碎机中破碎60min，筛分得到电池黑粉和铜铝渣；测量电池黑粉中的有价金属含量，并将测量结果记录在表1中。
(4)将所得电池黑粉与水以1kg：10L的固液比进行混合，在30℃下搅拌60min后固液分离，得到浸出渣和浸出液，将得到的225.6kg浸出渣转入湿法酸浸还原体系进行处理。测量浸出液中的主要离子浓度，并将浸出液中的主要离子浓度记录在下表2中，然后根据废旧电池正极材料(占整体电池质量40％)的理论含量组成预估浸出液中的主要离子浸出率，浸出液中的主要离子浸出率数据记录在下表3中，浸出液中的主要离子浓度及离子浸出率分别为：Li+浓度为3.37g/L，浸出率为90％；Ni2+浓度为10.11g/L，浸出率为71％；Co2+浓度为3.86g/L，浸出率为68％；Mn2+浓度为5.39g/L，浸出率为62％；
(5)使用P507萃取剂对浸出液进行萃取，分别得到硫酸镍钴锰溶液和硫酸锂溶液；
(6)在硫酸锂溶液中加入5.9kg氧化钙调节pH到8，溶液温度为60℃；再加入11.8kg的碳酸钠去除溶液中的钙杂质。
表1：实施例1-4和对比例1-2中的电池黑粉中有价金属含量(质量分数％)
由上表1可知，与对比例1相比，实施例1-4中方法所获得的电池黑粉中的Ni、Co和Mn离子的含量显著提高，而Al、Cu和Fe离子的含量显著降低，表明将废旧三元锂离子电池在硫酸铵溶液中进行放电，使硫酸盐充分浸润废旧三元锂离子电池材料，同时结合在热解炉中喷淋硫酸盐溶液实现对热解炉炉温的控制，可以实现对废旧三元锂离子电池中镍钴锰元素的选择性硫酸盐化，实现选择性浸出Ni、Co和Mn离子。此外，与对比例1-2相比，实施例1-4中的方法通过在热解炉内喷淋硫酸盐溶液，可以实现对热解炉炉温的控制，降低了电池黑粉中Ni、Co和Mn单质的含量，避免因热解炉温度过高而使废旧三元锂离子电池材料的 Ni、Co和Mn元素生成较多的合金料。与实施例1-3相比，实施例4采用的热解焙烧温度600℃，在此温度下，部分正极材料因电池中石墨的作用产生部分合金单质，因此，实施例4中的镍、钴和锰单质的含量均明显高于实施例1-3，进一步表明采用350-500℃的煅烧温度可以降低电池黑粉中镍钴锰单质和合金的含量。
表2：实施例1-4和对比例1-2中的浸出液中主要离子浓度
由上表2可知，与对比例1相比，实施例1-4中方法通过将废旧三元锂离子电池在硫酸铵溶液中进行放电，使硫酸盐充分浸润废旧三元锂离子电池材料，同时结合在热解炉中喷淋硫酸盐溶液实现对热解炉炉温的控制，可以显著提高Li、Ni、Co、Mn元素的硫酸盐化，提高Li、Ni、Co、Mn的回收率。与对比例2相比，实施例1-4中的方法通过在热解炉中喷淋硫酸盐溶液用以热解炉炉温，可以进一步提高Li、Ni、Co、Mn元素的硫酸盐化。
表3：实施例1-4和对比例1-2中浸出液中离子浸出率和浸出渣的质量
由上表3可知，与对比例1相比，实施例1-4中方法通过将废旧三元锂离子电池在硫酸铵溶液中进行放电，使硫酸盐充分浸润废旧三元锂离子电池材料，同时结合在热解炉中喷淋硫酸盐溶液实现对热解炉炉温的控制，可以显著提高Li、Ni、Co、Mn离子的浸出率，提高Li、Ni、Co、Mn的回收率，同时大幅度的降低了浸出渣的含量，降低了浸出渣处理量和浸 出渣处理成本。与对比例2相比，实施例1-4中的方法通过在热解炉中喷淋硫酸盐溶液用以热解炉炉温，可以进一步提高Li、Ni、Co、Mn离子的浸出率。
由上表2-3可知，与实施例1-2和实施例4相比，实施例3通过喷淋硫酸铵溶液将热解炉温度控制在500℃，硫酸盐与废旧三元锂离子电池材料的质量比控制在1：1，可以进一步提高对镍钴锰元素的选择性硫酸盐转化。与实施例2-4相比，实施例1通过喷淋硫酸铵溶液将热解炉温度控制在350℃，硫酸盐与废旧三元锂离子电池材料的质量比控制在1：1，可以使Mn转化的硫酸盐减少，主要选择性浸出镍钴离子。
上面对本发明实施例作了详细说明，但是本发明不限于上述实施例，在所属技术领域普通技术人员所具备的知识范围内，还可以在不脱离本发明宗旨的前提下做出各种变化。此外，在不冲突的情况下，本发明的实施例及实施例中的特征可以相互组合。

Claims (10)

  1. 一种从废旧锂离子电池中回收有价金属的方法，其特征在于：包括以下步骤：
    S1：将废旧锂离子电池在含硫酸盐的放电溶液中放电，然后与硫酸盐和/或硫化物混合并热解焙烧，在热解焙烧过程中喷淋含硫酸盐的溶液，回收电池黑粉；
    S2：使用溶剂浸取电池黑粉中的金属离子，萃取沉锂得到碳酸锂和硫酸镍钴锰。
  2. 根据权利要求1所述的从废旧锂离子电池中回收有价金属的方法，其特征在于：所述萃取沉锂得到碳酸锂和硫酸镍钴锰的步骤为：将电池黑粉使用溶剂浸取后的浸出液萃取，得到硫酸镍钴锰溶液和硫酸锂溶液，然后对硫酸锂溶液除杂后浓缩，再沉锂，得到碳酸锂和硫酸镍钴锰。
  3. 根据权利要求2所述的从废旧锂离子电池中回收有价金属的方法，其特征在于：所述除杂步骤为：在硫酸锂溶液中加入氧化钙，调节pH为8-9，温度为60-90℃，然后再加入碳酸盐并过滤。
  4. 根据权利要求3所述的从废旧锂离子电池中回收有价金属的方法，其特征在于：所述碳酸盐和氧化钙的物质的量比为1:1。
  5. 根据权利要求1所述的从废旧锂离子电池中回收有价金属的方法，其特征在于：所述步骤S1中的含硫酸盐的溶液为：将热解焙烧步骤产生的废气通入沉锂步骤产生的废液和硫酸的混合液中得到的含硫酸盐的溶液；或，回收所述放电步骤产生的废液得到的含硫酸盐的溶液。
  6. 根据权利要求5所述的从废旧锂离子电池中回收有价金属的方法，其特征在于：所述硫酸的用量与步骤S1中硫酸盐和/或硫化物的用量的物质的量比为(0.5-2.5)：1。
  7. 根据权利要求1或5所述的从废旧锂离子电池中回收有价金属的方法，其特征在于：所述含硫酸盐的溶液的喷淋流量为0.5-3.5m3/h。
  8. 根据权利要求1所述的从废旧锂离子电池中回收有价金属的方法，其特征在于：所述硫酸盐和/或硫化物中的硫酸根与废旧锂离子电池的质量比为(0.6-1.2)：1。
  9. 根据权利要求1所述的从废旧锂离子电池中回收有价金属的方法，其特征在于：所述热解焙烧温度为350-600℃。
  10. 权利要求1-9任一项所述的从废旧锂离子电池中回收有价金属的方法在电池回收中的应用。
PCT/CN2023/079351 2022-08-25 2023-03-02 一种从废旧锂离子电池中回收有价金属的方法 WO2024040910A1 (zh)

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Citations (6)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN108767354A (zh) * 2018-05-29 2018-11-06 中南大学 一种从废旧锂离子电池正极材料中回收有价金属的方法
CN109935922A (zh) * 2019-03-14 2019-06-25 北京矿冶科技集团有限公司 一种从废旧锂离子电池材料中回收有价金属的方法
CN111206148A (zh) * 2020-03-16 2020-05-29 宁波容百新能源科技股份有限公司 一种利用废旧三元锂电池回收制备三元正极材料的方法
CN114293029A (zh) * 2021-12-30 2022-04-08 上海电力大学 一种从废旧锂离子电池中选择性提锂的方法
KR20220101972A (ko) * 2021-01-12 2022-07-19 한국해양대학교 산학협력단 폐리튬이차전지 황산염 배소 산물의 수침출에 따른 유용금속 회수 방법
CN115466845A (zh) * 2022-08-25 2022-12-13 广东邦普循环科技有限公司 一种从废旧锂离子电池中回收有价金属的方法

Patent Citations (6)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN108767354A (zh) * 2018-05-29 2018-11-06 中南大学 一种从废旧锂离子电池正极材料中回收有价金属的方法
CN109935922A (zh) * 2019-03-14 2019-06-25 北京矿冶科技集团有限公司 一种从废旧锂离子电池材料中回收有价金属的方法
CN111206148A (zh) * 2020-03-16 2020-05-29 宁波容百新能源科技股份有限公司 一种利用废旧三元锂电池回收制备三元正极材料的方法
KR20220101972A (ko) * 2021-01-12 2022-07-19 한국해양대학교 산학협력단 폐리튬이차전지 황산염 배소 산물의 수침출에 따른 유용금속 회수 방법
CN114293029A (zh) * 2021-12-30 2022-04-08 上海电力大学 一种从废旧锂离子电池中选择性提锂的方法
CN115466845A (zh) * 2022-08-25 2022-12-13 广东邦普循环科技有限公司 一种从废旧锂离子电池中回收有价金属的方法

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