Step-by-step recycling method for decommissioned lithium batteries

Abstract

A step-by-step recycling method for decommissioned lithium batteries. The method comprises the following steps: performing pre-processing on a decommissioned lithium battery, to obtain a powder; mixing the powder and a first organic acid, and performing a leaching reaction, to obtain a leachate containing lithium and aluminum, and a first leaching residue; mixing the first leaching residue and a composite ammonia source, and performing a leaching reaction, to obtain a copper-containing leachate and a second leaching residue; performing acid leaching on the second leaching residue, to obtain a solution containing nickel, cobalt, manganese and iron, and a third leaching residue; and mixing the solution containing nickel, cobalt, manganese and iron with a pH value regulator, performing a reaction, and aging, to obtain a solution containing nickel, cobalt, and manganese, and an iron precipitation residue. The method reduces the loss of valuable metal elements lithium, nickel, cobalt and manganese, and the output of slag; increases the recovery rate of metallic lithium; implements high-value utilization of copper and aluminum in the leaching solution; and reduces process production costs. In addition, reclamation and reduction of solid waste in a lithium battery wet recycling process is achieved.

Description

A method for recycling retired lithium batteries step by step Technical Field

[0001] The present disclosure belongs to the technical field of lithium-ion battery recycling, and particularly relates to a method for step-by-step recycling of retired lithium batteries. Background Art

[0002] With the rapid development of the new energy industry, the full life cycle recycling industry of retired lithium batteries is booming. In the existing lithium battery recycling technology, retired lithium batteries are first fully discharged by physical and chemical methods, and then directly disassembled, calcined, and screened to separate battery powder, electrodes (aluminum foil, copper foil), diaphragms and other components. Then, a series of processes such as acid leaching, impurity removal, and extraction are carried out to obtain relatively pure ternary slurry and lithium-containing liquid. Since the raw materials are mixed with a large amount of copper, aluminum, iron and other impurities, the impurity removal efficiency during the acid leaching liquid removal process is low and the loss of precious metal elements lithium, nickel, cobalt, and manganese is large, and a large amount of low-value sponge copper is produced. At the same time, a large amount of low-grade iron-aluminum slag and graphite slag are produced, which increases production and manufacturing costs.

[0003] To this end, the prior art has proposed some solutions. For example, CN106191466A discloses a method for recovering lithium from waste lithium iron phosphate batteries, comprising disassembling and shelling the waste lithium iron phosphate batteries to obtain a battery core; calcining the battery core; crushing the battery core and sieving it to obtain a positive and negative electrode mixture; adding alkaline solution to the positive and negative electrode mixture to remove residual aluminum in the positive and negative electrode mixture, and filtering to obtain filter mud; leaching the filter mud with a lithium solution using a strong acid, filtering to remove carbon and iron phosphate, and obtaining a leachate; adjusting the pH of the leachate to precipitate a small amount of iron, and filtering to obtain a filtrate; removing copper impurities in the filtrate, and filtering to obtain a filtrate; and adding solid sodium carbonate to the filtrate to obtain a lithium carbonate precipitate. CN113285135A discloses a method for recycling multiple components of waste lithium iron phosphate batteries, comprising the following steps: breaking the shells of waste lithium iron phosphate batteries after discharge treatment, disassembling and separating them; treating the battery cores to obtain a solvent recovery liquid; crushing and sorting the battery cores to obtain coarse lithium iron phosphate powder, copper powder, and aluminum powder; adding the coarse lithium iron phosphate powder to an acid solution for reaction, filtering to obtain an acid leaching solution and carbon residue, and washing and drying the carbon residue to obtain high-carbon graphite; adjusting the pH value of the acid leaching solution, adding a reducing agent to remove copper, and filtering to obtain a copper removal solution and copper residue; adding an oxidant and an appropriate amount of a phosphorus source to the copper removal solution to obtain ferric orthophosphate; adding an alkaline solution to the iron precipitation solution to obtain an aluminum removal solution and aluminum residue; adding the aluminum precipitation solution to the alkaline solution to obtain an alkalized solution and an alkaline residue; evaporating and concentrating the alkalized solution to obtain a lithium-rich solution, which is added to a sodium carbonate solution to obtain lithium carbonate.

[0004] However, the loss of valuable metal elements in the above solutions is still significant, and the process production cost is high.

[0005] Therefore, how to effectively reduce the loss of valuable metal elements during the recycling process and reduce process production costs is an urgent problem that needs to be solved.

[0006] Summary of the Invention

[0007] The following is a summary of the subject matter described in detail herein. This summary is not intended to limit the scope of the claims.

[0008] The present disclosure aims to provide a method for the step-by-step recycling of retired lithium batteries. This method reduces the loss of valuable metal elements such as lithium, nickel, cobalt, and manganese, as well as the amount of slag produced. It also improves the recovery rate of metallic lithium and the high-value utilization of copper and aluminum in the leachate, reducing process production costs and achieving resource utilization and reduction of solid waste during the wet recycling of lithium batteries.

[0009] To achieve this disclosure purpose, the present disclosure adopts the following technical solutions:

[0010] In a first aspect, the present disclosure provides a method for recycling retired lithium batteries step by step, the method comprising the following steps:

[0011] (1) Pre-treating retired lithium batteries to obtain powder;

[0012] (2) mixing the powder and a first organic acid to perform a leaching reaction to obtain a lithium- and aluminum-containing leachate and a first leaching residue;

[0013] (3) mixing the first leaching residue and a composite ammonia source to perform a leaching reaction to obtain a copper-containing leachate and a second leaching residue;

[0014] (4) acid leaching the second leaching residue to obtain a solution containing nickel, cobalt, manganese, and iron and a third leaching residue;

[0015] (5) mixing the nickel, cobalt, manganese and iron-containing solution and a pH regulator, reacting and aging to obtain the nickel, cobalt and manganese-containing solution and iron precipitate slag.

[0016] The method provided by the present disclosure reduces the loss of valuable metal elements lithium, nickel, cobalt, and manganese and the output of slag, while improving the recovery rate of metallic lithium and the high-value utilization of copper and aluminum in the leachate, reducing process production costs, and realizing the resource utilization and reduction of solid waste in the wet recovery process of lithium batteries.

[0017] In the present disclosure, the powder and the first organic acid are mixed and subjected to a leaching reaction, which can preferentially extract more than 98% of the lithium element and dissolve 100% of the aluminum. This not only reduces the subsequent loss of lithium metal, but also prevents the introduction of aluminum in the subsequent leaching process, thereby achieving iron-aluminum separation. In addition, the leaching rate of nickel and cobalt in this process is <0.5%, and the leaching rate of manganese is <1.5%.

[0018] In the present disclosure, the first leaching residue and the composite ammonia source are mixed to carry out a leaching reaction to achieve selective copper removal. The copper leaching rate can reach 95.2%. The obtained copper-containing leachate can be used to prepare micro-nano copper with regular shapes. Compared with the existing technology, this method not only efficiently separates copper foil from lithium battery powder, but also can prepare higher-value micro-nano copper, thereby improving the utilization value of copper foil in lithium batteries.

[0019] In the present disclosure, acid leaching is performed on the second leaching residue, which can reduce high-valent elements such as nickel, cobalt, and manganese in the battery powder into low-valent metal elements, thereby improving the leaching rate of Ni, Co, and Mn.

[0020] In this disclosure, a solution containing nickel, cobalt, manganese, and iron is mixed with a pH adjuster to preferentially and selectively remove 99.58% of the iron in the leachate. During the entire selective iron removal process, the total loss of Ni, Co, and Mn is only 1.1%. Compared with existing technologies, this method significantly reduces the loss of valuable metal elements, while simultaneously achieving high-value resource utilization of lithium battery aluminum and copper foil and reducing waste residue (ferro-aluminum slag and graphite slag), achieving both good economic and environmental benefits.

[0021] As an optional technical solution of the present disclosure, the particle size D50 of the powder in step (1) is 48-78 μm, for example, it can be 48 μm, 53 μm, 58 μm, 63 μm, 68 μm, 73 μm or 78 μm, etc.

[0022] In one embodiment, the pre-processing step comprises:

[0023] Discharge, crush, calcine and screen retired lithium batteries.

[0024] In one embodiment, the specific steps of discharging include:

[0025] Soak retired lithium batteries in salt water to discharge them.

[0026] As an optional technical solution of the present disclosure, in step (2), the first organic acid includes any one of oxalic acid, tartaric acid, malic acid or acetic acid, or a combination of at least two of them.

[0027] It should be noted that the product obtained by mixing oxalic acid and powder includes single lithium oxalate and aluminum complex oxalate.

[0028] In one embodiment, the concentration of the first organic acid in step (2) is 0.5-0.6 mol / L, for example, 0.5 mol / L, 0.52 mol / L, 0.54 mol / L, 0.56 mol / L, 0.58 mol / L or 0.6 mol / L, etc., and can be optionally 0.5-0.55 mol / L.

[0029] In one embodiment, the mass volume ratio of the powder in step (2) and the first organic acid is 1g:(6-10)mL, for example, it can be 1g:6mL, 1g:7mL, 1g:8mL, 1g:9mL or 1g:10mL, etc., and can be optionally 1g:(7-8)mL.

[0030] In one embodiment, the temperature of the leaching reaction in step (2) is 50-80°C, for example, 50°C, 55°C, 60°C, 65°C, 70°C, 75°C or 80°C, and can be optionally 55-65°C.

[0031] In one embodiment, the leaching reaction time in step (2) is 40-80 min, for example, 40 min, 50 min, 60 min, 70 min or 80 min, etc., and can be optionally 50-70 min.

[0032] In one embodiment, after the leaching reaction in step (2) is completed, the obtained filter residue is washed with water and filtered to obtain a first leaching residue and a filtrate, and the filtrate is mixed with the reaction solution obtained by the leaching reaction to obtain a lithium- and aluminum-containing leachate.

[0033] In the present disclosure, the purpose of washing and filtering the filter residue again is to clean the metal ions attached to the surface of the residue and reduce the loss of metal ions. The same applies to the following.

[0034] As an optional technical solution of the present disclosure, the lithium-aluminum-containing leachate of step (2) is subjected to aluminum precipitation treatment, and the specific steps include:

[0035] The lithium-containing and aluminum-containing leachate is mixed with a second organic acid, subjected to a complex reaction, and aged to obtain a lithium-containing leachate and aluminum salt.

[0036] In the present disclosure, a lithium- and aluminum-containing leachate is mixed with a second organic acid and subjected to a complexation reaction, which can efficiently and selectively remove 97.08% of the aluminum in the leachate without causing lithium loss, and the obtained aluminum salt can be decomposed into a high-purity Al(PO3)3 product at 1000°C, thereby obtaining a high-grade aluminum concentrate.

[0037] In one embodiment, the second organic acid includes any one or a combination of at least two of phytic acid, diethyl phosphate, alkyl phosphate, or phenyl hypophosphite.

[0038] In one embodiment, the molar ratio of Al ions to phytic acid in the lithium- and aluminum-containing leachate is (2-4):1, for example, 2:1, 2.5:1, 3:1, 3.5:1 or 4:1, etc., and can be optionally (3-3.5):1.

[0039] In one embodiment, the temperature of the complexation reaction is 50-70°C, for example, 50°C, 55°C, 60°C, 65°C or 70°C, and can be 55-65°C.

[0040] In one embodiment, the complexation reaction time is 20-50 min, for example, 20 min, 25 min, 30 min, 35 min, 40 min, 45 min or 50 min.

[0041] In one embodiment, the aging time is 30-60 min, for example, 30 min, 40 min, 50 min or 60 min, etc., and can be 50-60 min.

[0042] In one embodiment, the aluminum salt includes any one or a combination of at least two of aluminum phytate, diethyl aluminum phosphate, alkyl aluminum phosphate, or phenyl aluminum hypophosphite.

[0043] In one embodiment, the aluminum salt is heat-treated to obtain aluminum metaphosphate.

[0044] In one embodiment, the heat treatment temperature is 800-1000°C, for example, 800°C, 850°C, 900°C, 950°C or 1000°C, and can be optionally 900-1000°C.

[0045] In one embodiment, the heat treatment time is 60-90 min, for example, it can be 60 min, 65 min, 70 min, 75 min, 80 min, 85 min or 90 min, etc., and can be optionally 65-75 min.

[0046] As an optional technical solution of the present disclosure, the composite ammonia source in step (3) includes ammonia water and ammonium chloride solution.

[0047] In one embodiment, the concentration of the ammonia water is 4-5 mol / L, for example, 4 mol / L, 4.2 mol / L, 4.4 mol / L, 4.6 mol / L, 4.8 mol / L or 5 mol / L, etc., and can be optionally 4.5-5 mol / L.

[0048] In one embodiment, the concentration of the ammonium chloride solution is 5-6.5 mol / L, for example, 5 mol / L, 5.5 mol / L, 6 mol / L or 6.5 mol / L, etc., and can be optionally 5.5-6 mol / L.

[0049] In one embodiment, the volume ratio of the ammonia water and the ammonium chloride solution is (15-25): (5-15), wherein the selection range of the ammonia water "15-25" can be, for example, 15, 20 or 25, and the selection range of the ammonium chloride solution "5-15" can be, for example, 5, 10 or 15, etc.

[0050] In one embodiment, the mass volume ratio of the first leaching residue and the composite ammonia source in step (3) is 1g:(5-10)mL, for example, it can be 1g:5mL, 1g:6mL, 1g:7mL, 1g:8mL, 1g:9mL or 1g:10mL, etc., and can be optionally 1g:(5-7)mL.

[0051] In one embodiment, the temperature of the leaching reaction in step (3) is 40-60°C, for example, 40°C, 45°C, 50°C, 55°C or 60°C, etc., and can be optionally 45-55°C.

[0052] In one embodiment, the leaching reaction time in step (3) is 40-80 min, for example, 40 min, 50 min, 60 min, 70 min or 80 min, etc., and can be optionally 55-65 min.

[0053] In one embodiment, after the leaching reaction in step (3) is completed, the obtained filter residue is washed with water and filtered to obtain a second leaching residue and a filtrate, and the filtrate is mixed with the reaction solution after the leaching reaction to obtain a copper-containing leachate.

[0054] As an optional technical solution of the present disclosure, the copper-containing leaching solution in step (3) is mixed with a reducing agent and subjected to a reduction reaction to obtain pure nano-copper.

[0055] In one embodiment, the reducing agent includes any one of glucose, ascorbic acid or hydrazine hydrate, or a combination of at least two thereof.

[0056] In the present disclosure, glucose can reduce copper ions in the solution to cuprous oxide.

[0057] In one embodiment, a protective solution is further added during the reduction reaction, and the protective solution includes hexadecyltrimethylammonium bromide solution.

[0058] It should be noted that the huge surface energy of copper nanoparticles not only causes them to spontaneously agglomerate, but also makes their chemical properties very active and very easy to be oxidized. The nitrogen atoms and oxygen atoms with lone pairs of electrons on the side chains of the molecular structure of hexadecyltrimethylammonium bromide can coordinate with the surface atoms of copper nanoparticles to form a protective layer, which prevents the copper nanoparticles from agglomerating while protecting them from oxidation.

[0059] In one embodiment, after the reduction reaction, sodium hydroxide and ascorbic acid are added to react.

[0060] It should be noted that the addition of ascorbic acid will directly affect the nucleation and growth process of the copper nanoparticles, thereby affecting the morphology of the copper nanoparticles.

[0061] In one embodiment, the temperature of the reduction reaction is 75-95°C, for example, 75°C, 80°C, 85°C, 90°C or 95°C.

[0062] In one embodiment, the reduction reaction time is 5-15 min, for example, 5 min, 10 min or 15 min.

[0063] In one embodiment, the particle size D50 of the pure nano-copper is 450-550 nm, for example, 450 nm, 475 nm, 500 nm, 525 nm or 550 nm.

[0064] As an optional technical solution of the present disclosure, the specific steps of the acid leaching in step (4) include:

[0065] The second leaching residue is mixed with sulfuric acid solution to carry out acid leaching reaction.

[0066] In one embodiment, the concentration of the sulfuric acid solution is 2-4 mol / L, for example, 2 mol / L, 2.5 mol / L, 3 mol / L, 3.5 mol / L or 4 mol / L, etc., and can be optionally 2.5-3.5 mol / L.

[0067] In one embodiment, after the second leaching residue and the sulfuric acid solution are mixed, a reducing gas is introduced into the mixed solution.

[0068] In the present disclosure, the introduction of reducing gas can provide a reducing atmosphere and accelerate the leaching rate of metal elements.

[0069] In one embodiment, the reducing gas includes any one of sulfur dioxide, sulfur monoxide, or hydrogen sulfide, or a combination of at least two thereof.

[0070] In the present disclosure, the high-temperature exhaust gas SO2 is put to industrial use to achieve the goal of treating waste with waste.

[0071] In one embodiment, the flow rate of the reducing gas is 1-2.5 L / min, for example, 1 L / min, 1.5 L / min, 2 L / min or 2.5 L / min, etc., and can be optionally 1.5-2 L / min.

[0072] In one embodiment, the mass volume ratio of the second leaching residue to the sulfuric acid solution is 1 g:(3-5) mL, for example, 1 g:3 mL, 1 g:3.5 mL, 1 g:4 mL, 1 g:4.5 mL or 1 g:5 mL, etc., and can be optionally 1 g:(4-5) mL.

[0073] In one embodiment, the temperature of the acid leaching reaction is 75-95°C, for example, 75°C, 80°C, 85°C, 90°C or 95°C.

[0074] In one embodiment, the acid leaching reaction time is 50-70 min, for example, it can be 50 min, 55 min, 60 min, 65 min or 70 min.

[0075] In one embodiment, after the acid leaching in step (4) is completed, the obtained filter residue is washed with water and filtered to obtain a filtrate and a third leaching residue, and the filtrate is mixed with the reaction solution after acid leaching to obtain a solution containing nickel, cobalt, manganese and iron.

[0076] In one embodiment, the third leached residue is graphite residue.

[0077] As an optional technical solution of the present disclosure, the pH regulator in step (5) includes any one of sodium bicarbonate, sodium carbonate, calcium carbonate, ammonium carbonate or ammonium bicarbonate, or a combination of at least two thereof.

[0078] In one embodiment, after the nickel, cobalt, manganese, and iron-containing solution and the pH regulator are mixed in step (5), the pH value of the mixed solution is 3-3.5, for example, it can be 3, 3.1, 3.2, 3.3, 3.4 or 3.5, and can be optionally 3.2-3.3.

[0079] In one embodiment, the reaction temperature in step (5) is 75-95°C, for example, 75°C, 80°C, 85°C, 90°C or 95°C, and can be 80-90°C.

[0080] In one embodiment, the reaction time of step (5) is 40-60 min, for example, 40 min, 45 min, 50 min, 55 min or 60 min, etc., and can be optionally 45-55 min.

[0081] In one embodiment, the aging time in step (5) is 180-300 min, for example, it can be 180 min, 200 min, 220 min, 240 min, 260 min, 280 min or 300 min, etc., and can be optionally 220-260 min.

[0082] In one embodiment, after the aging in step (5), the obtained filter residue is washed with water and filtered to obtain iron precipitate and filtrate, and the filtrate is mixed with the aging liquid to obtain a solution containing nickel, cobalt and manganese.

[0083] As an optional technical solution disclosed in the present invention, the iron precipitate slag is post-processed to obtain iron concentrate.

[0084] In one embodiment, the post-processing step comprises:

[0085] The iron precipitate slag is mixed with a carbon source and calcined to obtain iron concentrate.

[0086] In one embodiment, the carbon source comprises graphite slag.

[0087] In one embodiment, the mass ratio of the iron precipitate slag to the carbon source is 1:(0.1-0.2), for example, it can be 1:0.1, 1:0.12, 1:0.14, 1:0.16, 1:0.18 or 1:0.2, etc., and can be optionally 1:(0.13-0.17).

[0088] In one embodiment, the calcination temperature is 1000-1200°C, for example, 1000°C, 1050°C, 1100°C, 1150°C or 1200°C, and can be optionally 1050-1150°C.

[0089] In one embodiment, the calcination time is 60-90 min, for example, 60 min, 65 min, 70 min, 75 min, 80 min, 85 min or 90 min, etc., and can be optionally 70-80 min.

[0090] As an optional technical solution of the present disclosure, the method includes the following steps:

[0091] (1) Pre-processing of retired lithium batteries, specifically including the following steps:

[0092] Discharging, crushing, calcining and sieving retired lithium batteries to obtain powder with a particle size D50 of 48-78 μm;

[0093] (2) mixing the powder with oxalic acid having a concentration of 0.5-0.6 mol / L at a ratio of 1 g: (6-10) mL, stirring and reacting at a temperature of 50-80° C. for 40-80 min, filtering to obtain a filter residue and a reaction liquid, washing the obtained filter residue with water, filtering to obtain a first leaching residue and a filtrate, and mixing the filtrate and the reaction liquid to obtain a lithium- and aluminum-containing leachate;

[0094] (3) mixing the lithium-containing and aluminum-containing leachate with phytic acid, performing a complexation reaction at 50-70° C. for 20-50 minutes, aging for 30-60 minutes, filtering to obtain the lithium-containing leachate and aluminum-containing phytate, drying the aluminum-containing phytate, and then performing a heat treatment at 800-1000° C. for 60-90 minutes to obtain aluminum metaphosphate;

[0095] The molar ratio of Al ions to phytic acid in the lithium and aluminum-containing leachate is (2-4):1;

[0096] (4) mixing the first leaching residue and the composite ammonia source according to 1 g: (5-10) mL, performing a leaching and stirring reaction at 40-60 ° C for 40-80 min, filtering to obtain a filter residue and a reaction solution, washing the obtained filter residue with water, filtering to obtain a second leaching residue and a filtrate, mixing the filtrate and the reaction solution to obtain a copper-containing leaching solution; mixing the copper-containing leaching solution with a hexadecyltrimethylammonium bromide solution, adding a reducing agent and mixing, performing a reduction reaction to obtain a first mixed solution, mixing the first mixed solution with a sodium hydroxide and ascorbic acid solution, reacting to obtain pure nano-copper with a particle size D50 of 450-550 nm;

[0097] The composite ammonia source includes aqueous ammonia with a concentration of 4-5 mol / L and ammonium chloride solution with a concentration of 5-6.5 mol / L, and the volume ratio of aqueous ammonia to ammonium chloride solution is (15-25):(5-15);

[0098] (5) mixing the second leaching residue with a sulfuric acid solution having a concentration of 2-4 mol / L, and introducing sulfur dioxide gas at a flow rate of 1-2.5 L / min, performing an acid leaching reaction at 75-95° C. for 50-70 min, filtering to obtain a filter residue and a reaction liquid, washing the obtained filter residue with water, filtering to obtain a filtrate and graphite residue, and mixing the filtrate and the reaction liquid to obtain a solution containing nickel, cobalt, manganese, and iron;

[0099] Wherein, the mass volume ratio of the second leaching residue and the sulfuric acid solution is 1g:(3-5)mL;

[0100] (6) mixing the nickel, cobalt, manganese, and iron-containing solution with sodium bicarbonate to a pH value of 3-3.5, then reacting at 75-95° C. for 40-60 min, aging for 180-300 min, and filtering to obtain a filter residue and an aging liquid, washing the obtained filter residue with water, filtering to obtain an iron precipitate and a filtrate, and mixing the filtrate with the aging liquid to obtain a nickel, cobalt, and manganese-containing solution;

[0101] The iron precipitate slag and graphite slag are mixed and calcined at 1000-1200° C. for 60-90 minutes to obtain iron concentrate;

[0102] The mass ratio of iron precipitate slag to graphite slag is 1:(0.1-0.2).

[0103] The numerical range described in the present disclosure includes not only the point values ​​listed above, but also any point values ​​between the above numerical ranges that are not listed. Due to space limitations and for the sake of simplicity, the present disclosure no longer exhaustively lists the specific point values ​​included in the range.

[0104] Compared with the prior art, the present disclosure has the following beneficial effects:

[0105] (1) The present invention mixes the powder and the first organic acid to carry out a leaching reaction, which can preferentially extract more than 98% of the lithium element and dissolve 100% of the aluminum, not only reducing the subsequent loss of lithium metal, but also preventing the introduction of aluminum in the subsequent leaching process, thereby achieving iron-aluminum separation. In addition, the leaching rate of nickel and cobalt in this process is less than 0.5%, and the leaching rate of manganese is less than 1.5%.

[0106] (2) The present invention mixes a lithium- and aluminum-containing leachate with a second organic acid to carry out a complexation reaction, which can efficiently and selectively remove 97.08% of the aluminum in the leachate without causing lithium loss, and the obtained aluminum salt can be decomposed into a high-purity Al(PO3)3 product at 1000°C, thereby obtaining a high-grade aluminum concentrate.

[0107] (3) The present invention mixes the first leaching residue and the composite ammonia source to carry out a leaching reaction to achieve selective copper removal. The copper leaching rate can reach 95.2%. The obtained copper-containing leachate can be used to prepare micro-nano copper with regular shapes. Compared with the existing technology, this method not only efficiently separates the copper foil in the lithium battery powder, but also can prepare micro-nano copper with higher value, thereby improving the utilization value of the copper foil in the lithium battery.

[0108] (4) The present disclosure performs acid leaching on the second leaching residue, which can reduce high-valent elements such as nickel, cobalt, and manganese in the battery powder into low-valent metal elements, thereby improving the leaching rate of Ni, Co, and Mn.

[0109] (5) The present invention combines a solution containing nickel, cobalt, manganese, and iron with a pH regulator to selectively remove 99.58% of the iron in the leachate. During the entire selective iron removal process, the total loss of Ni, Co, and Mn is only 1.1%. Compared with existing technologies, this method significantly reduces the loss of valuable metal elements, while simultaneously achieving high-value resource utilization of lithium battery aluminum foil and copper foil and reducing waste slag (ferroaluminum slag and graphite slag), achieving both good economic and environmental benefits.

[0110] (6) The method provided by the present disclosure reduces the loss of valuable metal elements lithium, nickel, cobalt, and manganese and the output of slag, while improving the recovery rate of metallic lithium and the high-value utilization of copper and aluminum in the leachate, reducing the process production cost, and realizing the resource utilization and reduction of solid waste in the wet recycling process of lithium batteries.

[0111] Still other aspects will become apparent upon reading and understanding the accompanying drawings and detailed description. BRIEF DESCRIPTION OF THE DRAWINGS

[0112] The accompanying drawings are used to provide a further understanding of the technical solution of this article and constitute a part of the specification. Together with the embodiments of this application, they are used to explain the technical solution of this article and do not constitute a limitation on the technical solution of this article.

[0113] FIG1 is a process flow chart of the step-by-step recycling of retired lithium batteries provided in Example 1 of the present disclosure.

[0114] FIG2 is a process flow chart of the step-by-step recycling of retired lithium batteries provided in Example 2 of the present disclosure.

[0115] FIG3 is a process flow chart of the step-by-step recycling of retired lithium batteries provided in Example 3 of the present disclosure.

[0116] FIG4 is a SEM image of nano-copper prepared in Example 3 of the present disclosure.

[0117] FIG5 is an XRD diagram of nano-copper prepared in Example 3 of the present disclosure.

[0118] FIG6 is a process flow chart of the step-by-step recycling of retired lithium batteries provided in Example 4 of the present disclosure.

[0119] FIG7 is a process flow chart of the step-by-step recycling of retired lithium batteries provided in Example 5 of the present disclosure.

[0120] FIG8 is a process flow chart for the step-by-step recycling of retired lithium batteries provided in Examples 8-9 of the present disclosure.

[0121] FIG9 is a process flow chart of the step-by-step recycling of retired lithium batteries provided in Comparative Example 2 of the present disclosure.

[0122] FIG10 is a process flow chart of the step-by-step recycling of retired lithium batteries provided in Comparative Example 3 of the present disclosure. DETAILED DESCRIPTION

[0123] The technical solution of the present disclosure is further described below through specific implementation methods. Those skilled in the art should understand that the embodiments are only used to help understand the present disclosure and should not be regarded as specific limitations of the present disclosure.

[0124] Example 1

[0125] This embodiment provides a method for recycling retired lithium batteries step by step, the process flow diagram of which is shown in FIG1 . The method includes the following steps:

[0126] (1) Immerse the retired lithium battery in 5.0 wt% salt water, fully discharge it at room temperature (25°C) for 36 hours, remove it, and drain it;

[0127] (2) dismantling the discharged retired lithium batteries, crushing them, drying them (300°C, 60 min), calcining them (650°C, 60 min), and sieving them to separate the fine powder 1 (including negative electrode powder, positive electrode powder, and fine copper powder and aluminum powder) and large particle size particles (including aluminum foil, copper foil, and separator paper) of the retired lithium batteries;

[0128] (3) further crushing and sieving the large-size particles separated in step (2) to obtain a retired lithium battery fine powder 2 and a mixture of copper foil and aluminum foil, and collecting the retired lithium battery fine powder 1 and the retired lithium battery powder 2 to obtain a powder;

[0129] (4) One-stage leaching: The powder obtained in step (3) was mixed with oxalic acid (0.5 mol / L) at a material-liquid ratio of 1 g:8 mL, stirred and reacted at 60° C. for 60 min, and filtered to obtain a reaction solution and a filter residue, respectively. The filter residue was washed with water and filtered to obtain a first leaching residue and a filtrate, respectively. The filtrate was mixed with the reaction solution to obtain a lithium- and aluminum-containing leachate.

[0130] (5) Secondary leaching: The first leaching residue obtained in step (4) was mixed with a composite ammonia source (comprising 20 mL of 4.5 mol / L ammonia water and 10 mL of 6 mol / L ammonium chloride solution) at a material-liquid ratio of 1 g:5 mL, stirred and reacted at 50° C. for 60 min, filtered to obtain a filter residue and a reaction liquid, washed the filter residue with water, filtered to obtain a second leaching residue and a filtrate, and the filtrate and the reaction liquid were mixed to obtain a copper-containing leachate;

[0131] (6) three-stage leaching: the second leaching residue obtained in step (5) was mixed with a 3.0 mol / L sulfuric acid solution at a material-liquid ratio of 1 g:5 mL, and the mixture was leached at 85° C. for 60 min, filtered to obtain a filter residue and a reaction liquid, washed and filtered to obtain a filtrate and graphite residue, and the filtrate and reaction liquid were mixed to obtain a solution containing nickel, cobalt, manganese, and iron;

[0132] (7) Four-stage leaching: adding a certain amount of NaHCO3 to the solution containing nickel, cobalt, manganese and iron obtained in step (6) so that the pH value of the solution is 3.3, reacting at 85°C for 60 minutes, aging for 240 minutes, filtering to obtain yellow sodium iron alum residue and aging liquid, washing the obtained yellow sodium iron alum residue with water, filtering to obtain iron alum residue and filtrate, mixing the filtrate and aging liquid to obtain a solution containing nickel, cobalt and manganese.

[0133] The present disclosure tests the leaching rate of the recovery treatment in each step of this embodiment, as well as the concentration of each element in the final nickel, cobalt and manganese-containing solution. The test results are shown in Table 1.

[0134] Table 1

[0135] Example 2

[0136] The difference between this embodiment and embodiment 1 is that the lithium-aluminum-containing leachate obtained in step (4) is subjected to aluminum precipitation treatment, and the specific steps include:

[0137] The lithium-containing and aluminum-containing leachate and phytic acid are mixed, subjected to a complexation reaction at 60°C for 30 minutes, aged for 60 minutes, and then filtered to obtain the lithium-containing leachate and aluminum-containing phytate. The aluminum-containing phytate is dried and then heat-treated at 1000°C for 70 minutes to obtain a high-grade aluminum concentrate, namely aluminum metaphosphate.

[0138] The rest of the preparation methods and parameters remained the same as in Example 1.

[0139] FIG2 shows a process flow chart for the step-by-step recycling of retired lithium batteries according to this embodiment.

[0140] The present disclosure tests the leaching rate of the first stage leaching in this embodiment, the content of each element in the lithium-containing leachate and the aluminum concentrate, and the test results are shown in Table 2.

[0141] Table 2

[0142] Example 3

[0143] The difference between this embodiment and embodiment 1 is that the copper-containing leachate obtained in step (5) is subjected to copper extraction treatment, and the specific steps include:

[0144] The copper-containing leachate and 10 mL of hexadecyltrimethylammonium bromide solution (comprising 0.4 g of hexadecyltrimethylammonium bromide and 10 mL of water) were mixed evenly, and then 0.6 g of glucose was added, and a reduction reaction was carried out in an ultrasonic water bath at 85° C. for 30 minutes to obtain a first mixed solution;

[0145] 0.5 g of sodium hydroxide and 0.5 g of ascorbic acid particles were respectively dissolved in 10 mL of water. The first mixed solution was then mixed with the sodium hydroxide solution and reacted in an 85° C. ultrasonic water bath for 30 minutes to obtain a second mixed solution. The second mixed solution was then mixed with the ascorbic acid solution and reacted in an 85° C. ultrasonic water bath for 20 minutes. After centrifugation and vacuum drying, pure nano-copper particles with regular shapes and a particle size D50 of 450-550 nm were obtained.

[0146] The rest of the preparation methods and parameters remained the same as in Example 1.

[0147] FIG3 shows a process flow chart for the step-by-step recycling of retired lithium batteries according to this embodiment.

[0148] FIG4 shows an SEM image of the nano-copper prepared in this example. As can be seen from the image, the product is spherical copper nanoparticles with a relatively uniform size distribution.

[0149] FIG5 shows the XRD pattern of the nano-copper prepared in this embodiment. As can be seen from the figure, there are three obvious diffraction peaks at diffraction angles 2θ = 43.30°, 50.43°, and 74.13°, which correspond to the (111), (200), and (220) crystal planes of the copper phase in the face-centered cubic system, respectively. In addition, XRD shows that there are no impurities such as copper oxide, cuprous oxide, and sodium sulfate in the prepared product, indicating that the prepared product is pure elemental copper in a cubic crystal form.

[0150] The present disclosure tests the leaching rate of the second-stage leaching, the concentration of the copper-containing leachate, and the content of each element in the nano-copper particles in this embodiment. The test results are shown in Table 3.

[0151] Table 3

[0152] Example 4

[0153] The difference between this embodiment and embodiment 1 is that SO2 gas is introduced at a flow rate of 2.0 L / min in step (6).

[0154] The rest of the preparation methods and parameters remained the same as in Example 1.

[0155] FIG6 shows a process flow chart for the step-by-step recycling of retired lithium batteries according to this embodiment.

[0156] The present disclosure tests the leaching rates of the three leaching stages in this embodiment and the concentrations of the elements in the final nickel, cobalt and manganese-containing solution. The test results are shown in Table 4.

[0157] Table 4

[0158] Example 5

[0159] The difference between this embodiment and embodiment 1 is that the iron alum slag in step (7) is subjected to iron extraction treatment, and the specific steps include:

[0160] The iron alum slag and graphite slag were mixed evenly in a mass ratio of 1:0.15, and then calcined at 1100°C for 70 minutes, washed with water, and filtered to obtain high-grade iron concentrate.

[0161] The rest of the preparation methods and parameters remained the same as in Example 1.

[0162] FIG7 shows a process flow chart for the step-by-step recycling of retired lithium batteries according to this embodiment.

[0163] The present disclosure tests the removal rate of the iron extraction treatment in this embodiment and the concentration of each element in the final iron concentrate. The test results are shown in Table 5.

[0164] Table 5

[0165] Example 6

[0166] This embodiment provides a method for recycling retired lithium batteries step by step, the method comprising the following steps:

[0167] (1) Immerse the retired lithium battery in 5.0 wt% salt water, fully discharge it at room temperature (25°C) for 36 hours, remove it, and drain it;

[0168] (2) dismantling the discharged retired lithium batteries, crushing them, drying them (300°C, 60 min), calcining them (650°C, 60 min), and sieving them to separate the fine powder 1 (including negative electrode powder, positive electrode powder, and fine copper powder and aluminum powder) and large particle size particles (including aluminum foil, copper foil, and separator paper) of the retired lithium batteries;

[0169] (3) further crushing and sieving the large-size particles separated in step (2) to obtain a retired lithium battery fine powder 2 and a mixture of copper foil and aluminum foil, and collecting the retired lithium battery fine powder 1 and the retired lithium battery powder 2 to obtain a powder;

[0170] (4) One-stage leaching: The powder obtained in step (3) was mixed with oxalic acid (0.55 mol / L) at a material-liquid ratio of 1 g:7.5 mL, stirred and reacted at 50° C. for 80 min, and filtered to obtain a reaction solution and a filter residue, respectively. The filter residue was washed with water and filtered to obtain a first leaching residue and a filtrate, respectively. The filtrate was mixed with the reaction solution to obtain a lithium- and aluminum-containing leachate;

[0171] (5) Secondary leaching: The first leaching residue obtained in step (4) was mixed with a composite ammonia source (comprising 15 mL of 4 mol / L ammonia water and 15 mL of 5 mol / L ammonium chloride solution) at a material-liquid ratio of 1 g:10 mL, stirred and reacted at 40° C. for 80 min, filtered to obtain a filter residue and a reaction liquid, washed the filter residue with water, filtered to obtain a second leaching residue and a filtrate, and the filtrate and the reaction liquid were mixed to obtain a copper-containing leachate;

[0172] (6) three-stage leaching: the second leaching residue obtained in step (5) was mixed with a 2.0 mol / L sulfuric acid solution at a material-liquid ratio of 1 g:4 mL, and the mixture was leached at 75° C. for 70 min. The mixture was filtered to obtain a filter residue and a reaction liquid. The filter residue was washed with water and filtered to obtain a filtrate and graphite residue. The filtrate and the reaction liquid were mixed to obtain a solution containing nickel, cobalt, manganese, and iron.

[0173] (7) Four-stage leaching: adding a certain amount of NaHCO3 to the solution containing nickel, cobalt, manganese and iron obtained in step (6) so that the pH value of the solution is 3, reacting at 80°C for 50 minutes, aging for 300 minutes, filtering to obtain yellow sodium ferroalite residue and aging liquid, washing the obtained yellow sodium ferroalite residue with water, filtering to obtain ferroalite residue and filtrate, mixing the filtrate and aging liquid to obtain a solution containing nickel, cobalt and manganese.

[0174] The present disclosure tests the leaching rate of the recovery treatment in each step of this embodiment, as well as the concentration of each element in the final nickel, cobalt and manganese-containing solution. The test results are shown in Table 6.

[0175] Table 6

[0176] Example 7

[0177] This embodiment provides a method for recycling retired lithium batteries step by step, the method comprising the following steps:

[0178] (1) Immerse the retired lithium battery in 5.0 wt% salt water, fully discharge it at room temperature (25°C) for 36 hours, remove it, and drain it;

[0179] (2) dismantling the discharged retired lithium batteries, crushing them, drying them (300°C, 60 min), calcining them (650°C, 60 min), and sieving them to separate the fine powder 1 (including negative electrode powder, positive electrode powder, and fine copper powder and aluminum powder) and large particle size particles (including aluminum foil, copper foil, and separator paper) of the retired lithium batteries;

[0180] (3) further crushing and sieving the large-size particles separated in step (2) to obtain a retired lithium battery fine powder 2 and a mixture of copper foil and aluminum foil, and collecting the retired lithium battery fine powder 1 and the retired lithium battery powder 2 to obtain a powder;

[0181] (4) One-stage leaching: The powder obtained in step (3) was mixed with oxalic acid (0.6 mol / L) at a material-liquid ratio of 1 g:7 mL, stirred and reacted at 80° C. for 40 min, and filtered to obtain a reaction solution and a filter residue, respectively. The filter residue was washed with water and filtered to obtain a first leaching residue and a filtrate, respectively. The filtrate was mixed with the reaction solution to obtain a lithium- and aluminum-containing leachate;

[0182] (5) Secondary leaching: The first leaching residue obtained in step (4) was mixed with a composite ammonia source (comprising 25 mL of 5 mol / L ammonia water and 5 mL of 6.5 mol / L ammonium chloride solution) at a material-liquid ratio of 1 g:7.5 mL, stirred and reacted at 60° C. for 40 min, filtered to obtain a filter residue and a reaction liquid, washed the filter residue with water, filtered to obtain a second leaching residue and a filtrate, and the filtrate and the reaction liquid were mixed to obtain a copper-containing leachate;

[0183] (6) three-stage leaching: the second leaching residue obtained in step (5) was mixed with a 4.0 mol / L sulfuric acid solution at a material-liquid ratio of 1 g:3 mL, and the mixture was leached at 95° C. for 50 min. The mixture was filtered to obtain a filter residue and a reaction liquid. The filter residue was washed with water and filtered to obtain a filtrate and graphite residue. The filtrate and the reaction liquid were mixed to obtain a solution containing nickel, cobalt, manganese, and iron.

[0184] (7) Four-stage leaching: adding a certain amount of NaHCO3 to the solution containing nickel, cobalt, manganese and iron obtained in step (6) so that the pH value of the solution is 3.5, reacting at 90°C for 40 minutes, aging for 180 minutes, filtering to obtain yellow sodium iron alum residue and aging liquid, washing the obtained yellow sodium iron alum residue with water, filtering to obtain iron alum residue and filtrate, mixing the filtrate and aging liquid to obtain a solution containing nickel, cobalt and manganese.

[0185] The present disclosure tests the leaching rate of the recovery treatment in each step of this embodiment, as well as the concentration of each element in the final nickel, cobalt and manganese-containing solution. The test results are shown in Table 7.

[0186] Table 7

[0187] Example 8

[0188] The difference between this embodiment and embodiment 4 is that SO2 gas is replaced by O2.

[0189] The rest of the preparation methods and parameters remained the same as in Example 4.

[0190] The present disclosure tested the leaching rates of the three leaching stages in this embodiment, and the test results are shown in Table 8.

[0191] Table 8

[0192] Example 9

[0193] The difference between this embodiment and embodiment 4 is that SO2 gas is replaced by N2.

[0194] The rest of the preparation methods and parameters remained the same as in Example 4.

[0195] FIG8 shows a process flow chart for the step-by-step recovery of retired lithium batteries according to Examples 8 and 9.

[0196] The present disclosure tested the leaching rates of the three leaching stages in this embodiment, and the test results are shown in Table 9.

[0197] Table 9

[0198] Example 10

[0199] The difference between this embodiment and embodiment 3 is that the composite ammonia source in step (5) contains only 30 mL of ammonia water.

[0200] The rest of the preparation methods and parameters remained the same as in Example 3.

[0201] The present disclosure tests the leaching rate of the second-stage leaching, the concentration of the copper-containing leachate, and the content of each element in the nano-copper particles in this embodiment. The test results are shown in Table 10.

[0202] Table 10

[0203] Example 11

[0204] The difference between this embodiment and embodiment 3 is that no cetyltrimethylammonium bromide solution is added during the copper extraction process.

[0205] The rest of the preparation methods and parameters remained the same as in Example 3.

[0206] The present disclosure tests the leaching rate of the second-stage leaching, the concentration of the copper-containing leachate, and the content of each element in the nano-copper particles in this embodiment. The test results are shown in Table 11.

[0207] Table 11

[0208] Example 12

[0209] The difference between this embodiment and embodiment 3 is that no sodium hydroxide solution and ascorbic acid solution are added during the copper extraction process.

[0210] The rest of the preparation methods and parameters remained the same as in Example 3.

[0211] The present disclosure tests the leaching rate of the second-stage leaching, the concentration of the copper-containing leachate, and the content of each element in the nano-copper particles in this embodiment. The test results are shown in Table 12.

[0212] Table 12

[0213] Comparative Example 1

[0214] The difference between this comparative example and Example 2 is that oxalic acid is replaced by an inorganic acid (i.e., hydrochloric acid).

[0215] The rest of the preparation methods and parameters remained the same as in Example 2.

[0216] The present disclosure tested the leaching rate of the first stage leaching of this comparative example, and the test results are shown in Table 13.

[0217] Table 13

[0218] Comparative Example 2

[0219] The difference between this comparative example and comparative example 1 is that the specific steps of aluminum precipitation treatment include:

[0220] The lithium- and aluminum-containing leachate is mixed with a certain amount of sodium carbonate to adjust the pH value of the leachate for aluminum precipitation to obtain a lithium-containing solution and an aluminum precipitate. The aluminum precipitate is dried and then heat-treated at 1000° C. for 70 minutes to obtain an aluminum-containing product.

[0221] The rest of the preparation methods and parameters remained the same as those in Comparative Example 1.

[0222] FIG9 shows a process flow chart for the step-by-step recycling of retired lithium batteries in this comparative example.

[0223] The present disclosure tested the concentration of the lithium-containing solution in this comparative example and the content of each element in the aluminum concentrate, and the test results are shown in Table 14.

[0224] Table 14

[0225] Comparative Example 3

[0226] The difference between this comparative example and Example 3 is that the composite ammonia source is replaced by an inorganic acid (i.e., hydrochloric acid).

[0227] The rest of the preparation methods and parameters remained the same as in Example 3.

[0228] FIG10 shows a process flow chart for the step-by-step recycling of retired lithium batteries in this comparative example.

[0229] The present disclosure tested the leaching rate of the second-stage leaching in this comparative example, the concentration of the copper-containing leachate, and the content of each element in the nano-copper particles. The test results are shown in Table 15.

[0230] Table 15

[0231] analyze:

[0232] From the data of the above embodiments and comparative examples, it can be seen that the method provided by the present disclosure reduces the loss of valuable metal elements lithium, nickel, cobalt, and manganese and the output of slag, while improving the recovery rate of metallic lithium and the high-value utilization of copper and aluminum in the leachate, reducing the process production cost, and realizing the resource utilization and reduction of solid waste in the wet recovery process of lithium batteries.

[0233] It can be seen from the data of Example 1 and Example 4 that the introduction of SO2 gas during the three-stage leaching process can place the entire leaching process in a reducing atmosphere, thereby accelerating the leaching rate of Ni, Co, and Mn.

[0234] It can be seen from the data of Examples 4 and 8-9 that if the SO2 gas introduced during the three-stage leaching process is replaced by O2 or N2, the leaching rate of Ni, Co, and Mn will slow down and the reaction time will be prolonged.

[0235] It can be seen from the data of Example 3 and Example 10 that if the composite ammonia source contains only aqueous ammonia, the leaching rate of Cu is slowed down and the leaching effect is poor, resulting in partial Cu loss.

[0236] It can be seen from the data of Example 3 and Example 11 that if the cetyltrimethylammonium bromide solution is not added during the copper extraction process, the copper ions in the copper-containing solution will not be dispersed well and the yield of nano-copper will be reduced.

[0237] It can be seen from the data of Examples 3 and 12 that if the sodium hydroxide solution and the ascorbic acid solution are not added during the copper extraction process, the copper ions in the copper-containing solution cannot form a stable precipitate, resulting in a decrease in the yield of nano-copper.

[0238] It can be seen from the data of Example 2 and Comparative Example 1 that if oxalic acid is replaced by an inorganic acid in one stage of the leaching process, selective aluminum extraction and preferential lithium extraction cannot be achieved.

[0239] It can be seen from the data of Comparative Examples 1 and 2 that if sodium carbonate is added instead of phytic acid during the aluminum precipitation process, selective aluminum removal cannot be achieved and some lithium loss will occur.

[0240] It can be seen from the data of Example 3 and Comparative Example 3 that if the composite ammonia source is replaced by an inorganic acid in the second-stage leaching process, selective copper extraction cannot be achieved.

Claims

1. A method for recycling retired lithium batteries step by step, the method comprising the following steps: (1) Pre-treating retired lithium batteries to obtain powder; (2) mixing the powder and a first organic acid to perform a leaching reaction to obtain a leaching solution containing lithium and aluminum and a first leaching residue; (3) mixing the first leaching residue and a composite ammonia source to carry out a leaching reaction to obtain a copper-containing leaching solution and a second leaching residue; (4) acid leaching the second leaching residue to obtain a solution containing nickel, cobalt, manganese and iron and a third leaching residue; (5) mixing the nickel, cobalt, manganese and iron-containing solution and a pH adjuster, reacting and aging to obtain a nickel, cobalt and manganese-containing solution and iron precipitate slag.

2. The method according to claim 1, wherein: The particle size D50 of the powder in step (1) is 48-78 μm.

3. The method according to claim 1 or 2, wherein: The pre-processing steps include: Discharge, crush, calcine and screen retired lithium batteries.

4. The method according to any one of claims 1 to 3, wherein: Step (2) the first organic acid comprises any one of oxalic acid, tartaric acid, malic acid or acetic acid, or a combination of at least two thereof; Optionally, in step (2), the concentration of the first organic acid is 0.5-0.6 mol / L; Optionally, in step (2), the concentration of the first organic acid is 0.5-0.55 mol / L; Optionally, the mass volume ratio of the powder and the first organic acid in step (2) is 1 g: (6-10) mL; Optionally, the mass volume ratio of the powder and the first organic acid in step (2) is 1 g: (7-8) mL; Optionally, the temperature of the leaching reaction in step (2) is 50-80°C; Optionally, the temperature of the leaching reaction in step (2) is 55-65°C; Optionally, the leaching reaction time in step (2) is 40-80 min; Optionally, the leaching reaction time in step (2) is 50-70 minutes.

5. The method according to any one of claims 1 to 4, wherein: After the leaching reaction in step (2) is completed, the obtained filter residue is washed with water and filtered to obtain a first leaching residue and a filtrate, and the filtrate is mixed with the reaction solution obtained by the leaching reaction to obtain a leaching solution containing lithium and aluminum.

6. The method according to any one of claims 1 to 5, wherein: The lithium-aluminum leaching solution in step (2) is subjected to aluminum precipitation treatment, and the specific steps include: The lithium-containing and aluminum-containing leaching solution is mixed with a second organic acid, subjected to complexation reaction, and aged to obtain a lithium-containing leaching solution and an aluminum salt; Optionally, the second organic acid includes any one or a combination of at least two of phytic acid, diethyl phosphate, alkyl phosphate or phenyl hypophosphite; Optionally, the molar ratio of Al ions to phytic acid in the lithium-aluminum-containing leaching solution is (2-4):1; Optionally, the molar ratio of Al ions to phytic acid in the lithium-aluminum-containing leaching solution is (3-3.5):1; Optionally, the temperature of the complexation reaction is 50-70°C; Optionally, the temperature of the complexation reaction is 55-65°C; Optionally, the complexation reaction time is 20-50 min; Optionally, the aging time is 30-60 min; Optionally, the aging time is 50-60 min; Optionally, the aluminum salt includes any one or a combination of at least two of aluminum phytate, diethyl aluminum phosphate, alkyl aluminum phosphate or phenyl aluminum hypophosphite; Optionally, the aluminum salt is heat-treated to obtain aluminum metaphosphate; Optionally, the temperature of the heat treatment is 800-1000°C; Optionally, the heat treatment temperature is 900-1000°C; Optionally, the heat treatment time is 60-90 min; Optionally, the heat treatment time is 65-75 min.

7. The method according to any one of claims 1 to 6, wherein: The composite ammonia source in step (3) includes aqueous ammonia and ammonium chloride solution; Optionally, the concentration of the ammonia water is 4-5 mol / L; Optionally, the concentration of the ammonia water is 4.5-5 mol / L; Optionally, the concentration of the ammonium chloride solution is 5-6.5 mol / L; Optionally, the concentration of the ammonium chloride solution is 5.5-6 mol / L; Optionally, the volume ratio of the ammonia water to the ammonium chloride solution is (15-25):(5-15); Optionally, in step (3), the mass volume ratio of the first leached residue to the composite ammonia source is 1 g: (5-10) mL; Optionally, in step (3), the mass volume ratio of the first leached residue to the composite ammonia source is 1 g: (5-7) mL; Optionally, the temperature of the leaching reaction in step (3) is 40-60°C; Optionally, the temperature of the leaching reaction in step (3) is 45-55°C; Optionally, the leaching reaction time in step (3) is 40-80 min; Optionally, the leaching reaction time in step (3) is 55-65 minutes.

8. The method according to any one of claims 1 to 7, wherein: After the leaching reaction in step (3) is completed, the obtained filter residue is washed with water and filtered to obtain a second leaching residue and a filtrate, and the filtrate is mixed with the reaction solution after the leaching reaction to obtain a copper-containing leaching solution.

9. The method according to any one of claims 1 to 8, wherein: Step (3) the copper-containing leaching solution is mixed with a reducing agent to carry out a reduction reaction to obtain pure nano copper; Optionally, the reducing agent includes any one of glucose, ascorbic acid or hydrazine hydrate, or a combination of at least two thereof; Optionally, a protective solution is added during the reduction reaction, and the protective solution includes a hexadecyltrimethylammonium bromide solution; Optionally, after the reduction reaction, sodium hydroxide and ascorbic acid are added to react; Optionally, the temperature of the reduction reaction is 75-95°C; Optionally, the reduction reaction time is 5-15min; Optionally, the particle size D50 of the pure nano copper is 450-550 nm.

10. The method according to any one of claims 1 to 9, wherein: The specific steps of the acid leaching in step (4) include: The second leaching residue is mixed with a sulfuric acid solution to perform an acid leaching reaction; Optionally, the concentration of the sulfuric acid solution is 2-4 mol / L; Optionally, the concentration of the sulfuric acid solution is 2.5-3.5 mol / L; Optionally, after the second leaching residue and the sulfuric acid solution are mixed, a reducing gas is introduced into the mixed solution; Optionally, the reducing gas includes any one of sulfur dioxide, sulfur monoxide or hydrogen sulfide, or a combination of at least two thereof; Optionally, the flow rate of the reducing gas is 1-2.5 L / min; Optionally, the flow rate of the reducing gas is 1.5-2 L / min; Optionally, the mass volume ratio of the second leached residue to the sulfuric acid solution is 1 g: (3-5) mL; Optionally, the mass volume ratio of the second leached residue to the sulfuric acid solution is 1 g: (4-5) mL; Optionally, the temperature of the acid leaching reaction is 75-95°C; Optionally, the acid leaching reaction time is 50-70 minutes.

11. The method according to any one of claims 1 to 10, wherein: After the acid leaching in step (4) is completed, the obtained filter residue is washed and filtered to obtain a filtrate and a third leached residue, and the filtrate is mixed with the reaction solution after the acid leaching to obtain a solution containing nickel, cobalt, manganese and iron; Optionally, the third leached residue is graphite residue.

12. The method according to any one of claims 1 to 11, wherein: The pH adjuster in step (5) includes any one of sodium bicarbonate, sodium carbonate, calcium carbonate, ammonium carbonate or ammonium bicarbonate, or a combination of at least two thereof; Optionally, after the nickel, cobalt, manganese, iron solution and the pH adjuster are mixed in step (5), the pH value of the mixed solution is 3-3.5; Optionally, after the nickel, cobalt, manganese, iron solution and the pH adjuster are mixed in step (5), the pH value of the mixed solution is 3.2-3.3; Optionally, the reaction temperature in step (5) is 75-95°C; Optionally, the reaction temperature in step (5) is 80-90°C; Optionally, the reaction time in step (5) is 40-60 min; Optionally, the reaction time in step (5) is 45-55 min; Optionally, the aging time in step (5) is 180-300 min; Optionally, the aging time in step (5) is 220-260 min.

13. The method according to any one of claims 1 to 12, wherein: After the aging in step (5), the obtained filter residue is washed with water and filtered to obtain iron precipitate and filtrate, and the filtrate is mixed with the aging liquid to obtain a solution containing nickel, cobalt and manganese.

14. The method according to any one of claims 1 to 13, wherein: The iron precipitated slag is post-processed to obtain iron concentrate; Optionally, the post-processing step includes: The iron precipitate slag is mixed with a carbon source and calcined to obtain an iron concentrate; Optionally, the carbon source includes graphite slag; Optionally, the mass ratio of the iron precipitate slag to the carbon source is 1:(0.1-0.2); Optionally, the mass ratio of the iron precipitate slag to the carbon source is 1:(0.13-0.17); Optionally, the calcination temperature is 1000-1200°C; Optionally, the calcination temperature is 1050-1150°C; Optionally, the calcination time is 60-90 min; Optionally, the calcination time is 70-80 min.

15. The method according to any one of claims 1 to 14, wherein: The method comprises the following steps: (1) Pre-processing the retired lithium batteries includes the following steps: Discharging, crushing, calcining and sieving retired lithium batteries to obtain powder with a particle size D50 of 48-78 μm; (2) mixing the powder and oxalic acid with a concentration of 0.5-0.6 mol / L according to the ratio of 1 g: (6-10) mL, stirring and reacting at a temperature of 50-80° C. for 40-80 min, filtering to obtain a filter residue and a reaction liquid, washing the obtained filter residue with water, filtering to obtain a first leaching residue and a filtrate, and mixing the filtrate with the reaction liquid to obtain a lithium- and aluminum-containing leaching liquid; (3) mixing the lithium-containing and aluminum-containing leachate with phytic acid, performing a complexation reaction at 50-70° C. for 20-50 min, aging for 30-60 min, filtering to obtain the lithium-containing leachate and aluminum-containing phytate, drying the aluminum-containing phytate, and then performing a heat treatment at 800-1000° C. for 60-90 min to obtain aluminum metaphosphate; Wherein, the molar ratio of Al ions and phytic acid in the lithium-aluminum leaching solution is (2-4):1; (4) mixing the first leaching residue and the composite ammonia source according to 1 g: (5-10) mL, performing a leaching stirring reaction at 40-60° C. for 40-80 min, filtering to obtain a filter residue and a reaction solution, washing the obtained filter residue with water, filtering to obtain a second leaching residue and a filtrate, mixing the filtrate with the reaction solution to obtain a copper-containing leaching solution; mixing the copper-containing leaching solution with a hexadecyltrimethylammonium bromide solution, adding a reducing agent and mixing, performing a reduction reaction to obtain a first mixed solution, mixing the first mixed solution with a sodium hydroxide and ascorbic acid solution, reacting to obtain pure nano copper with a particle size D50 of 450-550 nm; The composite ammonia source includes ammonia water with a concentration of 4-5 mol / L and ammonium chloride with a concentration of 5-6.5 mol / L. The volume ratio of solution, aqueous ammonia and ammonium chloride solution is (15-25):(5-15); (5) mixing the second leached residue with a sulfuric acid solution having a concentration of 2-4 mol / L, and introducing sulfur dioxide gas at a flow rate of 1-2.5 L / min, performing an acid leaching reaction at 75-95° C. for 50-70 min, filtering to obtain a filter residue and a reaction liquid, washing the obtained filter residue with water, filtering to obtain a filtrate and graphite residue, and mixing the filtrate with the reaction liquid to obtain a solution containing nickel, cobalt, manganese and iron; Wherein, the mass volume ratio of the second leaching residue and the sulfuric acid solution is 1g:(3-5)mL; (6) mixing the nickel, cobalt, manganese, and iron-containing solution with sodium bicarbonate to make the pH value of the mixed solution 3-3.5, then reacting at 75-95° C. for 40-60 min, aging for 180-300 min, filtering to obtain a filter residue and an aging liquid, washing the obtained filter residue with water, filtering to obtain an iron precipitate and a filtrate, and mixing the filtrate with the aging liquid to obtain a nickel, cobalt, and manganese-containing solution; The iron precipitate slag and graphite slag are mixed, and calcined at 1000-1200° C. for 60-90 min to obtain iron concentrate; Among them, the mass ratio of iron precipitate slag to graphite slag is 1:(0.1-0.2).

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