Method for synergistically recovering valuable metals in waste lithium batteries and electrolytic aluminum waste residues

Abstract

The invention belongs to the technical field of resource recovery, and provides a method for synergistically recovering valuable metals in waste lithium batteries and electrolytic aluminum waste residues, which comprises the following steps: slurrying, alkaline leaching, acid leaching, precipitating, washing, separating and the like, and synergistically treating the electrolytic aluminum waste residues and waste lithium iron phosphate battery electrode materials. Valuable substances such as calcium fluoride and lithium phosphate are efficiently recycled. The process has the core characteristics of cooperative treatment of two industrial wastes, reduction of use of acid and alkali, cyclic utilization of fluorine precipitation liquid and alkaline lithium liquid, and efficient recovery of lithium. Due to the characteristics, the process has remarkable advantages in the aspects of comprehensive utilization of resources, environmental protection benefits and economic benefits, and an innovative solution is provided for treatment of industrial wastes. By further optimizing process parameters and equipment, the feasibility and the application prospect of the method can be further improved.

Description

Technical Field

[0001] The present invention belongs to the technical field of resource recovery, and particularly relates to a method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum slag. Background Art

[0002] With the continuous development of industrial production, the treatment of industrial wastes such as aluminum slag and lithium iron phosphate black powder has become increasingly prominent. Aluminum slag is a solid waste generated during the production of electrolytic aluminum, and its main components are aluminum oxide (Al2O3), aluminum fluoride (AlF3), sodium fluoride (NaF), and a small amount of lithium oxide (Li2O). Aluminum slag contains a large amount of fluorides and lithium, and has high recycling value. Traditional treatment methods usually cannot effectively recover the valuable substances therein, resulting in waste of resources and environmental pollution. Traditional treatment methods often cannot effectively recover the valuable substances therein, resulting in waste of resources and environmental pollution. Lithium iron phosphate black powder is a pulverized product of the cathode material of waste lithium-ion batteries, and its main components are lithium iron phosphate (LiFePO4), graphite, and a small amount of metal impurities (such as copper, aluminum, etc.). Lithium iron phosphate black powder contains valuable elements such as lithium, iron, and phosphorus, and has high recycling value. Existing recycling processes usually adopt acid leaching or pyrometallurgical treatment, but these methods have problems such as high energy consumption, serious equipment corrosion, and a large amount of impurity introduction.

[0003] The traditional treatment process for waste lithium iron phosphate batteries has the following problems: (1) The main goal is to recover valuable elements such as lithium, iron, and phosphorus, and the economic benefit is relatively high, but the treatment cost is also relatively high; (2) Usually, the acid leaching method is used to extract lithium, but other components (such as aluminum, iron, and fluorine) may be ignored or treated as waste slag, and the recycling value is not fully utilized; (3) The acidic waste liquid and waste gas generated during the acid leaching process need to be strictly treated; (4) The waste liquid and waste slag generated during the treatment process need to be treated additionally, increasing the complexity; (5) The electrolyte and heavy metals in waste lithium batteries may cause environmental pollution.

[0004] The traditional treatment process for electrolytic aluminum slag has the following problems: (1) The fluorides and alkaline substances in the slag may cause environmental pollution; (2) The waste liquid and waste gas generated during the treatment process need to be treated additionally, increasing the environmental protection pressure; (3) The main recovery is aluminum, and the economic benefit is limited; (4) Treating harmful components such as fluorides increases the cost.

[0005] Therefore, developing an efficient and environmentally friendly co-recovery process for the two has important economic and environmental significance.

[0006] The information disclosed in this background art section is only intended to enhance the overall understanding of the present invention and should not be regarded as an admission or any form of suggestion that this information constitutes prior art already known to those of ordinary skill in the art. Summary of the Invention

[0007] The present invention is made to solve the above problems, and aims to provide a method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum slag. Through a series of chemical treatment and separation steps, efficient recovery of valuable substances such as calcium fluoride, lithium hydroxide, and lithium phosphate is achieved.

[0008] The present invention provides a method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum slag, having the following characteristics, including the following steps: Step S1, mixing electrolytic aluminum slag with water at a mass ratio of 1:(1 - 5) to form a slurried mixture, i.e., a slurry;

[0009] Step S2, adding a sodium hydroxide solution to the slurry for alkali leaching, and heating the reaction to obtain an alkali leaching solution;

[0010] Step S3, adding sodium hydroxide to the alkali leaching solution, maintaining the pH value of the system at 10 - 14, and heating the reaction to obtain a fluorine leaching solution and a fluorine leaching residue;

[0011] Step S4, according to the molar amount of substances, adding calcium hydroxide with a fluorine amount 1.5 - 2.0 times that of the fluorine leaching solution to the fluorine leaching solution, heating the reaction to generate a crude calcium fluoride precipitate and a fluorine precipitation solution 1; and 10% - 50% of the fluorine precipitation solution 1 returns to Step S1 to participate in slurrying, 10% - 50% returns to Step S2 to participate in alkali leaching, 10% - 50% returns to Step S3 to participate in fluorine leaching, and 20% - 70% is used for subsequent iron precipitation;

[0012] Step S5, adding water with a mass 4 - 6 times that of the obtained crude calcium fluoride, heating for secondary countercurrent washing to obtain pure calcium fluoride and a washing solution 1; and the washing solution 1 returns to Step S3 for fluorine leaching;

[0013] Step S6, adding calcium hydroxide to the fluorine leaching residue obtained in Step S3, heating for secondary countercurrent lithium leaching to obtain an alkaline lithium solution and a lithium leaching residue 1;

[0014] Step S7, adding water to the lithium leaching residue 1 at a mass ratio of 1:(1 - 3), and obtaining a residue and a washing solution 2 after washing; the washing solution 2 returns to be used for secondary countercurrent lithium leaching;

[0015] Step S8, adding water to the waste lithium iron phosphate black powder for flotation to obtain a lithium iron phosphate battery black powder residue and graphite;

[0016] Step S9, adding sulfuric acid and hydrogen peroxide to the lithium iron phosphate battery black powder residue simultaneously, maintaining the pH value of the solution at 1.0 - 3.0, heating the reaction until all iron is converted into ferric ions to obtain an acid leaching solution and an acid leaching residue;

[0017] Step S10, adding the fluorine precipitation solution obtained in Step S4 to the acid leaching residue obtained in Step S9 to obtain an iron precipitation residue and an iron precipitation solution;

[0018] Step S11: Add the alkaline lithium solution obtained in step S6 to the acid leaching solution obtained in step S9 to obtain a lithium leaching solution and lithium leaching residue 2;

[0019] Step S12: Concentrate the lithium leaching solution obtained in step S11 by 3 to 5 times to obtain a lithium leaching mother liquor;

[0020] Step S13: Concentrate and separate lithium phosphate and Glauber's salt from the lithium leaching mother liquor.

[0021] In the method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residue provided by the present invention, it may also have the following characteristics: wherein, it further includes step S0: Crush the electrolytic aluminum waste residue into particles with a diameter less than 5 mm, wash away soluble impurities with water, and then dry to obtain electrolytic aluminum slag for use.

[0022] In the method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residue provided by the present invention, it may also have the following characteristics: wherein, in step S2, the alkali is sodium hydroxide or potassium hydroxide. In the present invention, sodium hydroxide is taken as an example for illustration, and potassium hydroxide can achieve the same technical effect.

[0023] In the method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residue provided by the present invention, it may also have the following characteristics: wherein, in step S2, the alkali is sodium hydroxide, and by mass ratio, electrolytic aluminum slag:sodium hydroxide = 1:(3 ～ 5), heat up to 40 °C ～ 50 °C, keep warm and react for 15 ～ 30 minutes; the mass fraction concentration of the sodium hydroxide solution is 10% ～ 15%.

[0024] In the method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residue provided by the present invention, it may also have the following characteristics: wherein, in step S3, heat up to 40 °C ～ 50 °C, keep warm and react for 0.5 ～ 2 hours; in step S4, heat up to 40 °C - 50 °C, keep warm and react for 0.5 - 2 hours.

[0025] In the method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residue provided by the present invention, it may also have the following characteristics: wherein, in step S5, heat up to 40 °C - 50 °C for secondary countercurrent washing.

[0026] In the method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residue provided by the present invention, it may also have the following characteristics: wherein, in step S6, add a calcium hydroxide solution with a concentration of 0.05 - 0.15 mol / L to the fluorine leaching residue, and the mass ratio of the fluorine leaching residue to calcium hydroxide is 1:(3 - 5), heat up to 60 ～80 °C, keep the reaction at this temperature for 0.5 ～ 3 hours.

[0027] In the method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum slag provided by the present invention, it may further have the following characteristics:

[0028] In the method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum slag provided by the present invention, it may further have the following characteristics: wherein, in step S9, sulfuric acid and hydrogen peroxide are simultaneously added to the black powder slag of lithium iron phosphate batteries, and the pH value of the solution is maintained at 1.0 ～ 3.0, and the temperature is raised to 60 °C ～ 80 °C, according to the molar ratio of substances, the addition amount of sulfuric acid is 3 ～ 6 times the lithium content in the black powder slag of lithium iron phosphate batteries, and the addition amount of hydrogen peroxide is 1.5 ～ 3 times the iron content in the black powder slag of lithium iron phosphate batteries. The concentration of the sulfuric acid solution is 2M.

[0029] In the method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum slag provided by the present invention, it may further have the following characteristics: wherein, it further includes step S14: the solution in step S13 is processed by the MVR process to obtain solid sodium sulfate.

[0030] In the method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum slag provided by the present invention, it may further have the following characteristics: wherein, the evaporation temperature of MVR is 80 - 100 °C, and the evaporation pressure is 0.1 - 0.3 MPa.

[0031] Functions and effects of the invention

[0032] The process of the present invention realizes the efficient recovery of resources and the reduction of waste by co-processing electrolytic aluminum slag and waste lithium iron phosphate battery electrode materials. The basis for co-recovery is as follows: (1) Both contain lithium resources, and the by-products can complement each other; (2) Lithium coupling: LiF in electrolytic aluminum slag and LiFePO4 in the black powder of lithium iron phosphate batteries achieve directional enrichment of lithium through acid-base reactions; (3) Reuse of by-products: The fluorine-precipitating solution (containing NaOH) generated by alkali leaching of electrolytic aluminum slag can replace traditional NaOH for treating the black powder of lithium iron phosphate batteries; the alkaline lithium solution (containing LiOH + NaOH) after acid leaching of the black powder of lithium iron phosphate batteries can be recycled for treating aluminum slag, forming a closed loop.

[0033] Compared with the prior art, this process has at least the following beneficial effects:

[0034] (1) Dual-waste co-processing mechanism

[0035] Through the acid-base coupling reaction of electrolytic aluminum waste residue (containing LiF and Al2O3) and black powder of lithium iron phosphate battery (LiFePO4 and C), the directional enrichment of lithium element and the resource recovery of fluorine are realized. The fluorine-precipitating solution (NaOH) produced by the alkali leaching of electrolytic aluminum slag is reused for the acid leaching of black powder of lithium iron phosphate battery. The fluorine-leached residue obtained after the fluorine leaching of electrolytic aluminum slag is added with calcium hydroxide and then undergoes secondary countercurrent lithium leaching to obtain the alkaline lithium solution (LiOH + NaOH), which is used for the lithium leaching of the acid leaching solution of black powder of lithium iron phosphate battery, forming a cross-process material cycle.

[0036] (2) Acid-base self-supply system

[0037] The fluorine-precipitating solution (mainly NaOH) replaces the initial NaOH addition in the traditional process, and only Ca(OH)2 needs to be supplemented; the alkaline lithium solution (LiOH + NaOH) realizes lithium enrichment through three-stage countercurrent leaching, reducing the NaOH consumption by 60% throughout the process.

[0038] (3) Zero-reagent startup technology

[0039] Only NaOH is initially added to start the alkali leaching of electrolytic aluminum waste residue, and subsequent operations completely rely on the closed-loop cycle of the fluorine-precipitating solution and the alkaline lithium solution, subverting the continuous reagent consumption mode of the traditional process.

[0040] This process realizes the resource closed-loop of electrolytic aluminum slag and black powder of lithium iron phosphate battery through "synergy of double wastes and complementary components", and constructs a four-level collaborative mechanism of "pretreatment - leaching - separation - high-value utilization": after the electrolytic aluminum slag is slurried with water, lithium is dissolved by alkali leaching (+NaOH), and aluminum is converted into soluble sodium aluminate (NaAlO) and sodium fluoride (NaF). The core of the fluorine leaching step is to dissolve the soluble sodium salts (NaAlO2, NaF) and excessive alkali (fluorine leaching solution) in the alkali leaching slurry with a high-concentration NaOH solution, and separate the insoluble LiF solid (fluorine leaching residue); a part of the fluorine precipitation solution generated in the fluorine precipitation step of the electrolytic aluminum slag is returned for slurrying, alkali leaching and fluorine leaching, and the other part is used for iron precipitation of the acid leaching residue of the black powder of lithium iron phosphate battery, synchronously completing the extraction of Li+ and the recovery of fluorine resources (utilization rate > 90%); the crude calcium fluoride is obtained as pure calcium fluoride after secondary countercurrent washing and then sold out; the fluorine leaching residue (LiF) is added with calcium hydroxide and then subjected to secondary countercurrent lithium leaching to obtain a solution containing sodium hydroxide and lithium hydroxide, as well as calcium fluoride (lithium leaching residue 1). The residue obtained after washing the lithium leaching residue 1 is used to make hollow bricks, and the obtained washing solution 2 is returned for secondary countercurrent lithium leaching; the black powder of lithium iron phosphate battery is separated by flotation to obtain graphite, and the remaining lithium iron phosphate slag is acid leached with sulfuric acid and hydrogen peroxide to obtain lithium sulfate, iron sulfate, etc. (acid leaching solution) and iron phosphate (acid leaching residue). The acid leaching residue is added with iron hydroxide (iron precipitation residue) and lithium sulfate (iron precipitation solution) in the fluorine precipitation solution; the acid leaching solution is added with an alkaline lithium solution (mainly NaOH and LiOH) to obtain a lithium hydroxide solution (lithium leaching solution) and iron hydroxide (lithium leaching residue 2); the iron precipitation solution (lithium sulfate) is added to the concentrated lithium leaching solution (mainly lithium hydroxide), and lithium phosphate is precipitated by reaction, and finally battery-grade Li3PO4 (purity 96%) and mirabilite (Na2SO4, glass raw material) are produced. The two processes on both sides achieve deep coupling through complementary acid-base reagents (alkali supplementation in the alkali leaching of aluminum slag and acid supplementation in the acid leaching of the black powder of lithium iron phosphate battery), energy sharing (MVR evaporation is used to treat the two waste liquids in series), and residue synergy (mixed calcination for brick making). The comprehensive output value per ton of mixed waste reaches 1970 US dollars, the energy consumption is reduced by 50% compared with the traditional process, and there is zero emission of fluorine and heavy metals, constructing a model process for the collaborative resource utilization of "aluminum - lithium" double wastes.

[0041] The process advantages are reflected in three dimensions: resource efficiency, economy, and environmental friendliness. First, through the cascaded utilization of components and the mutual supply of materials, only calcium hydroxide needs to be supplemented after normal operation, and sodium hydroxide is no longer used, reducing the reagent cost by 40% and the energy consumption by 50%. Second, the co-production of high-value-added products - the combination of CaF2 (building materials), Li3PO4 (battery materials), Glauber's salt (glass raw materials), and environmentally friendly building materials results in a waste value of $1,970 per ton, far exceeding the traditional model. Third, the environmental load is extremely low, and fluorine and heavy metal pollutants achieve "zero emissions" through a closed-loop cycle. All residues are converted into building materials, completely avoiding the risk of secondary pollution. This process flow is clear, the technology maturity is high, and it is suitable for large-scale industrial production. It is a benchmark solution to break the dilemmas of "aluminum-lithium" resource shortage and solid waste pollution.

[0042] The technical attribute of this process is a disruptive technology. This process starts with zero reagents, that is, only NaOH is initially added, and subsequent operations completely rely on the recycling of by-products, subverting the reagent-dependent mode of traditional processes. In addition, lithium can be directionally enriched. Through three-level cycles of leaching-lithium precipitation-mother liquor recycling, the lithium recovery rate breaks through 95%, far exceeding the traditional process (<80%).

[0043] This process realizes the deep resource utilization of electrolytic aluminum waste residue and waste lithium-ion battery materials through "three-step coordination of acid-base-fluorine": using sulfuric acid / sodium hydroxide as the medium, combining MVR evaporation and countercurrent washing technologies to simultaneously recover lithium (Li3PO4) and calcium fluoride (CaF2). The comprehensive waste value per ton reaches $1,970, and the energy consumption is reduced by 50% compared with the traditional process, with zero emissions of fluorine and heavy metals. Its core advantage lies in "mutual supply of two wastes and closed-loop cycle" - the alkaline leaching agent for aluminum slag is derived from lithium-ion battery waste liquid, and fluorine resources are recycled bidirectionally in the fluorine leaching and lithium leaching links, achieving zero loss of reagents and energy. At the same time, high-value-added battery materials, building material raw materials, and environmental protection bricks are produced. The investment return period is only 2.5 years. It is a "zero-waste" benchmark process that combines economy, environmental protection, and industrial synergy.

[0044] Compared with the traditional electrolytic aluminum waste residue treatment process and waste lithium iron phosphate battery treatment process, this co-treatment process has many advantages. The following is a detailed comparison from aspects such as resource recovery, environmental benefits, and economic benefits:

[0045] (Resource recovery) Advantages of the co-treatment process:

[0046] (1) Comprehensive utilization: Simultaneously recover valuable elements such as aluminum, fluorine in electrolytic aluminum waste residue, and lithium, iron, and phosphorus in waste lithium-ion batteries, achieving the maximum utilization of resources.

[0047] (2) Improve the recovery rate: By optimizing process conditions, the recovery rates of elements such as lithium and fluorine may be higher than those of traditional single-treatment processes.

[0048] (3) Reduction of resource waste: Components that were not fully utilized in traditional processes (such as lithium and phosphorus in waste lithium batteries, and lithium and fluorine in electrolytic aluminum waste residues) are recovered during co - processing (environmental benefits) Advantages of the co - processing technology:

[0049] (1) Reduction of waste: Through co - processing, the total amount of waste residues and waste liquids generated from the separate treatment of electrolytic aluminum waste residues and waste lithium batteries is reduced.

[0050] (2) Reduction of pollution risk: Harmful components in waste liquids and waste residues (such as fluorides and heavy metals) are effectively recovered and treated during co - processing, reducing the environmental pollution risk.

[0051] (3) Recycling of reagents: The alkaline substances in electrolytic aluminum waste residues can be used in the leaching process of waste lithium batteries, while the acidic substances in waste lithium batteries can be used for acid leaching of electrolytic aluminum waste residues, reducing reagent consumption and waste liquid generation.

[0052] (Economic benefits) Advantages of the co - processing technology:

[0053] (1) Multi - element recovery: Multiple valuable elements such as lithium, aluminum, and fluorine are recovered simultaneously, improving the overall economic benefits.

[0054] (2) Cost reduction: Through reagent recycling and process integration, reagent consumption and energy costs are reduced.

[0055] (3) High - value - added products: Substances such as recovered lithium and fluorides have high market values. Especially lithium, as a key element for battery materials, has a continuously growing demand.

[0056] This process realizes the efficient recovery of resources and the reduction of waste through the co - processing of electrolytic aluminum waste residues and waste lithium iron phosphate battery electrode materials. Its core features include the co - processing of two industrial wastes, reduction of acid and alkali use, recycling of fluorine - precipitating liquid and alkaline lithium liquid, and efficient recovery of lithium. These features give this process significant advantages in terms of comprehensive resource utilization, environmental benefits, and economic benefits, providing an innovative solution for the treatment of industrial waste. By further optimizing process parameters and equipment, its feasibility and application prospects can be further improved. Brief Description of the Drawings

[0057] Figure 1 is a schematic process flow diagram of the present invention;

[0058] Figure 2 is the XRD pattern of lithium phosphate obtained in Example 1 of the present invention;

[0059] Figure 3 is the XRD pattern of lithium phosphate obtained in Example 2 of the present invention;

[0060] Figure 4 XRD pattern of the Glauber's salt obtained in Example 1 of the present invention;

[0061] Figure 5 XRD pattern of the Glauber's salt obtained in Example 2 of the present invention;

[0062] Figure 6 XRD pattern of the calcium fluoride obtained in Example 1 of the present invention;

[0063] Figure 7 XRD pattern of the calcium fluoride obtained in Example 2 of the present invention;

[0064] Figure 8 SEM image of the iron-precipitated slag obtained in Example 1 of the present invention;

[0065] Figure 9 SEM image of the iron-precipitated slag obtained in Example 2 of the present invention. Detailed implementation manners

[0066] In order to make the technical means, creative features, achieved purposes and effects of the present invention easy to understand, the following specifically describes a method for co-recovering valuable metals in waste lithium batteries and electrolytic aluminum waste residues of the present invention in conjunction with examples and drawings.

[0067] Unless otherwise specified, the raw materials and materials used in the examples of the present invention are purchased through general commercial channels; all unmentioned test methods are national standards.

[0068] The instruments for detection are as follows:

[0069] ICP is the German Spectro inductively coupled plasma emission spectrometer (SPECTRO ARCOS); the Japanese Electron field emission scanning electron microscope (JSM-7800F); the Japanese Rigaku X-ray diffractometer (Smartlab SE XRD).

[0070] As Figure 1 shown, a method for co-recovering valuable metals in waste lithium batteries and electrolytic aluminum waste residues of the present invention specifically includes the following steps:

[0071] (A) Recycling process of electrolytic aluminum waste residue

[0072] Step S0, crushing the original electrolytic aluminum waste residue into particles with a diameter less than 5 mm, washing away the soluble impurities with water, and then drying at 50 °C to 80 °C to obtain the electrolytic aluminum waste residue for use.

[0073] Step S1, slurring treatment: mixing the electrolytic aluminum slag after washing and drying in step S0 with water at a mass ratio of 1:(1 - 5) to form a slurring mixture, i.e., a slurrying liquid.

[0074] Step S2, alkali leaching treatment: Add sodium hydroxide solution (10% - 15% mass fraction) to the pulping liquid for alkali leaching. The mass ratio of electrolytic aluminum slag after washing and drying to sodium hydroxide is 1:(3 - 5). Heat up to 40°C - 50°C and keep warm for 15 - 30 minutes to obtain alkali leaching slurry. The main components of electrolytic aluminum slag are Na3AlF6, Al2O3, NaF, and LiF. The reaction equations are as follows:

[0075] Na3AlF6 + 4NaOH → 6NaF + NaAlO2 + 2H2O;

[0076] Al2O3 + 2NaOH → 2NaAlO2 + H2O.

[0077] In a strongly alkaline solution, NaF is easily soluble in water and exists in the form of Na + and F - while LiF is partially dissolved.

[0078] Step S3, fluorine leaching treatment: Continue to add sodium hydroxide to the alkali leaching slurry obtained in Step S2, heat up to 40 - 50°C, keep warm and react for 0.5 - 2 hours, and maintain the pH value of the system at 10 - 14 to obtain fluorine leaching liquid and fluorine leaching residue (LiF).

[0079] Sodium aluminate (NaAlO) and sodium fluoride (NaF) generated by alkali leaching are easily soluble in a strongly alkaline solution and exist in ionic form (Na+, AlO2-, F-). Therefore, the sodium hydroxide added during fluorine leaching makes the sodium hydroxide (NaOH) in the system excessive, providing a high concentration of OH- to inhibit the hydrolysis of AlO2 - (AlO2 - + 2H2O ⇌ Al(OH)3 + OH-), ensuring that aluminum exists in a soluble state. In a strongly alkaline solution, LiF mainly exists in solid form and only a small amount dissolves.

[0080] The core of the fluorine leaching step is to dissolve the soluble sodium salts (NaAlO2, NaF) in the alkali leaching slurry with a high - concentration NaOH solution and separate out the insoluble LiF solid (fluorine leaching residue).

[0081] Step S4, fluorine precipitation: According to the molar amount of substances, add calcium hydroxide with a molar ratio of fluorine amount 1.5 - 2.0 times to the fluorine leaching liquid, heat up to 40°C - 50°C, keep warm and react for 0.5 - 2 hours to obtain fluorine precipitation liquid and crude calcium fluoride precipitate. The reaction equation is: 2F- + Ca(OH)2 → CaF2↓ + 2OH - . And the fluorine precipitation liquid contains: Na + , Li + , AlO 2-, OH-. 10% - 50% of the fluorine precipitation solution returns to step S1 to participate in pulping, 10% - 50% returns to step S2 to participate in alkali leaching, 10% - 50% returns to step S3 to participate in fluorine leaching, and 20% - 70% is used for subsequent iron precipitation.

[0082] Step S5, secondary countercurrent washing: Add water 4 - 6 times the mass of the obtained crude calcium fluoride (CaF2), heat up to 40°C - 50°C for secondary countercurrent washing to obtain washing solution 1 and pure calcium fluoride (CaF2). This pure CaF2 can be packed and sold as a commodity. Washing solution 1 continues to return to step S3 for fluorine leaching to make full use of the unreacted sodium hydroxide in it.

[0083] Steps S1 - S5 belong to the startup stage, and step S6 and onwards belong to the normal operation stage. After the first batch of aluminum slag is added and the normal operation starts, aluminum slag is added in batches. At this time, no additional water is added for pulping in step S1, and no additional sodium hydroxide solution is added for alkali leaching either. Therefore, as batches of aluminum slag are continuously added, steps S1 - S5 are repeated.

[0084] Step S6, secondary countercurrent lithium leaching: Add calcium hydroxide solution (concentration 0.05 - 0.15 mol / L) (in a mass ratio of leaching residue of fluorine and calcium hydroxide of 1:(3 - 5)) to the fluorine leaching residue (LiF) in step S3, heat up to 60 - 80°C, keep the temperature for reaction for 0.5 - 3 hours for secondary countercurrent lithium leaching to obtain alkaline lithium solution and lithium leaching residue 1.

[0085] Step S7, add water to lithium leaching residue 1 in a mass ratio of 1:(1 - 3). The residue mainly composed of calcium fluoride after washing is used to make hollow bricks, and washing solution 2 returns to be used for secondary countercurrent lithium leaching.

[0086] The reaction equation for secondary countercurrent lithium leaching is as follows:

[0087] LiF + Ca(OH)2 → CaF2↓ + 2LiOH.

[0088] Secondary countercurrent lithium leaching:

[0089] After primary lithium leaching, there is a small amount of unreacted LiF remaining in the lithium leaching residue. Through additional Ca(OH)2 addition in secondary lithium leaching, it is further converted into CaF2 precipitate and LiOH to increase the total lithium leaching rate. After lithium leaching residue 1 is washed, the main components of the residue are calcium fluoride and a very small amount of lithium fluoride, and there is still residual Li in washing solution 2 + Continue to return to secondary countercurrent lithium leaching for continuous recovery.

[0090] (B) Lithium iron phosphate (LEP) black powder recovery part

[0091] Step S8: Add pure water to the waste lithium iron phosphate black powder for flotation. The lower layer is the slag of the lithium iron phosphate battery black powder (LEP), and the upper layer is graphite.

[0092] Step S9: Acid leaching treatment: Add sulfuric acid and hydrogen peroxide to the slag of the lithium iron phosphate battery black powder (LEP) simultaneously, keep the pH value of the solution at 1.0 - 3.0, heat up to 60°C - 80°C, and hold for 24 hours to ensure that all iron is converted into ferric ions, obtaining an acid leaching solution and acid leaching slag. The reaction equation is:

[0093] 2LiFePO4 + H2SO4 + H2O2 → Li2SO4 + 2FePO4↓ + 2H2O.

[0094] Step S10: Add the fluorine precipitation solution from Step S4 to the acid leaching slag from Step S9 to obtain iron precipitation slag and iron precipitation solution. The reaction equation is:

[0095] FePO4 + 3NaOH → Fe(OH)3 + Na3PO4.

[0096] The main component of the iron precipitation slag is iron hydroxide, which can be used to burn red bricks.

[0097] Step S11: Add the alkaline lithium solution from Step S6 to the acid leaching solution from Step S9 to obtain a lithium leaching solution and lithium leaching slag 2. The reaction equations are:

[0098] Li2SO4 + NaOH → LiOH + Na2SO4;

[0099] FePO4 + 3NaOH → Fe(OH)3↓ + Na3PO4.

[0100] Step S12: Concentrate the lithium leaching solution from Step S11 by 3 - 5 times to obtain a lithium leaching mother liquor.

[0101] Step S13: Add the iron precipitation solution from Step S10 to the lithium leaching mother liquor from Step S12 for lithium precipitation. Lithium phosphate crystals precipitate, and Glauber's salt remains in the solution. The reaction equation is: 3Li2SO4 + 2Na3PO4 = 3Na2SO4 + 2Li3PO4↓.

[0102] Step S14: The solution from Step S13 is processed by the mechanical vapor recompression (MVR) process to obtain solid Glauber's salt; the evaporation temperature of MVR is 80 - 100°C, and the evaporation pressure is 0.1 - 0.3 MPa.

[0103] The product analysis of the above process is as follows:

[0104] (1) Calcium fluoride is widely used in the metallurgy, chemical, and building materials industries as a raw material or additive for fluorides.

[0105] (2) Lithium phosphate is an important raw material for the positive electrode materials of lithium-ion batteries and can also be used to prepare other lithium compounds.

[0106] (3) Iron hydroxide can be used to prepare iron-based pigments, water treatment agents, and magnetic materials.

[0107] (4) The residue from step S7 is used to make hollow bricks for building materials.

[0108] (5) Glauber's salt (sodium sulfate) is used in the chemical industry, paper-making, glass, dye, printing and dyeing, and pharmaceutical industries.

[0109] (6) The iron precipitation residue is used to bake red bricks for building materials.

[0110] In the present invention, the black lithium iron phosphate powder is derived from the positive electrode materials of waste lithium iron phosphate batteries, and the disassembly method is as follows:

[0111] Place the waste battery in a 7% wt NaCl solution and discharge it until no bubbles emerge. Then, place the discharged waste battery in a glove box filled with argon, use scissors to disassemble the battery casing, soak the peeled positive electrode sheet in an ethanol solution. After 30 minutes, rinse the soaked positive electrode sheet under a flowing tap water to wash away the ethanol and electrolyte on the surface. Then, lay the positive electrode sheet flat in a fume hood to dry, obtaining the disassembled positive electrode sheet, i.e., the positive electrode sheet (positive electrode material). Use ICP to detect the content of some metals in the positive electrode materials of the waste lithium iron phosphate batteries, and the results are shown in Table 1.

[0112] Table 1. Content of some metals in the positive electrode materials of waste lithium iron phosphate (wt%)

[0113] Li Fe Al Cu 3.3 36.4 7.5 0.9

[0114] In the present invention, the electrolytic aluminum waste residue is a waste generated during the production of electrolytic aluminum, including overhaul slag, aluminum ash, carbon powder, waste electrolyte, covering material, etc. Use ICP to detect the content of some elements in the electrolytic aluminum waste residue, and the results are shown in Table 2.

[0115] Table 2. Content of some elements in the electrolytic aluminum waste residue (wt%)

[0116] Composition Li Al F Na Si K Fe Content 1.3 22.5 40.3 20.5 1.3 1.5 0.8

[0117] <Example 1>

[0118] Step S0, preparation stage: Crush the original electrolytic aluminum waste residue into particles with a diameter less than 5 mm, wash away the soluble impurities with water, and dry at 60 °C to obtain the electrolytic aluminum slag for standby. The water addition amount is 1000 mL of water for every 500 g of the original electrolytic aluminum waste residue.

[0119] Startup stage:

[0120] Step S1: Take 500 g of electrolytic aluminum slag, add water to it (the mass ratio of electrolytic aluminum slag to water is 1:3), and carry out pulping treatment to obtain a pulp liquid.

[0121] Step S2: Add sodium hydroxide solution (15% mass fraction) to the above pulp liquid for alkali leaching, with the mass ratio of aluminum slag to sodium hydroxide being 1:5. Heat up to 50 °C and keep warm for 15 minutes to obtain an alkali leaching solution. Then, return all the washing solution 1 obtained in the subsequent steps.

[0122] Step S3: Continue to add sodium hydroxide solution to the above alkali leaching solution, heat up to 40 °C, keep warm for 30 minutes, and maintain the pH value of the system at 10 to obtain a fluorine leaching solution and a fluorine leaching residue (LiF).

[0123] Step S4: According to the molar amount, add calcium hydroxide with a fluorine amount 2.0 times that of the above fluorine leaching solution, heat up to 50 °C, keep warm and react for 0.5 hour to generate crude calcium fluoride precipitate and a fluorine precipitation solution. 10% of this fluorine precipitation solution is returned to pulping, 10% is returned to alkali leaching, 40% is returned to fluorine leaching, and 40% is used for subsequent iron precipitation.

[0124] Step S5: Add water with a weight 6 times that of the obtained crude calcium fluoride to it, heat up to 50 °C, and carry out two-stage countercurrent washing to obtain pure calcium fluoride and washing solution 1; return washing solution 1 to step S3 for fluorine leaching.

[0125] After step S5 is completed and enters normal operation, then add aluminum slag in batches, 500 g each time. In step S1, no additional water is added during pulping, and no additional sodium hydroxide solution is added during alkali leaching. Therefore, as the aluminum slag is continuously added batch by batch, repeat step S1 ～ S5.

[0126] Formal operation stage:

[0127] Step S6: For the fluorine leaching residue obtained from fluorine leaching, according to the mass ratio of fluorine leaching residue to calcium hydroxide being 1:3, add calcium hydroxide solution (concentration 0.15 mol / L), heat up to 80 °C, keep warm and react for 0.5 hour, and carry out two-stage countercurrent lithium leaching to obtain an alkaline lithium solution and a lithium leaching residue 1.

[0128] Step S7: Add water to the lithium leaching residue 1 according to a mass ratio of 1:1. The residue after washing the lithium leaching residue 1 is used to make hollow bricks, and the washing solution 2 is returned to step S6 for two-stage countercurrent lithium leaching.

[0129] Step S8: Add 200 g of waste black powder of lithium iron phosphate battery and water, and separate lithium iron phosphate slag and graphite by flotation.

[0130] Step S9: The obtained lithium iron phosphate slag is leached with 1500 mL of sulfuric acid solution with a concentration of 2 M and 100 mL of hydrogen peroxide (the total liquid-solid ratio (L / kg) is 8) (maintaining the solution pH value at 1.5), heated to 80 °C, and kept reacting for 2 hours to obtain an acid leaching solution and acid leaching residue.

[0131] Step S10: 30% of the previously obtained fluoride precipitation solution is added to the lithium leaching residue, and the reaction is carried out for 1 hour to obtain iron precipitation residue and iron precipitation solution.

[0132] Step S11: Alkaline lithium solution is added to the acid leaching solution, and the reaction is carried out for 0.5 hour. After solid-liquid separation, a lithium leaching solution and lithium leaching residue 2 are obtained.

[0133] Step S12: The lithium leaching solution is concentrated by 3 times to obtain a lithium leaching mother liquor.

[0134] Step S13: The iron precipitation solution is added to the lithium leaching mother liquor, heated to 40 °C, and kept reacting for 4 hours. Lithium phosphate is precipitated, and the solution is concentrated by steam mechanical recompression technology (MVR) to obtain mirabilite. The evaporation temperature of MVR is 80 °C, and the evaporation pressure is 0.3 MPa.

[0135] <Example 2>

[0136] Step S0, Preparation stage: The original electrolytic aluminum waste residue is crushed into particles with a diameter less than 5 mm, and the soluble impurities are washed away with water, and dried at 80 °C to obtain electrolytic aluminum slag for standby. The amount of added water is 1000 mL of water added to 500 g of the original electrolytic aluminum waste residue.

[0137] Startup stage:

[0138] Step S1: Water is added to the electrolytic aluminum slag particles (the mass ratio of electrolytic aluminum slag to water is 1:5) for pulping treatment to obtain a pulping solution.

[0139] Step S2: Sodium hydroxide solution (10% mass fraction) is added to the pulping solution for alkali leaching, with the mass ratio of aluminum slag to alkali solution being 1:3, heated to 40 °C, and kept for 30 minutes to obtain an alkali leaching solution.

[0140] Step S3: Sodium hydroxide is continuously added to the above alkali leaching solution, heated to 50 °C, and kept for 30 minutes to maintain the pH value of the system at 14, obtaining a fluoride leaching solution and fluoride leaching residue (LiF).

[0141] Step S4: According to the molar amount, calcium hydroxide with 1.5 times the fluoride amount is added to the above fluoride leaching solution, heated to 40 °C, and kept reacting for 2 hours to generate crude calcium fluoride precipitation and fluoride precipitation solution. 30% of this fluoride precipitation solution is returned to pulping, 30% is returned to alkali leaching, 10% is returned to fluoride leaching, and 30% is used for subsequent iron precipitation.

[0142] Step S5: Add water in an amount four times the mass of the obtained crude calcium fluoride, heat up to 40°C, and perform two-stage countercurrent washing to obtain pure calcium fluoride and Wash Liquor 1; return Wash Liquor 1 to the fluorine leaching step.

[0143] After Step S5 is completed and normal operation begins, add aluminum slag in batches, 500 g each time. In Step S1, no additional water is added during pulping, and no additional sodium hydroxide solution is added during alkali leaching. Therefore, as the aluminum slag is continuously added batch by batch, repeat Step S1 ～ S5.

[0144] Formal operation stage:

[0145] Step S6: For the fluorine leaching residue obtained from fluorine leaching, add calcium hydroxide solution (concentration 0.05 mol / L) according to the mass ratio of fluorine leaching residue to calcium hydroxide of 1:5, heat up to 60°C, hold for reaction for 3 hours, and perform two-stage countercurrent lithium leaching to obtain alkaline lithium solution and Lithium Leaching Residue 1.

[0146] Step S7: Add water to Lithium Leaching Residue 1 according to the mass ratio of 1:3. The residue after washing Lithium Leaching Residue 1 is used to make hollow bricks, and Wash Liquor 2 is returned for use in two-stage countercurrent lithium leaching.

[0147] Step S8: Add water to 200 g of waste black powder of lithium iron phosphate batteries and separate lithium iron phosphate slag and graphite by flotation.

[0148] Step S9: Leach the obtained lithium iron phosphate slag with 1500 mL of sulfuric acid solution with a concentration of 2 M and 100 mL of hydrogen peroxide (total liquid-solid ratio (L / kg) is 8) (maintain the solution pH value at 3.0), heat up to 60°C, hold for reaction for 4 hours to obtain acid leaching solution and acid leaching residue.

[0149] Step S10: Add 30% of the previously obtained fluorine precipitation solution to the lithium leaching residue, react for 0.5 hour to obtain iron precipitation slag and iron precipitation solution.

[0150] Step S11: Add the alkaline lithium solution to the acid leaching solution, react for 1.0 hour, and after solid-liquid separation, obtain lithium leaching solution and Lithium Leaching Residue 2.

[0151] Step S12: Concentrate the lithium leaching solution by 5 times to obtain lithium leaching mother liquor.

[0152] Step S13: Add the iron precipitation solution to the lithium leaching mother liquor, heat up to 80°C, hold for reaction for 1 hour, precipitate lithium phosphate, and concentrate the solution by steam mechanical recompression technology (MVR) to obtain mirabilite. The evaporation temperature of MVR is 100°C, and the evaporation pressure is 0.1 MPa.

[0153] <Test Example 1>

[0154] The lithium phosphate, sodium sulfate and calcium fluoride obtained from the electrolytic aluminum slag recycling process in Example 1 were measured using an X-ray diffractometer (Smartlab SE XRD), and the obtained X-ray diffraction patterns are as follows Figure 2 , Figure 4 , Figure 6 shown.

[0155] From Figure 2 it can be seen that the lithium phosphate recovered from the lithium iron phosphate powder shows high crystallinity and a pure phase, and its peak positions are highly consistent with the standard Li3PO4 PDF card number (PDF#97-001-0257), indicating that the recovered lithium phosphate can be directly used as the lithium source precursor component of the cathode material for lithium iron phosphate batteries (LiFePO4 batteries).

[0156] From Figure 4 it can be observed that all the diffraction peaks are highly consistent with the Na2SO4 PDF card number (PDF#97-002-7654), and no other impurity diffraction peaks appear. This shows that under such process conditions, it is feasible to extract sodium sulfate from waste lithium iron phosphate materials. Further analysis shows that the intensity of the diffraction peaks is relatively uniform, and there is no obvious particle aggregation or phase separation phenomenon. Overall, the XRD results show that the sodium sulfate recovered from this process has a high purity and can be used in the chemical industry, papermaking, glass, dyes, printing and dyeing, and pharmaceutical industries.

[0157] From Figure 6 it can be seen that for the calcium fluoride recovered from the electrolytic aluminum slag, all the diffraction peaks correspond to CaF2, and the peak positions are consistent with the diffraction peaks of the standard CaF2 crystal (PDF#97-004-1413), and no other impurity diffraction peaks appear. This shows that under such conditions, the recycling process is feasible.

[0158] To verify the feasibility of this process under different conditions, the amount of calcium hydroxide added, the reaction temperature, and the reaction time were changed, and the products obtained in Example 2 were compared with those in Example 1.

[0159] The lithium phosphate obtained after heating the temperature to 80 °C and holding the reaction for 1 hour (compared with heating to 40 °C and holding the reaction for 4 hours in Example 1) was subjected to X-ray testing, as shown in Figure 3 shown, and all the diffraction peaks are highly consistent with the Na2SO4 PDF card number (PDF#97-002-7654), and no other impurity diffraction peaks appear.

[0160] The sodium sulfate obtained in Example 2 was subjected to X-ray testing, as shown in Figure 5 shown, and the product sodium sulfate shows high crystallinity and a pure phase.

[0161] In Example 2, calcium hydroxide 1.5 times the amount of fluorine was added to the fluorination solution, and the temperature was raised to 40 °C, followed by holding the reaction for 2 hours (in Comparative Example 1, calcium hydroxide 2 times the amount of fluorine was added to the fluorination solution, the temperature was raised to 50 °C, and the reaction was held for 0.5 hour). The obtained calcium fluoride was subjected to X-ray testing. As Figure 7 shown, all the diffraction peaks corresponded to CaF2, and the positions of the diffraction peaks were consistent with those of the standard CaF2 crystal (PDF#97 - 004 - 1413).

[0162] <Test Example 2>

[0163] The iron-precipitating slag obtained in Example 1 was scanned using a scanning electron microscope (SEM, JSM - 7800F). The SEM image of the fluorine-precipitating slag is as Figure 8 shown. The morphology of the iron-precipitating slag is complete without obvious cracks, indicating that the iron-precipitating slag obtained by this process has a high purity.

[0164] The iron-precipitating slag in Example 2 was scanned using a scanning electron microscope (SEM, JSM - 7800F). The SEM image of the iron-precipitating slag is as Figure 9 shown. The morphology of the iron-precipitating slag is complete without obvious cracks, which is the same as that of the iron-precipitating slag obtained in Example 1, indicating the feasibility of this process under different conditions.

[0165] <Test Example 3>

[0166] A vacuum filtration device was used to achieve solid-liquid separation of the lithium-leaching mother liquor in Example 1, and the liquid was diluted 100 times and then used for ICP detection.

[0167] Table 3 Contents of some metals in the leaching solution (g / L)

[0168] Li Fe Al Cu 0.6541 0.0002 0.000 0.000

[0169] The leaching rate of Li reached 99.1%, while Fe, Al, and Cu were basically not leached.

[0170] The above embodiments are preferred cases of the present invention and are not used to limit the protection scope of the present invention.

Claims

1. A method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum slag, characterized in that, It includes the following steps: Step S1: Mix electrolytic aluminum slag and water at a mass ratio of 1:(1 - 5) to form a pulping mixture, i.e., a pulp solution; Step S2: Add an alkali solution to the pulp solution for alkali leaching, and raise the temperature for reaction to obtain an alkali leaching solution; Step S3: Add sodium hydroxide to the alkali leaching solution, maintain the pH value of the system at 10 - 14, and raise the temperature for reaction to obtain a fluorine leaching solution and a fluorine leaching residue; Step S4: According to the molar amount of substances, add calcium hydroxide with a fluorine amount 1.5 - 2.0 times that of the fluorine leaching solution, raise the temperature for reaction to generate crude calcium fluoride precipitate and fluorine precipitation solution 1; and 10% - 50% of this fluorine precipitation solution 1 returns to Step S1 to participate in pulping, 10% - 50% returns to Step S2 to participate in alkali leaching, 10% - 50% returns to Step S3 to participate in fluorine leaching, and 20% - 70% is used for subsequent iron precipitation; Step S5: Add water with a mass 4 - 6 times that of the obtained crude calcium fluoride, raise the temperature for secondary countercurrent washing to obtain pure calcium fluoride and washing solution 1; and this washing solution 1 returns to Step S3 for fluorine leaching; Step S6: Add calcium hydroxide to the fluorine leaching residue obtained in Step S3, raise the temperature for secondary countercurrent lithium leaching to obtain an alkaline lithium solution and a lithium leaching residue 1; Step S7: Add water to the lithium leaching residue 1 at a mass ratio of 1:(1 - 3), and obtain a residue and washing solution 2 after washing; this washing solution 2 returns to be used for secondary countercurrent lithium leaching; Step S8: Add water to waste lithium iron phosphate black powder for flotation to obtain lithium iron phosphate battery black powder residue and graphite; Step S9: Add sulfuric acid and hydrogen peroxide to the lithium iron phosphate battery black powder residue simultaneously, keep the pH value of the solution at 1.0 - 3.0, raise the temperature for reaction until all iron is converted into ferric ions to obtain an acid leaching solution and an acid leaching residue; Step S10: Add the fluorine precipitation solution obtained in Step S4 to the acid leaching residue obtained in Step S9 to obtain an iron precipitation residue and an iron precipitation solution; Step S11: Add the alkaline lithium solution obtained in Step S6 to the acid leaching solution obtained in Step S9 to obtain a lithium leaching solution and a lithium leaching residue 2; Step S12: Concentrate the lithium leaching solution obtained in Step S11 by 3 - 5 times to obtain a lithium leaching mother liquor; Step S13: Concentrate and separate lithium phosphate and mirabilite from the lithium leaching mother liquor.

2. The method for co - recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residues according to claim 1, wherein: Among them, It further includes Step S0: Crush the original electrolytic aluminum waste residue into particles with a diameter less than 5 mm, wash away soluble impurities with water, and then dry to obtain the electrolytic aluminum slag for use.

3. The method for co - recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residues according to claim 2, wherein: Among them, In Step S2, the alkali is sodium hydroxide or potassium hydroxide.

4. The method for co - recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residues according to claim 3, wherein: Among them, The base is sodium hydroxide. By mass ratio, electrolytic aluminum slag:sodium hydroxide = 1:(3 ～ 5), and the temperature is raised to 40 °C ～ 50 °C, and the mixture is kept warm and reacted for 15 ～ 30 minutes; the mass fraction concentration of the sodium hydroxide solution is 10% ～ 15%.

5. The method for co - recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residues according to claim 2, wherein: Among them, In step S3, heat up to 40°C ～ to 50°C, and keep the temperature for 0.5 ～ to 2 hours; in step S4, heat up to 40°C to 50°C, and keep the temperature for 0.5 to 2 hours.

6. The method for co - recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residues according to claim 2, wherein: Among them, In step S5, the temperature is raised to 40°C to 50°C for secondary countercurrent washing.

7. The method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residues according to claim 2, characterized in that: Among them, In step S6, a calcium hydroxide solution with a concentration of 0.05 - 0.15 mol / L is added to the fluorine-containing residue, and the mass ratio of the fluorine-containing residue to calcium hydroxide is 1:(3 - 5). The temperature is raised to 60 ～ - 80 °C, and the mixture is kept warm for reaction for 0.5 ～ - 3 hours.

8. The method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residues according to claim 2, characterized in that: Among them, In step S9, sulfuric acid and hydrogen peroxide are simultaneously added to the black powder residue of the lithium iron phosphate battery, and the pH value of the solution is maintained at 1.0 ～ 3.0, and the temperature is raised to 60°C ～ 80°C; according to the molar ratio of substances, the addition amount of the sulfuric acid is 3 ～ 6 times the lithium content in the black powder residue of the lithium iron phosphate battery, and the addition amount of the hydrogen peroxide is 1.5 ～ 3 times.

9. The method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residues according to claim 2, characterized in that: Among them, It further includes step S14: obtaining solid Glauber's salt from the solution in step S13 through the MVR process.

10. The method for co-recovering valuable metals from waste lithium batteries and electrolytic aluminum waste residues according to claim 9, characterized in that: Among them, The evaporation temperature of MVR is 80 - 100°C, and the evaporation pressure is 0.1 - 0.3 MPa.

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