Efficient separation and recovery method of retired ternary lithium battery positive electrode material

Abstract

The invention provides an efficient separation and recovery method of a retired ternary lithium battery positive electrode material, which comprises the following steps of: performing controlled reduction on Ni < 3 + > / Ni < 4 + > / Co < 3 + > / Co < 4 + > / Mn < 4 + > in the positive electrode material by utilizing a reducing atmosphere generated by pyrolysis of waste polypropylene at 600-800 DEG C; lithium is dissolved out preferentially by combining water leaching subsequently, the lithium recovery rate is increased, and multi-metal co-dissolution is avoided; a roasted product is subjected to wet magnetic separation, nickel, cobalt and manganese can be preliminarily enriched in magnetic slag, graphite can be preliminarily enriched in non-magnetic slag, and the separation process is greatly simplified; and after the magnetic slag is subjected to mild acid leaching, manganese, cobalt and nickel are efficiently separated through a P204 / P507 cascade extraction system, a high-purity sulfate solution is obtained, and the obtained high-purity sulfate solution can be prepared into high-purity metal salt through crystallization or precipitation. According to the process, strong acid and a high-price reducing agent required by a traditional wet method are abandoned, the problems of high energy consumption and fluorine-containing waste gas of pyrogenic process high-temperature smelting are avoided, low reagent consumption, low environmental load, high resource recovery efficiency and good raw material adaptability are achieved, and a new technically feasible path is provided for large-scale and green cyclic utilization of the retired power lithium battery.

Description

Technical Field

[0001] This invention belongs to the field of cathode material recycling technology, and relates to an efficient separation and recycling method for cathode materials of retired ternary lithium batteries. Background Technology

[0002] With the advancement of carbon emission reduction policies and the rapid development of the new energy vehicle industry, the demand for power lithium batteries has surged. Currently, the lifespan of mainstream power batteries is approximately 4-6 years, and my country will face a large-scale wave of battery obsolescence in the coming years. It is estimated that by 2026, the total amount of waste lithium-ion batteries awaiting disposal will reach 2.312 million tons, making efficient and green recycling extremely urgent. Ternary lithium batteries (NCM) have become one of the mainstream choices due to their high energy density and good low-temperature performance. The lithium, nickel, cobalt, and manganese contained in their cathode materials are all scarce strategic resources with uneven global distribution. Failure to effectively recycle them will result in resource waste; at the same time, the heavy metals and electrolyte components in waste batteries are toxic, and improper disposal will pose a serious threat to the ecological environment and human health.

[0003] Studies have shown that recovering valuable metals from spent ternary lithium batteries has significant advantages over extracting them from primary ores. Primary ores have complex compositions, many impurities, and complicated purification processes with large fluctuations in product purity; while the metal element ratios in spent battery cathode materials are uniform and high in purity, and the purity of the recovered metal salt products can reach over 90%, with excellent recycling adaptability, and can be directly used in the manufacture of new batteries, forming a closed-loop resource system.

[0004] However, current mainstream recycling technologies face significant bottlenecks. Hydrometallurgy relies on strong acids and large amounts of reducing agents, resulting in high costs and difficult-to-treat wastewater. Furthermore, simultaneous leaching of multiple metals leads to poor lithium selectivity and low leaching rates, and the subsequent nickel-cobalt-manganese separation process is lengthy and complex. Pyrometallurgy is extremely energy-intensive, with lithium easily volatilized and lost, resulting in low recovery rates, while also generating toxic waste gases containing fluorine. Emerging technologies such as direct regeneration have stringent requirements for the uniformity of raw material composition, making them unsuitable for the complex sources of retired batteries. Existing technologies face severe challenges in terms of economic viability, environmental friendliness, and lithium recovery efficiency.

[0005] Therefore, there is an urgent need to propose a green and efficient recycling strategy that fundamentally avoids the problems of lengthy processes, low lithium recovery rates, and secondary pollution caused by multi-metal simultaneous leaching. This strategy would significantly simplify the process, improve resource utilization efficiency, and significantly reduce chemical consumption and environmental impact, providing an innovative solution for the recycling of waste lithium battery resources that is both technically feasible and environmentally sustainable. Summary of the Invention

[0006] In view of the shortcomings of the existing technology, the present invention discloses an efficient separation and recycling method for retired ternary lithium battery cathode materials, which can solve or at least alleviate one or more of the above-mentioned problems or other problems existing in the prior art.

[0007] This invention discloses an efficient method for separating and recycling cathode materials from retired ternary lithium batteries, comprising the following steps:

[0008] (1) Weigh the raw materials

[0009] Weigh the dried black powder and polypropylene powder at a mass ratio of 1:(0.5~1.5);

[0010] (2) Mixing and grinding

[0011] Black powder, polypropylene powder, and grinding balls are loaded into an agate ball mill jar for mechanical ball milling.

[0012] (3) Calcination and reduction treatment

[0013] The mixed and ground materials are placed in a quartz crucible and calcined in a tube furnace at 600-800°C for 1-2 hours under argon protection. The furnace is then cooled to room temperature to obtain the calcined product.

[0014] (4) Magnetic separation-water immersion separation

[0015] The calcined product was refined by vibration grinding, then deionized water was added, and the mixture was stirred and reacted in a constant temperature water bath. The resulting mixture was separated by wet magnetic separation to obtain magnetic slag rich in nickel, cobalt and manganese, non-magnetic slag mainly composed of graphite, and lithium-containing leachate.

[0016] (5) Lithium extraction from leachate

[0017] The lithium-containing leachate was vacuum filtered to remove impurities. The filtrate was concentrated to 1 / 5 of its original volume, and then a saturated Na₂CO₃ solution was slowly added while stirring to allow the lithium to dissolve. + The solid-liquid mixture was fully precipitated as Li2CO3; the precipitate was filtered under vacuum, and the precipitate was repeatedly washed with deionized water until no impurity ions were present. Then it was dried at 100°C for 2 hours to obtain high-purity lithium carbonate.

[0018] (6) Recycling of valuable metals from magnetic slag

[0019] Acid leaching: Add an appropriate amount of dilute sulfuric acid to the dried magnetic slag at a certain liquid-solid ratio, stir thoroughly at room temperature to dissolve Ni, Co, and Mn, and then filter under vacuum to obtain Ni-containing... 2+ Co 2+ Mn 2+ Acid leaching solution;

[0020] Iron precipitation: NaOH solution was then added dropwise to adjust the pH to a suitable range, and the mixture was stirred thoroughly to hydrolyze the iron impurities and form ferric hydroxide precipitate. Iron was then removed by filtration.

[0021] Stepwise extraction: The iron-removing solution is first extracted with P204 at pH 3.0–3.5 to separate manganese. The loaded organic phase is washed and back-extracted to obtain a manganese sulfate solution. After manganese extraction, the aqueous phase is selectively extracted with P507 at pH 3.5–4.0. After washing and back-extraction, a cobalt sulfate solution is obtained. Nickel is enriched in the final aqueous phase. The obtained manganese sulfate, cobalt sulfate and nickel sulfate solutions are all crystallized or precipitated to obtain high-purity metal salts.

[0022] Preferably, in the mixed grinding, black powder, polypropylene powder and grinding balls are loaded into an agate ball mill jar at a ratio of 1:5 of the total mass of raw materials to agate grinding balls, and then mechanically ball-milled.

[0023] Preferably, in the mixed grinding, the ball mill speed is 300 r / min and the grinding time is 30 min.

[0024] Preferably, in the calcination and reduction process, the mixed and ground material is transferred to a quartz crucible, and then the crucible is placed in a tube furnace and heated to 600~800°C at a heating rate of 10°C / min under an argon protective atmosphere.

[0025] Preferably, in the magnetic separation-water immersion separation, the roasted product is ground for 10 minutes using a vibratory mill.

[0026] Preferably, in the magnetic separation-water leaching separation, the ground roasted product and deionized water are mixed at a solid-liquid ratio of 1:(5~25), and then placed in a constant temperature water bath at 20~80℃ and stirred for 20~100 min.

[0027] Preferably, in the magnetic separation-water immersion separation, the magnetic field strength of the wet magnetic separator is 300mT.

[0028] Preferably, in the lithium extraction from the leachate, the clarified filtrate obtained by vacuum filtration is transferred to a rotary evaporator and concentrated to 1 / 5 of its original volume at 60°C.

[0029] Preferably, in the lithium extraction from the leachate, the concentrated solution is slowly added to a saturated Na₂CO₃ solution at a lithium ion to sodium carbonate molar ratio of 1:1, and stirred continuously for 30 minutes to allow the Li₂CO₃ to be extracted. + It fully precipitates as Li2CO3;

[0030] Preferably, in the recovery of valuable metals from the magnetic slag, after the magnetic slag is dried, 3.5 mol / L dilute sulfuric acid is added at a liquid-to-solid ratio of 5:1, and the mixture is stirred and leached at room temperature for 60 minutes.

[0031] The present invention has the following beneficial effects:

[0032] (1) Waste polypropylene is used as an endogenous reducing agent to replace expensive chemical reducing agents (such as NaHSO3) and high-concentration strong acids in traditional wet processes, which greatly reduces acid and alkali consumption, avoids the generation of a large amount of difficult-to-treat acidic saline wastewater, and achieves green and clean production.

[0033] (2) By selective water leaching, lithium can be preferentially and efficiently leached out, effectively avoiding the lithium entrainment loss caused by traditional wet multi-metal simultaneous leaching and the problem of lithium volatilization at high temperature in pyrometallurgical processes.

[0034] (3) By utilizing the magnetic differences of the products after roasting, the magnetic slag rich in Ni / Co / Mn and the non-magnetic slag mainly composed of graphite are separated in one step by wet magnetic separation, simplifying the process. Compared with the limitations of traditional wet leaching solution which requires complex impurity removal due to the coexistence of multiple metals or direct regeneration which has high requirements for the consistency of raw materials, this process is more adaptable and more flexible in operation.

[0035] (4) After the magnetic slag is mildly acid-leached, the P204 / P507 stepwise extraction system is used to achieve efficient stepwise recovery of Mn–Co–Ni under precise pH control. The process is short, the amount of reagent used is small, and the product purity is high, which overcomes the problems of complex operation and cross-contamination caused by traditional wet multi-stage precipitation or extraction.

[0036] (5) The overall roasting temperature is controlled in the medium temperature range of 600~800°C, which is much lower than that of pyrometallurgical smelting (>1400°C), resulting in significant energy savings; at the same time, it avoids the high-temperature decomposition of fluorine-containing binders such as PVDF to produce toxic gases such as HF and dioxins, thus ensuring high environmental safety.

[0037] (6) This process combines the advantages of pyrometallurgical pretreatment and hydrometallurgical refining to build a low-consumption, low-emission, and high-value recycling path for waste ternary lithium batteries, producing high-grade metal-rich aggregates or battery-grade salt products, which has both technical feasibility and industrialization prospects. Attached Figure Description

[0038] To more clearly illustrate the specific embodiments of this disclosure, the accompanying drawings used in the specific embodiments will be briefly described below. In all the drawings, the elements or parts are not necessarily drawn to scale.

[0039] Figure 1 This is a process flow diagram of the present invention. Detailed Implementation

[0040] The embodiments of the technical solutions disclosed herein will now be described in detail with reference to the accompanying drawings. These embodiments are merely illustrative of the technical solutions disclosed herein and are therefore intended to limit the scope of protection of this disclosure.

[0041] It should be noted that, unless otherwise stated, the technical or scientific terms used in this application should have the ordinary meaning as understood by one of ordinary skill in the art to which this invention pertains.

[0042] Unless otherwise specified, all test materials and reagents used in the following examples are commercially available.

[0043] Unless otherwise specified in the embodiments, the techniques or conditions described in the literature in this field or in accordance with the product manual may be followed.

[0044] Raw material preparation:

[0045] To ensure safety and achieve efficient recycling, the spent lithium batteries are first immersed in a 5 wt.% NaCl solution for 24 hours to completely discharge them and eliminate the risk of short circuits or thermal runaway. The batteries are then cleaned and air-dried. The outer casing is removed using mechanical cutting tools (such as scissors), the battery cells are extracted, and the positive electrode, negative electrode (containing graphite), separator, and residual electrolyte are separated. The obtained positive electrode (composed of an NCM active coating and an aluminum foil substrate) is collected and cut into uniformly sized sheet materials.

[0046] To remove the organic components from the NCM active coating, the sheet-like positive electrode material was heat-treated at 400°C for 2 hours to fully decompose the binder polyvinylidene fluoride (PVDF) and residual electrolyte, achieving initial desorption of the NCM coating. Subsequently, the pyrolyzed material was transferred to a container containing a sodium hydroxide solution with a solid-liquid ratio of 1:10 and a sodium hydroxide solution concentration of 5%. Under mechanical stirring, a chemical-mechanical synergistic exfoliation was carried out: the sodium hydroxide solution corroded the oxide layer on the aluminum foil surface, while the PVDF lost its adhesive properties due to pyrolysis, causing the NCM coating to detach from the aluminum foil in the form of black powder and settle to the bottom of the container along with some of the sheet-like aluminum foil.

[0047] The mixed slurry in the above container was filtered, and the resulting solid product was repeatedly washed with deionized water and filtered multiple times until the filtrate was neutral (pH≈7) to remove residual alkali and soluble impurities. The filter cake was then dried in a 100°C vacuum drying oven for 12 hours to completely remove adsorbed water. Finally, the dried material was sieved through a standard sieve to effectively separate residual aluminum foil fragments, obtaining a high-purity black powder (mainly composed of NCM positive electrode active material).

[0048] Example 1

[0049] like Figure 1 As shown, this invention discloses an efficient method for separating and recycling retired ternary lithium battery cathode materials, comprising the following steps:

[0050] (1) Weigh the raw materials

[0051] Weigh the dried black powder (mainly NCM positive electrode active material) and polypropylene powder at a mass ratio (blending ratio) of 1:1; the polypropylene powder is obtained by crushing waste polypropylene (such as post-consumer waste plastic products); the particle size of both the black powder and the polypropylene powder is less than 300μm.

[0052] (2) Mixing and grinding

[0053] The total mass of the raw materials was weighed to a 1:5 ratio with the mass of the agate grinding balls. The black powder, polypropylene powder, and agate grinding balls were then loaded into an agate ball mill jar. The jar was then placed in a ball mill for mechanical grinding. This process aimed to achieve uniform mixing and particle refinement of the cathode material and polypropylene powder, thereby strengthening the contact interface between the solid phases and providing a good physical basis for subsequent reactions. The ball mill speed was 300 r / min, and the grinding time was 30 min.

[0054] (3) Calcination and reduction treatment

[0055] The mixed and ground materials were transferred to a quartz crucible, which was then placed in a tube furnace and heated to 700°C at a heating rate of 10°C / min under an argon protective atmosphere. The mixture was then kept at this calcination temperature for 1.5 hours. After calcination, the heating program of the tube furnace was turned off, and the material was allowed to cool naturally to room temperature to obtain the calcined product. Polypropylene pyrolysis at high temperatures mainly produces low-carbon olefin gases, primarily ethylene, propylene, and butene, along with alkanes such as methane, ethane, and propane, and a small amount of hydrogen. Hydrogen is the most valuable strong reducing agent and can be directly used for the reduction of metal oxides. Olefins and alkanes also possess reducing capabilities at high temperatures and can participate in specific chemical reactions as reducing agents. Therefore, the entire pyrolysis gaseous product can be utilized as a mixed reducing gas resource. Under the action of olefins, alkanes, and hydrogen, Ni in the black powder is... 3+ / Ni 4+ Co 3+ / Co 4+ Mn 4+ Reduced to soluble Ni 2+ Co 2+ Mn 2+ This enables highly selective leaching of the target metal.

[0056] (4) Magnetic separation-water immersion separation

[0057] The calcined product was ground in a vibratory mill for 10 minutes to refine it thoroughly. Then, deionized water was added at a solid-liquid ratio of 1:(5~25) and the mixture was stirred in a constant temperature water bath at 20~80℃ for 20~100 minutes. After the reaction was completed, the resulting mixture was transferred to a wet magnetic separator and separated under a magnetic field strength of 300mT. The separation process yielded three phases: the magnetic slag was rich in nickel, cobalt and manganese, the non-magnetic slag was mainly graphite and was collected and dried, and the liquid phase was a lithium-containing leachate.

[0058] (5) Lithium extraction from leaching solution (preparation of lithium carbonate)

[0059] The lithium-containing leachate was vacuum filtered to remove suspended impurities. The resulting clarified filtrate was then transferred to a rotary evaporator and concentrated to 1 / 5 of its original volume at 60°C. Next, the concentrated solution was slowly added to a saturated Na₂CO₃ solution at a lithium ion to sodium carbonate molar ratio of 1:1, with continuous stirring for 30 minutes to allow the lithium to precipitate. + The mixture was fully precipitated as Li2CO3. After the reaction was completed, the resulting solid-liquid mixture was vacuum filtered. The precipitate was repeatedly washed with deionized water and filtered multiple times until no impurity ions remained. Finally, the washed precipitate was dried at 100℃ for 2 hours to obtain high-purity lithium carbonate.

[0060] (6) Recycling of valuable metals from magnetic slag

[0061] Acid leaching: After drying the magnetic slag, add 3.5 mol / L dilute sulfuric acid at a liquid-to-solid ratio of 5:1, and stir at room temperature for 60 minutes to selectively dissolve nickel, cobalt, and manganese. The resulting acid leaching solution (containing Ni) is obtained by vacuum filtration. 2+ Co 2+ Mn 2+ );

[0062] NiO + H₂SO₄ = NiSO₄ + H₂O

[0063] CoO + H₂SO₄ = CoSO₄ + H₂O

[0064] Ni + H₂SO₄ = NiSO₄ + H₂

[0065] Co + H₂SO₄ = CoSO₄ + H₂

[0066] MnO + H₂SO₄ = MnSO₄ + H₂O

[0067] Iron removal by precipitation: NaOH solution is then slowly added dropwise to the acid leaching solution to adjust the pH to 3.5-4.5, and the mixture is stirred thoroughly for 30 minutes to promote the hydrolysis of iron impurities to form ferric hydroxide precipitate, which is then removed by filtration.

[0068] FeO + H₂SO₄ = FeSO₄ + H₂O (Reaction of iron impurities during acid leaching)

[0069] Fe + H₂SO₄ = FeSO₄ + H₂ (Reaction of iron impurities during acid leaching)

[0070] 4Fe 2+ +O₂ + 10H₂O = 4Fe(OH)₃ + 8H₂O +

[0071] Stepwise extraction: First-stage extraction is performed by adding P204 (di(2-ethylhexyl)phosphoric acid) extractant to the iron-removed solution to selectively separate manganese. The resulting manganese-loaded organic phase is washed and back-extracted to obtain a manganese salt solution. Subsequently, P507 (2-ethylhexylphosphonic acid mono-2-ethylhexyl ester) extractant is added to the manganese-extracted aqueous phase to utilize P507 for the extraction of Co. 2+ By leveraging the preferential complexing ability of the solution, the pH of the solution is further adjusted to 3.5–4.0, allowing cobalt to be selectively extracted while nickel remains in the aqueous phase. The cobalt-loaded organic phase is then washed and back-extracted to obtain a cobalt salt solution. Finally, high-purity manganese sulfate, cobalt sulfate, and nickel sulfate solution in the aqueous phase are obtained, all of which can be used to prepare corresponding high-purity metal salt products through crystallization or precipitation.

[0072] Stepwise extraction and the back-extraction of manganese and cobalt are existing technologies and will not be elaborated here.

[0073] The preparation of high-purity metal salt products through crystallization or precipitation is detailed below:

[0074] For sulfate systems, high-purity nickel sulfate heptahydrate (NiSO4·7H2O), cobalt sulfate heptahydrate (CoSO4·7H2O), and manganese sulfate monohydrate (MnSO4·7H2O) crystals can usually be obtained by evaporation concentration followed by cooling crystallization.

[0075] If carbonate or hydroxide precursors are required, a saturated Na₂CO₃ or NaOH solution is slowly added to a purified metal salt solution (manganese sulfate / cobalt sulfate / nickel sulfate). Under strict control of pH, temperature, and stirring rate, the target metal is uniformly precipitated as a carbonate (such as NiCO₃, CoCO₃, MnCO₃) or hydroxide (such as Ni(OH)₂, Co(OH)₂, Mn(OH)₂). After solid-liquid separation, deionized water washing, and vacuum drying, a high-purity metal salt product meeting the standards for battery materials or chemical raw materials is finally obtained. The precipitation process is existing technology and will not be described in detail here.

[0076] In stepwise extraction, to achieve efficient and selective separation of manganese and nickel, P2O4-based methods were used for the separation of Mn. 2+With stronger affinity, the pH of the solution after iron removal is adjusted to 3.0–3.5, at which point the separation coefficient (D) of manganese and nickel is higher. Mn / D Ni The figure reached 10.7.

[0077] Example 2

[0078] The difference from Example 1 is that the mass ratio (blending ratio) of black powder (mainly composed of NCM positive electrode active material) to polypropylene powder is 1:0.5.

[0079] Example 3

[0080] The difference from Example 1 is that the mass ratio (blending ratio) of black powder (mainly composed of NCM positive electrode active material) to polypropylene powder is 1:1.5.

[0081] Example 4

[0082] The difference from Example 1 is that, in the calcination reduction process, the mixed and ground materials are transferred to a quartz crucible, which is then placed in a tube furnace and heated to 600°C at a heating rate of 10°C / min under an argon protective atmosphere, and the reaction is carried out at this calcination temperature for 1.5 hours.

[0083] Example 5

[0084] The difference from Example 1 is that, in the calcination reduction process, the mixed and ground materials are transferred to a quartz crucible, which is then placed in a tube furnace and heated to 800°C at a heating rate of 10°C / min under an argon protective atmosphere, and the reaction is carried out at this calcination temperature for 1.5 hours.

[0085] Example 6

[0086] The difference from Example 1 is that, in the calcination reduction process, the mixed and ground materials are transferred to a quartz crucible, which is then placed in a tube furnace and heated to 700°C at a heating rate of 10°C / min under an argon protective atmosphere, and the reaction is carried out at this calcination temperature for 1 hour.

[0087] Example 7

[0088] The difference from Example 1 is that in the calcination reduction process, the mixed and ground materials are transferred to a quartz crucible, which is then placed in a tube furnace and heated to 700°C at a heating rate of 10°C / min under an argon protective atmosphere, and the reaction is carried out at this calcination temperature for 2 hours.

[0089] Comparative Example 1

[0090] The difference from Example 1 is that the mass ratio (blending ratio) of black powder (mainly composed of NCM positive electrode active material) to polypropylene powder is 1:0.2.

[0091] Comparative Example 2

[0092] The difference from Example 1 is that the mass ratio (blending ratio) of black powder (mainly composed of NCM positive electrode active material) to polypropylene powder is 1:2.5.

[0093] Comparative Example 3

[0094] The difference from Example 1 is that, in the calcination reduction process, the mixed and ground materials are transferred to a quartz crucible, which is then placed in a tube furnace and heated to 400°C at a heating rate of 10°C / min under an argon protective atmosphere, and the reaction is carried out at this calcination temperature for 1.5 hours.

[0095] Comparative Example 4

[0096] The difference from Example 1 is that in the calcination reduction process, the mixed and ground materials are transferred to a quartz crucible, which is then placed in a tube furnace and heated to 900°C at a heating rate of 10°C / min under an argon protective atmosphere, and the reaction is carried out at this calcination temperature for 1.5 hours.

[0097] Comparative Example 5

[0098] The difference from Example 1 is that, in the calcination reduction process, the mixed and ground materials are transferred to a quartz crucible, which is then placed in a tube furnace and heated to 700°C at a heating rate of 10°C / min under an argon protective atmosphere, and the reaction is carried out at this calcination temperature for 0.5 h.

[0099] Comparative Example 6

[0100] The difference from Example 1 is that in the calcination reduction process, the mixed and ground materials are transferred to a quartz crucible, which is then placed in a tube furnace and heated to 700°C at a heating rate of 10°C / min under an argon protective atmosphere, and the reaction is carried out at this calcination temperature for 3 hours.

[0101] Comparative Example 7

[0102] The difference from Example 1 is that no polypropylene was added.

[0103] Comparative Example 8

[0104] The difference from Example 1 is that a wet magnetic separator was not used for magnetic separation.

[0105] Comparative Example 9

[0106] The traditional hydrometallurgical method for recycling nickel, cobalt, manganese, and lithium from waste ternary lithium batteries typically includes the following steps: First, the disassembled cathode material is co-leached with acid (such as sulfuric acid or hydrochloric acid) and reducing agent (such as H2O2 or NaHSO3) under heating conditions, allowing metals such as Ni, Co, Mn, and Li to dissolve simultaneously into the solution; then, impurities (such as Fe) are removed sequentially and the target metals are separated through multi-stage pH-controlled precipitation or solvent extraction—usually, iron and aluminum are removed first by precipitation, and then Mn, Co, and Ni are separated by stepwise extraction using extractants such as P2O4 and P5O7, while Li is enriched in the final mother liquor; finally, each metal solution is back-extracted, crystallized, or precipitated to obtain sulfate or carbonate products.

[0107] In Examples 1-7 and Comparative Examples 1-9, the amount of black powder / cathode material added was 1 kg.

[0108] In the magnetic separation-water leaching separation step, a solid-liquid ratio of 1:(5~25) is acceptable, the water bath temperature of the constant temperature water bath is acceptable from 20 to 80°C, and the stirring reaction time is acceptable from 20 to 100 min. In Examples 1-7 and Comparative Examples 1-8, a solid-liquid ratio of 1:20 was selected, the water bath temperature of the constant temperature water bath was 20~30°C, and the stirring reaction time was 60 min.

[0109] The black powders involved in Examples 1-7 and Comparative Examples 1-9 are all high-purity black powders (mainly NCM positive electrode active material) obtained after preparation from the aforementioned raw materials.

[0110] The specific parameters of Examples 1-7 and Comparative Examples 1-9 of the present invention are shown in Table 1.

[0111] Table 1

[0112]

[0113] Performance testing

[0114] (1) Calculation of leaching rate of nickel, cobalt, manganese and lithium

[0115] Let the mass (dry) of the roasted product / magnetic slag be W (kg), the metal grade be R (%), and the mass (dry) of the leaching residue be W. r (kg), slag grade R r (%), then the metal leaching rate η is:

[0116]

[0117] Metal content was tested using ICP-MS (inductively coupled plasma mass spectrometry).

[0118] The lithium leaching rate was calculated based on the roasted product, and the leaching residue consisted of magnetic and non-magnetic residues obtained after water leaching.

[0119] The leaching rates of nickel, cobalt, and manganese are calculated based on the magnetic slag obtained by magnetic separation enrichment. The leaching slag is acid leaching slag, that is, acid leaching filter slag obtained after sufficient acid leaching and filtration.

[0120] (2) Lithium carbonate purity (LiP / %): determined according to GB / T 11064.1-2024 "Chemical analysis methods for lithium carbonate, lithium hydroxide monohydrate and lithium chloride - Part 1: Determination of lithium carbonate content by titration";

[0121] (3) Nickel mass fraction (Ni wt%) in nickel sulfate: determined according to GB / T 26524-2023 "Refined Nickel Sulfate" standard;

[0122] (4) Cobalt mass fraction (Co wt%) in cobalt sulfate: determined according to GB / T 26523-2022 "Refined Cobalt Sulfate" standard;

[0123] Test results: The performance test results are shown in Table 2.

[0124] Table 2

[0125]

[0126] According to the data in Table 1, Examples 1-7 of this invention demonstrate that the various process parameters work together to achieve a synergistic effect, with the absence of any one parameter weakening the effect. Waste polypropylene is used as a green reducing agent, and selective release of lithium and solid-phase enrichment of nickel, cobalt, and manganese in retired ternary lithium battery cathode materials are achieved through medium-low temperature (600-800°C) reduction roasting coupled with mechanical activation. This avoids the dependence on strong acids and high-valent chemical reducing agents in traditional wet processes. Combined with one-step wet magnetic separation, lithium-containing leachate, graphite non-magnetic slag, and high-grade nickel-cobalt-manganese magnetic slag are efficiently obtained, significantly simplifying the process. Lithium is recovered in the form of high-purity Li₂CO₃ (precipitation rate >99%). The magnetic slag undergoes mild acid leaching and P₂O₄ / P₅O₇ stepwise extraction to achieve highly selective separation of Ni, Co, and Mn. The leaching rates of Ni, Co, and Mn are higher than 90%, and the leaching rate of Li is higher than 88%, with product purity meeting battery material requirements. The entire process consumes few reagents, produces no high-salt wastewater, has low energy consumption, and achieves a high degree of resource utilization of waste, combining economic efficiency with environmental friendliness.

[0127] Analysis of Example 1 and Comparative Examples 1-2 shows that when the amount of polypropylene added is too low, a sufficient reducing atmosphere cannot be provided during the calcination process, resulting in poor reduction effect during calcination and high-valence metal ions (such as Ni) in the black powder. 3+ Co 3+ Mn 4+ It failed to be fully reduced to a soluble low-valence state (such as Ni). 2+ Co2+ Mn 2+ This leads to a significant decrease in the leaching rates of nickel, cobalt, manganese, and lithium in subsequent leaching stages. Simultaneously, lithium leaching is incomplete due to the difficulty in effectively "loosening" its crystal structure, resulting in a relative enrichment of residual impurities (such as Fe) in the leachate. Although these residual impurities are precipitated, they may still be carried into the lithium carbonate product, reducing its purity. More importantly, due to the low Ni / Co / Mn content and high impurity content in the magnetic slag, the concentration of metal ions in the acid leaching solution is low. Even after P204 / P507 staged extraction, the target metal mass fraction in the cobalt sulfate and nickel sulfate solutions obtained from back-extraction is significantly reduced, affecting the purity of the subsequent crystallized products. When the amount of polypropylene added is too high, although a strong reducing atmosphere can be provided during the roasting process, the intense reducing roasting can easily lead to excessive sintering and agglomeration of particles, and generate a large amount of carbonaceous residue that encapsulates unreacted active components. At the same time, nickel and cobalt may be excessively reduced to dense metallic phases or alloys, manganese forms insoluble low-valence oxides, and lithium is also trapped inside the sintered body. These deteriorations in physical structure significantly hinder the effective contact between water or acid and the target metal, resulting in a decrease in the leaching rates of nickel, cobalt, manganese, and lithium. In addition, carbon residues that are not completely removed in subsequent water leaching, magnetic separation, or washing processes enter the acid leaching solution, which will interfere with the stability of the extraction system and introduce organic or inorganic impurities into the back-extraction solution, slightly reducing the mass fraction of cobalt and nickel in the cobalt sulfate and nickel sulfate products.

[0128] Analysis of Example 1 and Comparative Examples 3-4 shows that when the calcination temperature is too low, the polypropylene pyrolysis is incomplete, the reducing atmosphere is weak, and the high-valence metal ions (such as Ni) in the black powder are concentrated. 3+ Co 3+ Mn 4+ The leaching process failed to effectively reduce the slag, resulting in a highly stable crystal structure and significantly lower leaching rates for nickel, cobalt, manganese, and lithium during subsequent leaching. Furthermore, the increased proportion of impurity ions (such as Fe) in the relatively concentrated system led to their easy co-precipitation into the product, reducing the purity of lithium carbonate (<98 wt%). Simultaneously, the insufficient enrichment of nickel, cobalt, and manganese in the magnetic slag, coupled with high impurity content and low metal concentration in the acid leaching solution, meant that even after P204 / P507 staged extraction, the target metal concentration in the back-extraction solution remained limited. Ultimately, this resulted in a significant decrease in the mass fraction of cobalt and nickel in the cobalt sulfate and nickel sulfate products. When the calcination temperature is too high, although the reduction is sufficient, the high temperature can easily cause significant sintering or even local melting, resulting in particle densification and lithium encapsulation, which leads to a decrease in lithium leaching rate. Although nickel, cobalt, and manganese are mostly reduced to metals or low-valence oxides, the acid leaching kinetics are hindered due to sintering and agglomeration, and the leaching rate also decreases to varying degrees. In addition, volatile or entrained impurities contaminate the lithium precipitation process and affect the purity of lithium carbonate. Although the Ni / Co grade in the magnetic slag is high, the agglomerated phase is not completely dissolved, and the actual effective metal concentration in the acid leaching solution is limited. After back-extraction, the mass fraction of cobalt and nickel in cobalt sulfate and nickel sulfate does not increase significantly, and may even decrease slightly due to the co-dissolution of impurities.

[0129] Analysis of Example 1 and Comparative Examples 5-6 shows that when the calcination reaction time is too short, the polypropylene pyrolysis is insufficient, and the reducing atmosphere fails to continuously act on the entire material system, resulting in high-valence metal ions (such as Ni) in the black powder. 3+ Co 3+ Mn 4 + Only a portion of the lithium is reduced, and the lattice structure is not completely loosened. Therefore, lithium is difficult to release completely during subsequent water or acid leaching, resulting in a low leaching rate. The leaching rates of nickel, cobalt, and manganese are also significantly limited because they are not fully converted into soluble low-valence states. The low lithium leaching rate reduces the amount of lithium carbonate precipitate, and impurity ions are relatively enriched, making them easy to be carried into the product and reducing the purity of lithium carbonate. At the same time, the enrichment of nickel, cobalt, and manganese in the magnetic slag is insufficient, the concentration of metal ions in the acid leaching solution is low, and the mass fraction of cobalt and nickel in the cobalt sulfate and nickel sulfate solutions obtained after extraction and back-extraction is low. When the calcination reaction time is too long, although the initial reduction is sufficient, prolonged high-temperature holding can cause excessive sintering of particles, grain coarsening, or even local melting, forming a dense structure that encapsulates unreacted components, resulting in a decrease in lithium leaching rate. Although nickel and cobalt have been reduced to metals or low-valence oxides, severe agglomeration increases the resistance to acid leaching mass transfer, leading to a significant drop in leaching rate. Manganese is more likely to form insoluble phases due to prolonged exposure to a reducing environment, which may significantly reduce the leaching rate. In addition, volatile impurities or carbon residues contaminate the lithium precipitation system, slightly affecting the purity of lithium carbonate. Although the Ni / Co content in the magnetic slag is high, the agglomerated phase is not completely dissolved, resulting in a limited increase in the effective metal concentration in the back-extraction solution. The increase in the mass fraction of cobalt and nickel in cobalt sulfate and nickel sulfate slows down, and may even decrease slightly due to the accumulation of impurities.

[0130] Analysis of Example 1 and Comparative Example 7 shows that when polypropylene is not added, the calcination process lacks an endogenous reducing agent and cannot provide an effective reducing atmosphere, resulting in high-valence metal ions (such as Ni) in the black powder. 3+ Co 3+ Mn 4+ The lithium carbonate is almost unreducible, and its lattice structure remains highly stable, making it difficult for lithium ions to be released from the layered structure. Therefore, during subsequent water leaching or weak acid leaching, the leaching rate of lithium is significantly reduced, while the leaching rates of nickel, cobalt, and manganese are also extremely low because they still exist in the form of insoluble high-valence oxides. Low lithium recovery not only reduces the production of lithium carbonate, but also makes impurities (such as Fe) in the leaching solution relatively enriched and co-precipitated into the product, which seriously reduces the purity of lithium carbonate. At the same time, because nickel, cobalt, and manganese in the magnetic slag cannot be effectively enriched (or even cannot form a significant magnetic phase), the concentration of metal ions in the resulting acid leaching solution is extremely low. Even after P204 / P507 stepwise extraction, the mass fraction of cobalt and nickel in the cobalt sulfate and nickel sulfate solutions decreases significantly after back-extraction.

[0131] Analysis of Example 1 and Comparative Example 8 shows that when magnetic separation is not performed, the calcined mixed product (containing nickel, cobalt, and manganese oxides, graphite, residual lithium compounds, and conductive agents, etc.) directly enters the acid leaching process, resulting in a large number of non-target components (such as graphite, aluminum foil fragments, carbon black, and binder residues) being treated with acid along with it. These impurities not only consume additional acid but also introduce co-soluble metal ions such as Fe, complicating the composition of the acid leaching solution. Although the total leaching rate of nickel, cobalt, manganese, and lithium may not decrease significantly, the concentration of impurities in the leaching solution increases dramatically, seriously interfering with subsequent separation. During the precipitation iron removal and stepwise extraction processes, high concentrations of impurities compete for the extractant and reduce selectivity, leading to a decrease in the separation efficiency of the P204 / P507 system, and a significant reduction in the mass fractions of cobalt and nickel in the back-extraction cobalt sulfate and nickel sulfate solutions. Simultaneously, lithium leaching solutions containing impurities are prone to co-precipitation during the lithium carbonate precipitation stage, resulting in a decrease in the purity of lithium carbonate.

[0132] Analysis of Example 1 and Comparative Example 9 shows that while the traditional hydrometallurgical process (i.e., direct acid leaching + chemical reducing agent, omitting polypropylene-assisted roasting and magnetic separation enrichment) can achieve high simultaneous leaching rates of nickel, cobalt, manganese, and lithium under the action of strong acid (e.g., 2 mol / L H2SO4) and reducing agent (e.g., H2O2, NaHSO3), all metals and impurities (Fe, graphite, binders, etc.) simultaneously enter the leachate, resulting in a highly complex system composition. Subsequent separation requires multi-stage pH-controlled precipitation and solvent extraction, a lengthy process with high reagent consumption. During this process, some impurities are prone to co-precipitation during the lithium precipitation stage, reducing the purity of lithium carbonate. Simultaneously, due to the high impurity concentration in the initial leachate, the selectivity of the P204 / P507 extraction system is limited. Although the mass fractions of cobalt and nickel in the back-extraction cobalt sulfate and nickel sulfate solutions can reach industrial-grade levels, batch stability is poor, and trace amounts of Na are often present. + SO4 2- The presence of organic entrainment makes it difficult to consistently meet the requirements for high-purity battery materials (e.g., Co wt% ≥ 20.5%, Ni wt% ≥ 22%). Furthermore, this method generates a large amount of hydrochloric acid wastewater, imposing a heavy environmental burden, and its overall economic efficiency and environmental friendliness are significantly lower than the technical solution of this invention.

[0133] The above embodiments are only used to illustrate the technical solutions of this disclosure, and are not intended to limit it. Although this disclosure has been described in detail with reference to the foregoing embodiments, those skilled in the art should understand that modifications can still be made to the technical solutions described in the foregoing embodiments, or equivalent substitutions can be made to some or all of the technical features therein. Such modifications or substitutions do not cause the essence of the corresponding technical solutions to deviate from the scope of the technical solutions of the embodiments of this disclosure, and they should all be covered within the scope of the claims and specification of this disclosure.

Claims

1. A method for the efficient separation and recycling of cathode materials from retired ternary lithium batteries, characterized in that, Includes the following steps: (1) Weigh the raw materials Weigh the dried black powder and polypropylene powder at a mass ratio of 1:(0.5~1.5); (2) Mixing and grinding Black powder, polypropylene powder, and grinding balls are loaded into an agate ball mill jar for mechanical ball milling. (3) Calcination and reduction treatment The mixed and ground materials are placed in a quartz crucible and calcined in a tube furnace at 600-800°C for 1-2 hours under argon protection. The furnace is then cooled to room temperature to obtain the calcined product. (4) Magnetic separation-water immersion separation The calcined product was refined by vibration grinding, then deionized water was added, and the mixture was stirred and reacted in a constant temperature water bath. The resulting mixture was separated by wet magnetic separation to obtain magnetic slag rich in nickel, cobalt and manganese, non-magnetic slag mainly composed of graphite, and lithium-containing leachate. (5) Lithium extraction from leachate The lithium-containing leachate was vacuum filtered to remove impurities. The filtrate was concentrated to 1 / 5 of its original volume, and then a saturated Na₂CO₃ solution was slowly added while stirring to allow the lithium to dissolve. + The solid-liquid mixture was fully precipitated as Li2CO3; the precipitate was filtered under vacuum, and the precipitate was repeatedly washed with deionized water until no impurity ions were present. Then it was dried at 100°C for 2 hours to obtain high-purity lithium carbonate. (6) Recycling of valuable metals from magnetic slag Acid leaching: Add an appropriate amount of dilute sulfuric acid to the dried magnetic slag at a certain liquid-solid ratio, stir thoroughly at room temperature to dissolve Ni, Co, and Mn, and then filter under vacuum to obtain Ni-containing... 2+ Co 2+ Mn 2+ Acid leaching solution; Iron precipitation: NaOH solution was then added dropwise to adjust the pH to a suitable range, and the mixture was stirred thoroughly to hydrolyze the iron impurities and form ferric hydroxide precipitate. Iron was then removed by filtration. Stepwise extraction: The iron-removing solution is first extracted with P204 at pH 3.0–3.5 to separate manganese. The loaded organic phase is washed and back-extracted to obtain a manganese sulfate solution. After manganese extraction, the aqueous phase is selectively extracted with P507 at pH 3.5–4.

0. After washing and back-extraction, a cobalt sulfate solution is obtained. Nickel is enriched in the final aqueous phase. The obtained manganese sulfate, cobalt sulfate and nickel sulfate solutions are all crystallized or precipitated to obtain high-purity metal salts.

2. The efficient separation and recycling method for retired ternary lithium battery cathode materials according to claim 1, characterized in that: In the aforementioned mixed grinding process, black powder, polypropylene powder, and grinding balls are loaded into an agate ball mill jar at a ratio of 1:5 (total mass of raw materials to agate grinding balls) for mechanical ball milling.

3. The efficient separation and recycling method for retired ternary lithium battery cathode materials according to claim 1 or 2, characterized in that: In the mixed grinding, the ball mill speed is 300 r / min and the grinding time is 30 min.

4. The efficient separation and recycling method for retired ternary lithium battery cathode materials according to claim 1, characterized in that: In the calcination and reduction process, the mixed and ground materials are transferred to a quartz crucible, which is then placed in a tube furnace and heated to 600-800°C at a heating rate of 10°C / min under an argon protective atmosphere.

5. The efficient separation and recycling method for retired ternary lithium battery cathode materials according to claim 1, characterized in that: In the magnetic separation-water immersion separation, the roasted product is ground for 10 minutes using a vibratory mill.

6. The efficient separation and recycling method for retired ternary lithium battery cathode materials according to claim 1, characterized in that: In the magnetic separation-water leaching process, the ground calcined product and deionized water are mixed at a solid-liquid ratio of 1:(5~25), and then placed in a constant temperature water bath at 20~80℃ and stirred for 20~100 min.

7. The efficient separation and recycling method for retired ternary lithium battery cathode materials according to claim 1, characterized in that: In the aforementioned magnetic separation-water immersion separation, the magnetic field strength of the wet magnetic separator is 300 mT.

8. The efficient separation and recycling method for retired ternary lithium battery cathode materials according to claim 1, characterized in that: In the lithium extraction from the leaching solution, the clarified filtrate obtained by vacuum filtration is transferred to a rotary evaporator and concentrated to 1 / 5 of its original volume at 60°C.

9. The efficient separation and recycling method for retired ternary lithium battery cathode materials according to claim 1, characterized in that: In the lithium extraction process from the leaching solution, the concentrated solution was slowly added to a saturated Na₂CO₃ solution at a lithium ion to sodium carbonate molar ratio of 1:1, and stirred continuously for 30 minutes to allow the lithium to precipitate. + Fully precipitated as Li2CO 3。 10. The efficient separation and recycling method for retired ternary lithium battery cathode materials according to claim 1, characterized in that: In the recovery of valuable metals from the magnetic slag, after the magnetic slag is dried, 3.5 mol / L dilute sulfuric acid is added at a liquid-to-solid ratio of 5:1, and the mixture is stirred and leached at room temperature for 60 minutes.

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