Method for step-by-step recycling of decommissioned lithium batteries

Abstract

This disclosure provides a method for the step-by-step recycling of retired lithium batteries. The method includes the following steps: pretreating the retired lithium batteries to obtain powder; mixing the powder with a first organic acid and performing a leaching reaction to obtain a lithium- and aluminum-containing leachate and a first leaching residue; mixing the first leaching residue with a composite ammonia source and performing a leaching reaction to obtain a copper-containing leachate and a second leaching residue; acid leaching the second leaching residue to obtain a nickel-, cobalt-, manganese-, and iron-containing solution and a third leaching residue; mixing the nickel-, cobalt-, manganese-, and iron-containing solution with a pH adjuster, reacting, and aging to obtain a nickel-, cobalt-, and manganese-containing solution and iron-laden slag. The method provided by this disclosure reduces the loss of valuable metal elements lithium, nickel, cobalt, and manganese, and the amount of slag produced, 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 recycling process of lithium batteries.

Description

Technical Field

[0001] This disclosure belongs to the field of lithium-ion battery recycling technology, specifically relating to a method for the step-by-step recycling of retired lithium batteries. Background Technology

[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 through physical and chemical methods, and then directly disassembled, calcined, and screened to separate battery powder, electrode sheets (aluminum foil, copper foil), separator and other components. Then, a series of processes such as acid leaching, impurity removal and extraction are used to obtain relatively pure ternary slurry and lithium-containing liquid. Because the raw materials contain a large amount of copper, aluminum, iron and other impurities, the impurity removal efficiency of the acid leaching solution is low and the loss of precious metal elements lithium, nickel, cobalt and manganese is large. A large amount of low-value sponge copper is also produced, as well as a large amount of low-grade iron-aluminum slag and graphite slag, which increases the production cost.

[0003] To address this, existing technologies have proposed several solutions. For example, CN106191466A discloses a method for recovering lithium from waste lithium iron phosphate batteries, which includes disassembling and removing the casing of waste lithium iron phosphate batteries to obtain battery cores; calcining the battery cores; crushing the battery cores and sieving them to obtain a mixture of positive and negative electrodes; adding alkaline solution to the mixture to remove residual aluminum, and filtering to obtain filter mud; leaching the filter mud with strong acid to obtain a lithium solution, filtering to remove carbon and iron phosphate, and obtaining a leachate; adjusting the pH of the leachate to produce a small amount of iron precipitate, and filtering to obtain a filtrate; removing copper impurities from the filtrate, and filtering to obtain a filtrate; adding solid sodium carbonate to the filtrate to obtain lithium carbonate precipitate. CN113285135A discloses a method for multi-component recycling of waste lithium iron phosphate batteries, comprising the following steps: disassembling and separating the waste lithium iron phosphate batteries after discharge treatment; processing the battery cells to obtain solvent recovery liquid; crushing and sorting the battery cells to obtain lithium iron phosphate coarse powder, copper powder, and aluminum powder; adding the lithium iron phosphate coarse powder to an acid solution for reaction, filtering to obtain an acid leaching solution and carbon slag, washing and drying the carbon slag to obtain high-carbon graphite; adjusting the pH value of the acid leaching solution, adding a reducing agent to remove copper, filtering to obtain a copper-removed solution and copper slag; adding an oxidant and an appropriate amount of phosphorus source to the copper-removed solution to obtain orthophosphate; adding the iron precipitation solution to an alkaline solution to obtain an aluminum-removed solution and aluminum slag; adding the aluminum precipitation solution to an alkaline solution to obtain an alkalizing solution and alkaline slag; evaporating and concentrating the alkalizing solution to obtain a lithium-rich solution, adding it 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 production costs are high.

[0005] Therefore, how to efficiently reduce the loss of valuable metal elements during the recycling process and lower the production cost is an urgent problem to be solved. Summary of the Invention

[0006] The following is an overview of the subject matter described in detail herein. This overview is not intended to limit the scope of the claims.

[0007] The purpose of this disclosure is to provide a method for the step-by-step recycling of retired lithium batteries. The method provided by this disclosure reduces the loss of valuable metal elements such as lithium, nickel, cobalt, and manganese, as well as the amount of slag produced, while improving the recovery rate of metallic lithium and the high-value utilization of copper and aluminum in the leachate. It also reduces the process production cost and realizes the resource utilization and reduction of solid waste in the wet recycling process of lithium batteries.

[0008] To achieve this purpose of disclosure, the following technical solution is adopted:

[0009] In a first aspect, this disclosure provides a method for the tiered recycling of retired lithium batteries, the method comprising the following steps:

[0010] (1) Pre-treat retired lithium batteries to obtain powder;

[0011] (2) The powder and the first organic acid are mixed and leaching reaction is carried out to obtain lithium and aluminum leaching solution and first leaching residue;

[0012] (3) The first leaching residue and the composite ammonia source are mixed and leaching reaction is carried out to obtain copper-containing leaching solution and second leaching residue;

[0013] (4) The second leaching residue is acid-leached to obtain a solution containing nickel, cobalt, manganese and iron and a third leaching residue;

[0014] (5) The nickel-, cobalt-, manganese-, and iron-containing solution and pH adjuster are mixed, reacted, and aged to obtain a nickel-, cobalt-, and manganese-containing solution and iron slag.

[0015] The method disclosed herein reduces the loss of valuable metal elements lithium, nickel, cobalt, and manganese, as well as the amount of slag produced. At the same time, it improves the recovery rate of metallic lithium and the high-value utilization of copper and aluminum in the leachate, reduces the process production cost, and realizes the resource utilization and reduction of solid waste in the wet recycling process of lithium batteries.

[0016] In this disclosure, the powder and the first organic acid are mixed and leached together, 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, thus achieving iron-aluminum separation. Moreover, the leaching rate of nickel and cobalt in this process is <0.5%, and the leaching rate of manganese is <1.5%.

[0017] In this disclosure, the first leaching residue and a composite ammonia source are mixed and subjected to a leaching reaction to achieve selective copper removal. The copper leaching rate can reach 95.2%, and the resulting copper-containing leachate can be used to prepare micro- and nano-copper with regular shapes. Compared with the prior art, this method not only efficiently separates copper foil from lithium battery powder, but also prepares higher-value micro- and nano-copper, thereby improving the utilization value of copper foil in lithium batteries.

[0018] In this disclosure, acid leaching of the second leaching residue can reduce high-valence nickel, cobalt, manganese and other elements in the battery powder to low-valence metal elements, thereby increasing the leaching rate of Ni, Co and Mn.

[0019] In this disclosure, a solution containing nickel, cobalt, manganese, and iron is mixed with a pH adjuster to selectively remove 99.58% of the iron element in the leaching solution. During the entire selective iron removal process, the total loss rate of Ni, Co, and Mn is only 1.1%. Compared with existing technologies, this method significantly reduces the loss of valuable metal elements while achieving high-value resource utilization of lithium battery aluminum and copper foils and reducing the volume of waste residue (iron-aluminum slag and graphite slag), thus possessing both good economic and environmental benefits.

[0020] As an optional technical solution of this disclosure, the particle size D50 of the powder in step (1) is 48-78μm, for example, it can be 48μm, 53μm, 58μm, 63μm, 68μm, 73μm or 78μm, etc.

[0021] In one embodiment, the preprocessing step includes:

[0022] The retired lithium batteries are discharged, crushed, calcined, and screened.

[0023] In one embodiment, the specific steps of the discharge include:

[0024] Retired lithium batteries are immersed in salt water to discharge them.

[0025] As an optional technical solution of this disclosure, in step (2), the first organic acid includes any one or a combination of at least two of oxalic acid, tartaric acid, malic acid or acetic acid.

[0026] It should be noted that when oxalic acid is mixed with powder, the resulting product includes both lithium oxalate and aluminum complexed oxalate.

[0027] In one embodiment, the concentration of the first organic acid in step (2) is 0.5-0.6 mol / L, for example, it can be 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 selected as 0.5-0.55 mol / L.

[0028] In one embodiment, the mass-volume ratio of the powder and the first organic acid in step (2) 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 selected as 1g:(7-8)mL.

[0029] In one embodiment, the leaching reaction temperature in step (2) is 50-80°C, for example, it can be 50°C, 55°C, 60°C, 65°C, 70°C, 75°C or 80°C, etc., and can be selected as 55-65°C.

[0030] In one embodiment, the leaching reaction time in step (2) is 40-80 min, for example, it can be 40 min, 50 min, 60 min, 70 min or 80 min, etc., and can be selected as 50-70 min.

[0031] 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. The filtrate and the reaction solution obtained from the leaching reaction are mixed to obtain a lithium-aluminum leaching solution.

[0032] In this disclosure, the purpose of washing and filtering the filter residue again is to remove metal ions adhering to the surface of the residue and reduce metal ion loss. The same principle applies to the following.

[0033] As an optional technical solution of this disclosure, the lithium-containing aluminum leaching solution in step (2) is subjected to aluminum precipitation treatment, and the specific steps include:

[0034] A lithium-containing and aluminum-containing leachate and a second organic acid are mixed, subjected to a complexation reaction, and aged to obtain a lithium-containing leachate and an aluminum salt.

[0035] In this disclosure, a lithium-containing aluminum leaching solution and a second organic acid are mixed and subjected to a complexation reaction, which can efficiently and selectively remove 97.08% of the aluminum in the leaching solution without causing lithium loss. Furthermore, the resulting aluminum salt can be decomposed into high-purity Al(PO3)3 products at 1000°C, thereby obtaining high-grade aluminum concentrate.

[0036] In one embodiment, the second organic acid includes any one or a combination of at least two of phytic acid, diethylphosphoric acid, alkylphosphoric acid, or phenylphosphine.

[0037] In one embodiment, the molar ratio of Al ions to phytic acid in the lithium-aluminum leaching solution is (2-4):1, for example, it can be 2:1, 2.5:1, 3:1, 3.5:1 or 4:1, etc., and can be selected as (3-3.5):1.

[0038] In one embodiment, the temperature of the complexation reaction is 50-70°C, for example, it can be 50°C, 55°C, 60°C, 65°C or 70°C, and can be selected as 55-65°C.

[0039] In one embodiment, the complexation reaction time is 20-50 min, for example, it can be 20 min, 25 min, 30 min, 35 min, 40 min, 45 min or 50 min.

[0040] In one embodiment, the aging time is 30-60 minutes, for example, it can be 30 minutes, 40 minutes, 50 minutes or 60 minutes, and can be selected as 50-60 minutes.

[0041] In one embodiment, the aluminum salt includes any one or a combination of at least two of aluminum phytate, diethylaluminum phosphate, alkylaluminum phosphate, or phenylaluminum hypophosphite.

[0042] In one embodiment, the aluminum salt is heat-treated to obtain aluminum metaphosphate.

[0043] In one embodiment, the heat treatment temperature is 800-1000℃, for example, it can be 800℃, 850℃, 900℃, 950℃ or 1000℃, etc., and can be selected as 900-1000℃.

[0044] In one embodiment, the heat treatment time is 60-90 minutes, for example, it can be 60 minutes, 65 minutes, 70 minutes, 75 minutes, 80 minutes, 85 minutes or 90 minutes, and can be selected as 65-75 minutes.

[0045] As an optional technical solution of this disclosure, the composite ammonia source in step (3) includes ammonia water and ammonium chloride solution.

[0046] In one embodiment, the concentration of the ammonia water is 4-5 mol / L, for example, it can be 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 selected as 4.5-5 mol / L.

[0047] In one embodiment, the concentration of the ammonium chloride solution is 5-6.5 mol / L, for example, it can be 5 mol / L, 5.5 mol / L, 6 mol / L or 6.5 mol / L, etc., and can be selected as 5.5-6 mol / L.

[0048] In one embodiment, the volume ratio of ammonia water to ammonium chloride solution is (15-25):(5-15), wherein the ammonia water is selected from the range of "15-25", for example, 15, 20 or 25, and the ammonium chloride solution is selected from the range of "5-15", for example, 5, 10 or 15.

[0049] 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 selected as 1g:(5-7)mL.

[0050] In one embodiment, the leaching reaction temperature in step (3) is 40-60°C, for example, it can be 40°C, 45°C, 50°C, 55°C or 60°C, and can be selected as 45-55°C.

[0051] In one embodiment, the leaching reaction time in step (3) is 40-80 min, for example, it can be 40 min, 50 min, 60 min, 70 min or 80 min, etc., and can be selected as 55-65 min.

[0052] 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 filtrate. The filtrate and the reaction liquid after the leaching reaction are mixed to obtain a copper-containing leaching solution.

[0053] As an optional technical solution of this disclosure, the copper-containing leaching solution and reducing agent in step (3) are mixed to carry out a reduction reaction to obtain pure nano-copper.

[0054] In one embodiment, the reducing agent includes any one or a combination of at least two of glucose, ascorbic acid, or hydrazine hydrate.

[0055] In this disclosure, glucose can reduce copper ions in solution to cuprous oxide.

[0056] In one embodiment, a protective liquid is added during the reduction reaction, the protective liquid comprising a hexadecyltrimethylammonium bromide solution.

[0057] It should be noted that the huge surface energy of copper nanoparticles not only causes them to spontaneously aggregate, but also makes them chemically very active and easily oxidized. The nitrogen and oxygen atoms with lone pairs of electrons on the side chains of the hexadecyltrimethylammonium bromide molecule can coordinate with the surface atoms of copper nanoparticles to form a protective layer, which not only prevents the aggregation of copper nanoparticles, but also protects them from oxidation.

[0058] In one embodiment, after the reduction reaction, sodium hydroxide and ascorbic acid are added to carry out the reaction.

[0059] It should be noted that the addition of ascorbic acid directly affects the nucleation and growth process of copper nanoparticles, thereby affecting the morphology of copper nanoparticles.

[0060] In one embodiment, the temperature of the reduction reaction is 75-95°C, for example, it can be 75°C, 80°C, 85°C, 90°C or 95°C.

[0061] In one embodiment, the reduction reaction takes 5-15 minutes, for example, 5 minutes, 10 minutes, or 15 minutes.

[0062] In one embodiment, the particle size D50 of the pure nano-copper is 450-550nm, for example, it can be 450nm, 475nm, 500nm, 525nm or 550nm, etc.

[0063] As an optional technical solution of this disclosure, the specific steps of acid leaching in step (4) include:

[0064] The second leaching residue is mixed with sulfuric acid solution to carry out an acid leaching reaction.

[0065] In one embodiment, the concentration of the sulfuric acid solution is 2-4 mol / L, for example, it can be 2 mol / L, 2.5 mol / L, 3 mol / L, 3.5 mol / L or 4 mol / L, etc., and can be selected as 2.5-3.5 mol / L.

[0066] In one embodiment, after the second leaching residue and sulfuric acid solution are mixed, a reducing gas is also introduced into the mixed solution.

[0067] In this disclosure, the introduction of a reducing gas can provide a reducing atmosphere and accelerate the leaching rate of metal elements.

[0068] In one embodiment, the reducing gas includes any one or a combination of at least two of sulfur dioxide, sulfur monoxide, or hydrogen sulfide.

[0069] In this disclosure, the industrial application of high-temperature waste gas SO2 can achieve the goal of treating waste with waste.

[0070] In one embodiment, the flow rate of the reducing gas is 1-2.5 L / min, for example, it can be 1 L / min, 1.5 L / min, 2 L / min or 2.5 L / min, etc., and can be selected as 1.5-2 L / min.

[0071] In one embodiment, the mass-to-volume ratio of the second leaching residue to the sulfuric acid solution is 1g:(3-5)mL, for example, it can be 1g:3mL, 1g:3.5mL, 1g:4mL, 1g:4.5mL or 1g:5mL, etc., and can be selected as 1g:(4-5)mL.

[0072] In one embodiment, the temperature of the acid leaching reaction is 75-95°C, for example, it can be 75°C, 80°C, 85°C, 90°C or 95°C.

[0073] In one embodiment, the acid leaching reaction time is 50-70 minutes, for example, 50 minutes, 55 minutes, 60 minutes, 65 minutes, or 70 minutes.

[0074] In one embodiment, after the acid leaching in step (4) is completed, the obtained filter residue is washed with water and filtered to obtain filtrate and third leaching residue. The filtrate and the reaction solution after acid leaching are mixed to obtain a solution containing nickel, cobalt, manganese and iron.

[0075] In one embodiment, the third leaching residue is graphite residue.

[0076] As an optional technical solution of this disclosure, the pH adjuster in step (5) includes any one or a combination of at least two of sodium bicarbonate, sodium carbonate, calcium carbonate, ammonium carbonate, or ammonium bicarbonate.

[0077] In one embodiment, after the nickel-, cobalt-, manganese-, and iron-containing solution and pH adjuster in step (5) are mixed, 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, etc., and can be selected as 3.2-3.3.

[0078] In one embodiment, the reaction temperature in step (5) is 75-95°C, for example, it can be 75°C, 80°C, 85°C, 90°C or 95°C, and can be selected as 80-90°C.

[0079] In one embodiment, the reaction time in step (5) is 40-60 min, for example, it can be 40 min, 45 min, 50 min, 55 min or 60 min, etc., and can be selected as 45-55 min.

[0080] 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 selected as 220-260 min.

[0081] In one embodiment, after aging in step (5), the obtained filter residue is washed with water and filtered to obtain iron slag and filtrate. The filtrate and aging liquid are mixed to obtain a solution containing nickel, cobalt and manganese.

[0082] As an optional technical solution of this disclosure, the iron slag is post-processed to obtain iron concentrate.

[0083] In one implementation, the post-processing steps include:

[0084] Iron concentrate is obtained by mixing iron slag and carbon source and calcining it.

[0085] In one embodiment, the carbon source comprises graphite slag.

[0086] In one embodiment, the mass ratio of the iron 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 selected as 1:(0.13-0.17).

[0087] In one embodiment, the calcination temperature is 1000-1200℃, for example, it can be 1000℃, 1050℃, 1100℃, 1150℃ or 1200℃, etc., and can be selected as 1050-1150℃.

[0088] In one embodiment, the calcination time is 60-90 minutes, for example, it can be 60 minutes, 65 minutes, 70 minutes, 75 minutes, 80 minutes, 85 minutes or 90 minutes, and can be selected as 70-80 minutes.

[0089] As an optional technical solution of this disclosure, the method includes the following steps:

[0090] (1) Pre-treatment of retired lithium batteries, specifically including the following steps:

[0091] Retired lithium batteries are discharged, crushed, calcined, and sieved to obtain powder with a particle size D50 of 48-78μm;

[0092] (2) The powder and oxalic acid with a concentration of 0.5-0.6 mol / L are mixed at a ratio of 1 g: (6-10) mL. After stirring and reacting at a temperature of 50-80℃ for 40-80 min, the mixture is filtered to obtain filter residue and reaction solution. The obtained filter residue is washed with water and filtered to obtain the first leaching residue and filtrate. The filtrate and reaction solution are mixed to obtain a lithium-aluminum leaching solution.

[0093] (3) The lithium-containing aluminum leaching solution and phytic acid are mixed and subjected to a complexation reaction at 50-70℃ for 20-50 min. After aging for 30-60 min, the mixture is filtered to obtain the lithium-containing leaching solution and aluminum-containing phytate. The aluminum-containing phytate is dried and then subjected to heat treatment at 800-1000℃ for 60-90 min to obtain aluminum metaphosphate.

[0094] The molar ratio of Al ions to phytic acid in the lithium- and aluminum-containing leachate is (2-4):1.

[0095] (4) The first leaching residue and the composite ammonia source are mixed at a ratio of 1g:(5-10)mL and leaching and stirring reaction is carried out at 40-60℃ for 40-80min. After filtration, filter residue and reaction solution are obtained. The obtained filter residue is washed with water and filtered to obtain the second leaching residue and filtrate. The filtrate and reaction solution are mixed to obtain copper-containing leaching solution. The copper-containing leaching solution is mixed with hexadecyltrimethylammonium bromide solution, and then a reducing agent is added and mixed to carry out a reduction reaction to obtain the first mixed solution. The first mixed solution is mixed with sodium hydroxide and ascorbic acid solution and reacted to obtain pure nano-copper with a particle size D50 of 450-550nm.

[0096] The composite ammonia source includes ammonia water with a concentration of 4-5 mol / L and ammonium chloride solution with a concentration of 5-6.5 mol / L, with a volume ratio of (15-25):(5-15).

[0097] (5) The second leaching residue is mixed with a sulfuric acid solution with a concentration of 2-4 mol / L, and sulfur dioxide gas with a flow rate of 1-2.5 L / min is introduced. The acid leaching reaction is carried out at 75-95℃ for 50-70 min. After filtration, filter residue and reaction solution are obtained. The obtained filter residue is washed with water and filtered to obtain filtrate and graphite residue. The filtrate and reaction solution are mixed to obtain a solution containing nickel, cobalt, manganese and iron.

[0098] The mass-to-volume ratio of the second leaching residue to the sulfuric acid solution is 1 g:(3-5) mL;

[0099] (6) The nickel-, cobalt-, manganese-, and iron-containing solution is mixed with sodium bicarbonate to make the pH of the mixed solution 3-3.5, and then reacted at 75-95℃ for 40-60 min. After aging for 180-300 min, the mixture is filtered to obtain filter residue and aging liquid. The obtained filter residue is washed with water and filtered to obtain iron slag and filtrate. The filtrate and aging liquid are mixed to obtain a nickel-, cobalt-, and manganese-containing solution.

[0100] The iron slag and graphite slag are mixed and calcined at 1000-1200℃ for 60-90 minutes to obtain iron concentrate;

[0101] The mass ratio of iron slag to graphite slag is 1:(0.1-0.2).

[0102] The numerical range described in this 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 brevity, this disclosure will not exhaustively list the specific point values ​​included in the range.

[0103] Compared with the prior art, this disclosure has the following beneficial effects:

[0104] (1) The present invention mixes powder and a first organic acid to carry out a leaching reaction, which can preferentially extract more than 98% of lithium elements and dissolve 100% of aluminum. This not only reduces the subsequent loss of lithium metal, but also does not introduce aluminum in the subsequent leaching process, thus achieving iron-aluminum separation. Moreover, the leaching rate of nickel and cobalt in this process is <0.5%, and the leaching rate of manganese is <1.5%.

[0105] (2) The present invention mixes lithium-containing aluminum leaching solution and a second organic acid to carry out a complexation reaction, which can efficiently and selectively remove 97.08% of the aluminum in the leaching solution without causing lithium loss, and the obtained aluminum salt can be decomposed into high-purity Al(PO3)3 product at 1000℃, thereby obtaining high-grade aluminum concentrate.

[0106] (3) In this disclosure, the first leaching residue and the composite ammonia source are mixed and leaching reaction is carried out to achieve selective copper removal. The copper leaching rate can reach 95.2%. The copper-containing leaching solution obtained can be used to prepare micro-nano copper with regular shape. Compared with the prior art, this method not only efficiently separates copper foil in lithium battery powder, but also prepares micro-nano copper with higher value, thereby improving the utilization value of copper foil in lithium batteries.

[0107] (4) The present invention performs acid leaching on the second leaching residue, which can reduce high-valence nickel, cobalt, manganese and other elements in the battery powder to low-valence metal elements, thereby increasing the leaching rate of Ni, Co and Mn.

[0108] (5) This disclosure mixes a solution containing nickel, cobalt, manganese, and iron with a pH adjuster, which can preferentially and selectively remove 99.58% of the iron element in the leaching solution. During the entire selective iron removal process, the total loss rate of Ni, Co, and Mn is only 1.1%. Compared with the prior art, this method significantly reduces the loss of valuable metal elements, while realizing the high-value resource utilization of lithium battery aluminum foil and copper foil and the reduction of waste residue (iron-aluminum slag and graphite slag), thus achieving both good economic and environmental benefits.

[0109] (6) The method provided in this disclosure reduces the loss of valuable metal elements lithium, nickel, cobalt and manganese and the amount of slag produced, while improving the recovery rate of lithium metal 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.

[0110] After reading and understanding the accompanying diagrams and detailed descriptions, the other aspects can be understood. Attached Figure Description

[0111] The accompanying drawings are used to provide a further understanding of the technical solutions in this paper and form part of the specification. They are used together with the embodiments of this application to explain the technical solutions in this paper and do not constitute a limitation on the technical solutions in this paper.

[0112] Figure 1 This is a process flow diagram for the step-by-step recycling of retired lithium batteries provided in Embodiment 1 of this disclosure.

[0113] Figure 2 This is a process flow diagram for the step-by-step recycling of retired lithium batteries provided in Embodiment 2 of this disclosure.

[0114] Figure 3 This is a process flow diagram for the step-by-step recycling of retired lithium batteries provided in Embodiment 3 of this disclosure.

[0115] Figure 4 This is a SEM image of the nano-copper prepared in Example 3 of this disclosure.

[0116] Figure 5 The image shows the XRD pattern of the nano-copper prepared in Example 3 of this disclosure.

[0117] Figure 6 This is a process flow diagram for the step-by-step recycling of retired lithium batteries provided in Embodiment 4 of this disclosure.

[0118] Figure 7 This is a process flow diagram for the step-by-step recycling of retired lithium batteries provided in Embodiment 5 of this disclosure.

[0119] Figure 8 The process flow diagrams for the step-by-step recycling of retired lithium batteries provided in Embodiments 8-9 of this disclosure are shown.

[0120] Figure 9 The process flow diagram for the step-by-step recycling of retired lithium batteries provided as Comparative Example 2 of this disclosure.

[0121] Figure 10 The process flow diagram for the step-by-step recycling of retired lithium batteries provided in Comparative Example 3 of this disclosure is shown. Detailed Implementation

[0122] The technical solutions of this disclosure will be further illustrated below through specific embodiments. Those skilled in the art should understand that the embodiments described are merely illustrative of this disclosure and should not be construed as specific limitations thereof.

[0123] Example 1

[0124] This embodiment provides a method for the step-by-step recycling of retired lithium batteries, and its process flow diagram is as follows: Figure 1 As shown, the method includes the following steps:

[0125] (1) Soak the retired lithium battery in 5.0wt% salt water, discharge it completely at room temperature (25℃) for 36 hours, then take it out and drain it;

[0126] (2) Disassemble, crush, dry (300℃, 60min), calcine (650℃, 60min), sieve, and separate the retired lithium battery to obtain fine powder 1 (including negative electrode powder, positive electrode powder, and fine copper powder and aluminum powder) and large particle size (including aluminum foil, copper foil and separator paper) of retired lithium battery.

[0127] (3) Further crush the large-diameter particles obtained in step (2), sieve them, and obtain retired lithium battery fine powder 2 and copper foil and aluminum foil mixture. Collect retired lithium battery fine powder 1 and retired lithium battery powder 2 to obtain powder.

[0128] (4) First stage leaching: The powder obtained in step (3) is mixed with oxalic acid (0.5 mol / L) at a ratio of 1 g: 8 mL. After stirring and reacting at 60°C for 60 min, the mixture is filtered to obtain a reaction solution and a filter residue. The filter residue is washed with water and filtered to obtain a first leaching residue and a filtrate. The filtrate is mixed with the reaction solution to obtain a lithium and aluminum leaching solution.

[0129] (5) Second-stage leaching: The first leaching residue obtained in step (4) is mixed with a composite ammonia source (including 20 mL of 4.5 mol / L ammonia water and 10 mL of 6 mol / L ammonium chloride solution) at a material-to-liquid ratio of 1 g: 5 mL. The mixture is stirred and reacted at 50 °C for 60 min. After filtration, filter residue and reaction solution are obtained. The filter residue is washed with water and filtered to obtain the second leaching residue and filtrate. The filtrate and reaction solution are mixed to obtain copper-containing leaching solution.

[0130] (6) Three-stage leaching: The second leaching residue obtained in step (5) is mixed with 3.0 mol / L sulfuric acid solution at a material-to-liquid ratio of 1 g: 5 mL, and leached at 85°C for 60 min. After filtration, filter residue and reaction solution are obtained. The filter residue is washed with water and filtered to obtain filtrate and graphite residue. The filtrate and reaction solution are mixed to obtain a solution containing nickel, cobalt, manganese and iron.

[0131] (7) Four-stage leaching: Add a certain amount of NaHCO3 to the nickel, cobalt, manganese and iron solution obtained in step (6) to make the pH value of the solution 3.3. React at 85℃ for 60 min, then age for 240 min, filter to obtain yellow sodium iron alum residue and aging liquid. Wash the obtained yellow sodium iron alum residue with water and filter to obtain iron alum residue and filtrate. Mix the filtrate and aging liquid to obtain a nickel, cobalt and manganese solution.

[0132] This disclosure tested the leaching rate of each step in the recovery process in 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.

[0133] Table 1

[0134]

[0135] Example 2

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

[0137] Lithium-containing and aluminum-containing leachate and phytic acid are mixed and subjected to a complexation reaction at 60°C for 30 minutes. After aging for 60 minutes, the mixture is filtered to obtain 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 high-grade aluminum concentrate, namely aluminum metaphosphate.

[0138] The remaining preparation methods and parameters are consistent with those in Example 1.

[0139] Figure 2 The process flow diagram for the step-by-step recycling of retired lithium batteries in this embodiment is shown.

[0140] This disclosure tested the leaching rate of the first leaching stage, the lithium-containing leachate, and the content of various elements in the aluminum concentrate in this embodiment. The test results are shown in Table 2.

[0141] Table 2

[0142] Process segment index lithium nickel cobalt manganese aluminum copper a period of leaching Leaching rate / % 98.16 0.22 0.06 0.36 100.02 0.18 Lithium-containing leachate <![CDATA[Concentration / mg·L -1 > 1094 9.67 0.62 6.54 5.32 1.09 Aluminum concentrate content / % 0.13 0.01 0.08 0.35 51.12 0.02

[0143] Example 3

[0144] The difference between this embodiment and Embodiment 1 is that the copper-containing leachate obtained in step (5) undergoes copper extraction treatment, specifically including the following steps:

[0145] The copper-containing leaching solution and 10 mL of hexadecyltrimethylammonium bromide solution (including 0.4 g of hexadecyltrimethylammonium bromide and 10 mL of water) were mixed evenly, and then 0.6 g of glucose was added. The mixture was then subjected to a reduction reaction in an ultrasonic water bath at 85 °C for 30 min to obtain the first mixture.

[0146] 0.5g of sodium hydroxide and 0.5g of ascorbic acid granules were dissolved in 10mL of water respectively. The first mixture was then mixed with the sodium hydroxide solution and reacted in an ultrasonic water bath at 85℃ for 30min to obtain a second mixture. The second mixture was then mixed with the ascorbic acid solution and reacted in an ultrasonic water bath at 85℃ for 20min. After centrifugation and vacuum drying, pure copper nanoparticles with regular shapes and a particle size D50 of 450-550nm were obtained.

[0147] The remaining preparation methods and parameters are consistent with those in Example 1.

[0148] Figure 3 The process flow diagram for the step-by-step recycling of retired lithium batteries in this embodiment is shown.

[0149] Figure 4 The SEM image of the nano-copper prepared in this embodiment is shown. As can be seen from the image, the product consists of spherical copper nanoparticles with a relatively uniform size distribution.

[0150] Figure 5 The XRD pattern of the nano-copper prepared in this embodiment is shown. 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 crystal system, respectively. The 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 with a cubic crystal form.

[0151] This disclosure tested the leaching rate of the two-stage leaching process in this embodiment, 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 3.

[0152] Table 3

[0153] Process segment index lithium nickel cobalt manganese aluminum copper Second stage leaching Leaching rate / % 0.01 0.01 0.02 0.01 0.01 95.35 Copper-containing leachate <![CDATA[Concentration / mg·L -1 > 0.23 0.44 0.21 0.18 0.01 319.6 Nano copper particles content / % 0.03 0.01 0.02 0.05 0.01 99.2

[0154] Example 4

[0155] The difference between this embodiment and embodiment 1 is that SO2 gas with a flow rate of 2.0 L / min is introduced in step (6).

[0156] The remaining preparation methods and parameters are consistent with those in Example 1.

[0157] Figure 6The process flow diagram for the step-by-step recycling of retired lithium batteries in this embodiment is shown.

[0158] This disclosure tested the leaching rate of the three leaching stages in this embodiment and the concentration of each element in the final solution containing nickel, cobalt, and manganese. The test results are shown in Table 4.

[0159] Table 4

[0160]

[0161]

[0162] Example 5

[0163] The difference between this embodiment and embodiment 1 is that the iron ore slag in step (7) undergoes iron extraction treatment, specifically including the following steps:

[0164] Iron ore slag and graphite slag were mixed evenly at a mass ratio of 1:0.15, then calcined at 1100℃ for 70 minutes, washed with water and filtered to obtain high-grade iron concentrate.

[0165] The remaining preparation methods and parameters are consistent with those in Example 1.

[0166] Figure 7 The process flow diagram for the step-by-step recycling of retired lithium batteries in this embodiment is shown.

[0167] This disclosure tested the iron removal rate of the iron extraction process in this embodiment and the concentration of each element in the final iron concentrate. The test results are shown in Table 5.

[0168] Table 5

[0169] Process segment index lithium nickel cobalt manganese iron copper Iron extraction processing Removal rate / % 0.01 0.02 0.02 0.05 99.62 0.08 Iron concentrate content / % 0.03 0.06 0.06 0.10 60.12 0.19

[0170] Example 6

[0171] This embodiment provides a method for the tiered recycling of retired lithium batteries, the method comprising the following steps:

[0172] (1) Soak the retired lithium battery in 5.0wt% salt water, discharge it completely at room temperature (25℃) for 36 hours, then take it out and drain it;

[0173] (2) Disassemble, crush, dry (300℃, 60min), calcine (650℃, 60min), sieve, and separate the retired lithium battery to obtain fine powder 1 (including negative electrode powder, positive electrode powder, and fine copper powder and aluminum powder) and large particle size (including aluminum foil, copper foil and separator paper) of retired lithium battery.

[0174] (3) Further crush the large-diameter particles obtained in step (2), sieve them, and obtain retired lithium battery fine powder 2 and copper foil and aluminum foil mixture. Collect retired lithium battery fine powder 1 and retired lithium battery powder 2 to obtain powder.

[0175] (4) First-stage leaching: The powder obtained in step (3) is mixed with oxalic acid (0.55mol / L) at a ratio of 1g:7.5mL. After stirring and reacting at 50℃ for 80min, the mixture is filtered to obtain a reaction solution and a filter residue. The filter residue is washed with water and filtered to obtain a first leaching residue and a filtrate. The filtrate is mixed with the reaction solution to obtain a lithium and aluminum leaching solution.

[0176] (5) Second-stage leaching: The first leaching residue obtained in step (4) is mixed with a composite ammonia source (including 15 mL of 4 mol / L ammonia water and 15 mL of 5 mol / L ammonium chloride solution) at a material-to-liquid ratio of 1 g: 10 mL. The mixture is stirred and reacted at 40 °C for 80 min. After filtration, filter residue and reaction solution are obtained. The filter residue is washed with water and filtered to obtain the second leaching residue and filtrate. The filtrate and reaction solution are mixed to obtain a copper-containing leaching solution.

[0177] (6) Three-stage leaching: The second leaching residue obtained in step (5) is mixed with 2.0 mol / L sulfuric acid solution at a material-to-liquid ratio of 1 g: 4 mL. The mixture is leached at 75°C for 70 min. After filtration, filter residue and reaction solution are obtained. The filter residue is washed with water and filtered to obtain filtrate and graphite residue. The filtrate and reaction solution are mixed to obtain a solution containing nickel, cobalt, manganese and iron.

[0178] (7) Four-stage leaching: Add a certain amount of NaHCO3 to the nickel, cobalt, manganese and iron solution obtained in step (6) to make the pH value of the solution 3. React at 80℃ for 50 min, then age for 300 min, filter to obtain yellow sodium iron alum residue and aging liquid. Wash the obtained yellow sodium iron alum residue with water and filter to obtain iron alum residue and filtrate. Mix the filtrate and aging liquid to obtain a nickel, cobalt and manganese solution.

[0179] This disclosure tested the leaching rate of each step in the recovery process in 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.

[0180] Table 6

[0181]

[0182]

[0183] Example 7

[0184] This embodiment provides a method for the tiered recycling of retired lithium batteries, the method comprising the following steps:

[0185] (1) Soak the retired lithium battery in 5.0wt% salt water, discharge it completely at room temperature (25℃) for 36 hours, then take it out and drain it;

[0186] (2) Disassemble, crush, dry (300℃, 60min), calcine (650℃, 60min), sieve, and separate the retired lithium battery to obtain fine powder 1 (including negative electrode powder, positive electrode powder, and fine copper powder and aluminum powder) and large particle size (including aluminum foil, copper foil and separator paper) of retired lithium battery.

[0187] (3) Further crush the large-diameter particles obtained in step (2), sieve them, and obtain retired lithium battery fine powder 2 and copper foil and aluminum foil mixture. Collect retired lithium battery fine powder 1 and retired lithium battery powder 2 to obtain powder.

[0188] (4) First-stage leaching: The powder obtained in step (3) is mixed with oxalic acid (0.6 mol / L) at a ratio of 1 g: 7 mL. After stirring and reacting at 80°C for 40 min, the mixture is filtered to obtain a reaction solution and a filter residue. The filter residue is washed with water and filtered to obtain a first leaching residue and a filtrate. The filtrate is mixed with the reaction solution to obtain a lithium and aluminum leaching solution.

[0189] (5) Second-stage leaching: The first leaching residue obtained in step (4) is mixed with a composite ammonia source (including 25 mL of 5 mol / L ammonia water and 5 mL of 6.5 mol / L ammonium chloride solution) at a material-to-liquid ratio of 1 g: 7.5 mL. The mixture is stirred and reacted at 60 °C for 40 min. After filtration, filter residue and reaction solution are obtained. The filter residue is washed with water and filtered to obtain the second leaching residue and filtrate. The filtrate and reaction solution are mixed to obtain copper-containing leaching solution.

[0190] (6) Three-stage leaching: The second leaching residue obtained in step (5) is mixed with 4.0 mol / L sulfuric acid solution at a material-to-liquid ratio of 1 g: 3 mL. The mixture is leached at 95°C for 50 min, filtered, and the residue and reaction solution are obtained. The residue is washed with water and filtered to obtain filtrate and graphite residue. The filtrate and reaction solution are mixed to obtain a solution containing nickel, cobalt, manganese and iron.

[0191] (7) Four-stage leaching: Add a certain amount of NaHCO3 to the nickel, cobalt, manganese and iron solution obtained in step (6) to make the pH value of the solution 3.5. React at 90℃ for 40 min, then age for 180 min, filter to obtain yellow sodium iron alum residue and aging liquid. Wash the obtained yellow sodium iron alum residue with water and filter to obtain iron alum residue and filtrate. Mix the filtrate and aging liquid to obtain a nickel, cobalt and manganese solution.

[0192] This disclosure tested the leaching rate of each step in the recovery process in 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.

[0193] Table 7

[0194]

[0195] Example 8

[0196] The difference between this embodiment and embodiment 4 is that SO2 gas is replaced with O2.

[0197] The remaining preparation methods and parameters are consistent with those in Example 4.

[0198] The leaching rate of the three leaching stages in this embodiment was tested, and the test results are shown in Table 8.

[0199] Table 8

[0200]

[0201]

[0202] Example 9

[0203] The difference between this embodiment and embodiment 4 is that SO2 gas is replaced with N2.

[0204] The remaining preparation methods and parameters are consistent with those in Example 4.

[0205] Figure 8 The process flow diagrams for the step-by-step recycling of retired lithium batteries in Examples 8 and 9 are shown.

[0206] The leaching rate of the three leaching stages in this embodiment was tested, and the test results are shown in Table 9.

[0207] Table 9

[0208] index lithium nickel cobalt manganese aluminum copper Example 4 Leaching rate / % 0.01 98.24 99.41 98.11 0.00 1.08 Comparative Example 5 Leaching rate / % 0.01 85.69 87.35 84.45 0.00 0.99

[0209] Example 10

[0210] The difference between this embodiment and embodiment 3 is that the composite ammonia source in step (5) contains only 30 mL of ammonia water.

[0211] The remaining preparation methods and parameters are consistent with those in Example 3.

[0212] This disclosure tested the leaching rate of the two-stage leaching process in this embodiment, 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 10.

[0213] Table 10

[0214] Process segment index lithium nickel cobalt manganese aluminum copper Second stage leaching Leaching rate / % 0.01 0.01 0.02 0.01 0.01 65.18 Copper-containing leachate <![CDATA[Concentration / mg·L -1 > 0.23 0.44 0.21 0.18 0.01 207.9 Nano copper content / % 0.03 0.01 0.02 0.05 0.01 98.9

[0215] Example 11

[0216] The difference between this embodiment and Embodiment 3 is that cetyltrimethylammonium bromide solution is not added during the copper extraction process.

[0217] The remaining preparation methods and parameters are consistent with those in Example 3.

[0218] This disclosure tested the leaching rate of the two-stage leaching process in this embodiment, 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 11.

[0219] Table 11

[0220] Process segment index lithium nickel cobalt manganese aluminum copper Second stage leaching Leaching rate / % 0.01 0.01 0.02 0.01 0.01 95.35 Copper-containing leachate <![CDATA[Concentration / mg·L -1 > 0.23 0.44 0.21 0.18 0.01 325.3 Nano copper content / % 0.03 0.01 0.02 0.05 0.01 63.5

[0221] Example 12

[0222] The difference between this embodiment and Embodiment 3 is that sodium hydroxide solution and ascorbic acid solution are not added during the copper extraction process.

[0223] The remaining preparation methods and parameters are consistent with those in Example 3.

[0224] This disclosure tested the leaching rate of the two-stage leaching process in this embodiment, 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 12.

[0225] Table 12

[0226]

[0227] Comparative Example 1

[0228] The difference between this comparative example and Example 2 is that oxalic acid is replaced with an inorganic acid (i.e., hydrochloric acid).

[0229] The remaining preparation methods and parameters are consistent with those in Example 2.

[0230] The leaching rate of the first leaching step of this comparative example was tested, and the test results are shown in Table 13.

[0231] Table 13

[0232] Process segment index lithium nickel cobalt manganese aluminum copper a period of leaching Leaching rate / % 99.15 92.32 91.24 92.13 99.02 90.11

[0233] Comparative Example 2

[0234] The difference between this comparative example and Comparative Example 1 is that the specific steps of the aluminum plating treatment include:

[0235] A lithium-containing aluminum leaching solution is mixed with a certain amount of sodium carbonate to adjust the pH value of the leaching solution for aluminum precipitation, resulting in a lithium-containing solution and an aluminum precipitate. The aluminum precipitate is dried and then heat-treated at 1000℃ for 70 minutes to obtain an aluminum-containing product.

[0236] The remaining preparation methods and parameters are consistent with those of Comparative Example 1.

[0237] Figure 9 The process flow diagram for the step-by-step recycling of retired lithium batteries in this comparative example is shown.

[0238] This disclosure tested the concentration of the lithium-containing solution and the content of each element in the aluminum concentrate in this comparative example, and the test results are shown in Table 14.

[0239] Table 14

[0240] Process segment index lithium nickel cobalt manganese aluminum copper lithium-containing solution <![CDATA[Concentration / mg·L -1 > 1099 4062 940.8 1676 5.02 295.9 Aluminum concentrate content / % 0.73 3.51 1.28 2.45 15.12 3.52

[0241] Comparative Example 3

[0242] The difference between this comparative example and Example 3 is that the composite ammonia source is replaced with an inorganic acid (i.e., hydrochloric acid).

[0243] The remaining preparation methods and parameters are consistent with those in Example 3.

[0244] Figure 10 The process flow diagram for the step-by-step recycling of retired lithium batteries in this comparative example is shown.

[0245] This disclosure tested the leaching rate of the two-stage leaching process, the concentration of the copper-containing leachate, and the content of each element in the nano-copper particles in this comparative example. The test results are shown in Table 15.

[0246] Table 15

[0247] Process segment index lithium nickel cobalt manganese aluminum copper Second stage leaching Leaching rate / % 0.91 93.38 91.99 92.89 0.01 95.55 Copper-containing leachate <![CDATA[Concentration / mg·L -1 > 0.23 4108 947.5 1691 0.01 320.3 Nano copper content / % 0.05 7.81 1.82 3.55 0.51 61.62

[0248] analyze:

[0249] As can be seen from the data in the above embodiments and comparative examples, the method provided in this disclosure reduces the loss of valuable metal elements lithium, nickel, cobalt, and manganese, as well as the amount of slag produced. At the same time, it improves the recovery rate of metallic lithium and the high-value utilization of copper and aluminum in the leachate, reduces the process production cost, and realizes the resource utilization and reduction of solid waste in the wet recycling process of lithium batteries.

[0250] As can be seen from the data of Examples 1 and 4, introducing SO2 gas during the three-stage leaching process can keep the entire leaching process in a reducing atmosphere, thereby accelerating the leaching rate of Ni, Co, and Mn.

[0251] Data from Examples 4 and 8-9 show that if the SO2 gas introduced during the three-stage leaching process is replaced with O2 or N2, the leaching rate of Ni, Co, and Mn slows down and the reaction time is prolonged.

[0252] Data from Examples 3 and 10 show that if the composite ammonia source contains only ammonia water, the leaching rate of Cu is slowed down and the leaching effect is poor, resulting in the loss of some Cu.

[0253] Data from Examples 3 and 11 show that if cetyltrimethylammonium bromide solution is not added during the copper extraction process, the dispersion of copper ions in the copper-containing solution is poor, and the yield of nano-copper is reduced.

[0254] Data from Examples 3 and 12 show that if sodium hydroxide solution and ascorbic acid solution are not added during the copper extraction process, copper ions in the copper-containing solution cannot form stable precipitates, resulting in a reduction in the yield of nano-copper.

[0255] As can be seen from the data of Example 2 and Comparative Example 1, if oxalic acid is replaced with inorganic acid during a leaching process, selective aluminum extraction and preferential lithium extraction cannot be achieved.

[0256] The data from Comparative Examples 1 and 2 show that if sodium carbonate is added instead of phytic acid during the aluminum precipitation process, selective aluminum removal cannot be achieved, and some lithium will be lost.

[0257] As can be seen from the data of Example 3 and Comparative Example 3, if the composite ammonia source is replaced with inorganic acid in the two-stage leaching process, selective copper extraction cannot be achieved.

Claims

1. A method for the tiered recycling of retired lithium batteries, the method comprising the following steps: (1) Pre-treat retired lithium batteries to obtain powder; (2) The powder and the first organic acid are mixed and leached to obtain a lithium and aluminum leaching solution and a first leaching residue, and more than 98% of the lithium element is extracted and 100% of the aluminum is dissolved. (3) The first leaching residue and the composite ammonia source are mixed and leaching reaction is carried out to obtain copper-containing leachate and second leaching residue; (4) The second leaching residue is acid-leached to obtain a solution containing nickel, cobalt, manganese and iron and a third leaching residue; The specific steps of acid leaching in step (4) include: mixing the second leaching residue with sulfuric acid solution to carry out an acid leaching reaction; After the second leaching residue and sulfuric acid solution are mixed, a reducing gas is also introduced into the mixed solution; (5) The nickel-, cobalt-, manganese-, and iron-containing solution and pH adjuster are mixed, reacted, and aged to obtain a nickel-, cobalt-, and manganese-containing solution and iron slag; Step (2) The first organic acid includes any one or a combination of at least two of oxalic acid, tartaric acid, malic acid or acetic acid; In step (2), the concentration of the first organic acid is 0.5-0.6 mol / L; The mass-to-volume ratio of the powder and the first organic acid in step (2) is 1 g:(6-10) mL; The lithium- and aluminum-containing leaching solution described in step (2) undergoes an aluminum precipitation treatment, specifically including the following steps: A lithium-containing and aluminum-containing leachate and a second organic acid are mixed, subjected to a complexation reaction, and aged to obtain a lithium-containing leachate and an aluminum salt. The second organic acid includes phytic acid; The molar ratio of Al ions to phytic acid in the lithium- and aluminum-containing leachate is (2-4):1; The composite ammonia source in step (3) includes ammonia water and ammonium chloride solution; The concentration of the ammonia solution is 4-5 mol / L; The concentration of the ammonium chloride solution is 5-6.5 mol / L; The volume ratio of the ammonia water to the ammonium chloride solution is (15-25):(5-15); Step (3) The mass-to-volume ratio of the first leaching residue to the composite ammonia source is 1g:(5-10)mL; After the solution containing nickel, cobalt, manganese, and iron in step (5) is mixed with the pH adjuster, the pH value of the mixed solution is 3-3.5; The reaction temperature in step (5) is 75-95℃.

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, wherein, The preprocessing steps include: The retired lithium batteries are discharged, crushed, calcined, and screened.

4. The method according to claim 1, wherein, In step (2), the concentration of the first organic acid is 0.5-0.55 mol / L.

5. The method according to claim 1, wherein, The mass-to-volume ratio of the powder and the first organic acid in step (2) is 1 g: (7-8) mL.

6. The method according to claim 1, wherein, The leaching reaction in step (2) is carried out at a temperature of 50-80℃.

7. The method according to claim 1, wherein, The leaching reaction in step (2) is carried out at a temperature of 55-65℃.

8. The method according to claim 1, wherein, The leaching reaction time in step (2) is 40-80 min.

9. The method according to claim 1, wherein, The leaching reaction time in step (2) is 50-70 min.

10. The method according to claim 1, wherein, After the leaching reaction in step (2) is completed, the obtained filter residue is washed with water and filtered to obtain the first leaching residue and filtrate. The filtrate and the reaction solution obtained from the leaching reaction are mixed to obtain a lithium-aluminum leaching solution.

11. The method according to claim 1, wherein, The molar ratio of Al ions to phytic acid in the lithium- and aluminum-containing leachate is (3-3.5):

1.

12. The method according to claim 1, wherein, The temperature of the complexation reaction is 50-70℃.

13. The method according to claim 1, wherein, The temperature of the complexation reaction is 55-65℃.

14. The method according to claim 1, wherein, The complexation reaction takes 20-50 minutes.

15. The method according to claim 1, wherein, The aging time is 30-60 minutes.

16. The method according to claim 1, wherein, The aging time is 50-60 minutes.

17. The method according to claim 1, wherein, The aluminum salts include aluminum-containing phytates.

18. The method according to claim 1, wherein, The aluminum salt was subjected to heat treatment to obtain aluminum metaphosphate.

19. The method according to claim 18, wherein, The heat treatment temperature is 800-1000℃.

20. The method according to claim 18, wherein, The heat treatment temperature is 900-1000℃.

21. The method according to claim 18, wherein, The heat treatment time is 60-90 minutes.

22. The method according to claim 18, wherein, The heat treatment time is 65-75 minutes.

23. The method according to claim 1, wherein, The concentration of the ammonia water is 4.5-5 mol / L.

24. The method according to claim 1, wherein, The concentration of the ammonium chloride solution is 5.5-6 mol / L.

25. The method according to claim 1, wherein, In step (3), the mass-to-volume ratio of the first leaching residue to the composite ammonia source is 1 g:(5-7) mL.

26. The method according to claim 1, wherein, The leaching reaction in step (3) is carried out at a temperature of 40-60℃.

27. The method according to claim 1, wherein, The leaching reaction in step (3) is carried out at a temperature of 45-55℃.

28. The method according to claim 1, wherein, The leaching reaction time in step (3) is 40-80 min.

29. The method according to claim 1, wherein, The leaching reaction time in step (3) is 55-65 min.

30. The method according to claim 1, 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 filtrate. The filtrate and the reaction solution after the leaching reaction are mixed to obtain a copper-containing leaching solution.

31. The method according to claim 1, wherein, In step (3), the copper-containing leaching solution and reducing agent are mixed to carry out a reduction reaction and obtain pure nano-copper.

32. The method according to claim 31, wherein, The reducing agent includes any one or a combination of at least two of glucose, ascorbic acid, or hydrazine hydrate.

33. The method according to claim 31, wherein, A protective solution is also added during the reduction reaction, which includes a hexadecyltrimethylammonium bromide solution.

34. The method according to claim 31, wherein, After the reduction reaction, sodium hydroxide and ascorbic acid are added to carry out the reaction.

35. The method according to claim 31, wherein, The reduction reaction is carried out at a temperature of 75-95℃.

36. The method according to claim 31, wherein, The reduction reaction takes 5-15 minutes.

37. The method according to claim 31, wherein, The particle size D50 of the pure nano-copper is 450-550 nm.

38. The method according to claim 1, wherein, The concentration of the sulfuric acid solution is 2-4 mol / L.

39. The method according to claim 1, wherein, The concentration of the sulfuric acid solution is 2.5-3.5 mol / L.

40. The method according to claim 1, wherein, The reducing gas includes any one or a combination of at least two of sulfur dioxide, sulfur monoxide, or hydrogen sulfide.

41. The method according to claim 1, wherein, The flow rate of the reducing gas is 1-2.5 L / min.

42. The method according to claim 1, wherein, The flow rate of the reducing gas is 1.5-2 L / min.

43. The method according to claim 1, wherein, The mass-to-volume ratio of the second leaching residue to the sulfuric acid solution is 1 g:(3-5) mL.

44. The method according to claim 1, wherein, The mass-to-volume ratio of the second leaching residue to the sulfuric acid solution is 1 g:(4-5) mL.

45. The method according to claim 1, wherein, The acid leaching reaction is carried out at a temperature of 75-95℃.

46. ​​The method according to claim 1, wherein, The acid leaching reaction takes 50-70 minutes.

47. The method according to claim 1, wherein, After the acid leaching in step (4) is completed, the obtained filter residue is washed with water and filtered to obtain filtrate and third leaching residue. The filtrate and the reaction solution after acid leaching are mixed to obtain a solution containing nickel, cobalt, manganese and iron.

48. The method according to claim 1, wherein, The third leaching residue is graphite residue.

49. The method according to claim 1, wherein, The pH adjuster in step (5) includes any one or a combination of at least two of sodium bicarbonate, sodium carbonate, calcium carbonate, ammonium carbonate, or ammonium bicarbonate.

50. The method according to claim 1, wherein, After the solution containing nickel, cobalt, manganese and iron in step (5) is mixed with the pH adjuster, the pH value of the mixed solution is 3.2-3.

3.

51. The method according to claim 1, wherein, The reaction temperature in step (5) is 80-90℃.

52. The method according to claim 1, wherein, The reaction time in step (5) is 40-60 min.

53. The method according to claim 1, wherein, The reaction time in step (5) is 45-55 min.

54. The method according to claim 1, wherein, The aging time in step (5) is 180-300 min.

55. The method according to claim 1, wherein, The aging time in step (5) is 220-260 min.

56. The method according to claim 1, wherein, After aging in step (5), the obtained filter residue is washed and filtered to obtain iron slag and filtrate. The filtrate and aging liquid are mixed to obtain a solution containing nickel, cobalt and manganese.

57. The method according to claim 1, wherein, The iron slag is further processed to obtain iron concentrate.

58. The method according to claim 57, wherein, The post-processing steps include: Iron concentrate is obtained by mixing iron slag and carbon source and calcining it.

59. The method according to claim 58, wherein, The carbon source includes graphite slag.

60. The method according to claim 58, wherein, The mass ratio of the iron slag to the carbon source is 1:(0.1-0.2).

61. The method according to claim 58, wherein, The mass ratio of the iron slag to the carbon source is 1:(0.13-0.17).

62. The method according to claim 58, wherein, The calcination temperature is 1000-1200℃.

63. The method according to claim 58, wherein, The calcination temperature is 1050-1150℃.

64. The method according to claim 58, wherein, The calcination time is 60-90 minutes.

65. The method according to claim 58, wherein, The calcination time is 70-80 minutes.

66. The method according to claim 1, wherein, The method includes the following steps: (1) Pre-treatment of retired lithium batteries, specifically including the following steps: Retired lithium batteries are discharged, crushed, calcined, and sieved to obtain powder with a particle size D50 of 48-78μm; (2) The powder and oxalic acid with a concentration of 0.5-0.6 mol / L are mixed at a ratio of 1 g: (6-10) mL. After stirring and reacting at a temperature of 50-80℃ for 40-80 min, the mixture is filtered to obtain filter residue and reaction solution. The obtained filter residue is washed with water and filtered to obtain the first leaching residue and filtrate. The filtrate and reaction solution are mixed to obtain a lithium and aluminum leaching solution. (3) The lithium-containing aluminum leaching solution and phytic acid are mixed and subjected to a complexation reaction at 50-70℃ for 20-50 min. After aging for 30-60 min, the mixture is filtered to obtain the lithium-containing leaching solution and the aluminum-containing phytate. The aluminum-containing phytate is dried and then subjected to heat treatment at 800-1000℃ for 60-90 min to obtain aluminum metaphosphate. The molar ratio of Al ions to phytic acid in the lithium- and aluminum-containing leachate is (2-4):

1. (4) The first leaching residue and the composite ammonia source are mixed at a ratio of 1g:(5-10)mL and leaching and stirring reaction is carried out at 40-60℃ for 40-80min. After filtration, filter residue and reaction solution are obtained. The obtained filter residue is washed with water and filtered to obtain the second leaching residue and filtrate. The filtrate and reaction solution are mixed to obtain copper-containing leaching solution. The copper-containing leaching solution is mixed with hexadecyltrimethylammonium bromide solution, and then a reducing agent is added and mixed to carry out a reduction reaction to obtain the first mixed solution. The first mixed solution is mixed with sodium hydroxide and ascorbic acid solution and reacted to obtain pure nano-copper with a particle size D50 of 450-550nm. The composite ammonia source includes ammonia water with a concentration of 4-5 mol / L and ammonium chloride solution with a concentration of 5-6.5 mol / L, with a volume ratio of (15-25):(5-15). (5) The second leaching residue is mixed with a sulfuric acid solution with a concentration of 2-4 mol / L, and sulfur dioxide gas with a flow rate of 1-2.5 L / min is introduced. The acid leaching reaction is carried out at 75-95℃ for 50-70 min. The mixture is filtered to obtain filter residue and reaction solution. The obtained filter residue is washed with water and filtered to obtain filtrate and graphite residue. The filtrate and reaction solution are mixed to obtain a solution containing nickel, cobalt, manganese and iron. The mass-to-volume ratio of the second leaching residue to the sulfuric acid solution is 1 g:(3-5) mL; (6) Mix the nickel, cobalt, manganese and iron solution with sodium bicarbonate to make the pH of the mixed solution 3-3.5, then react at 75-95℃ for 40-60 min, age for 180-300 min, filter to obtain filter residue and aging liquid, wash the obtained filter residue with water and filter to obtain iron slag and filtrate, mix the filtrate and aging liquid to obtain nickel, cobalt and manganese solution; The iron slag and graphite slag are mixed and calcined at 1000-1200℃ for 60-90 minutes to obtain iron concentrate; The mass ratio of iron slag to graphite slag is 1:(0.1-0.2).

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