Method for recycling lithium iron phosphate powder with iron salts and recovering all components

Abstract

This invention discloses a method for leaching lithium iron phosphate mixed powder with iron salts and recovering all components, belonging to the field of battery recycling. The invention uses an iron salt solution to leach the mixed powder, obtaining a lithium-containing leachate and leaching residue. Ferrous ions in the leachate are regenerated into ferric iron through acidification and oxidation, and recycled for leaching the next batch of mixed powder. After reaching a preset number of cycles, the leachate is used for re-leaching multiple batches of leaching residue to improve the lithium leaching rate. Ultimately, a enriched solution containing Li, Fe, Cu, and Al and graphite-containing iron phosphate residue are obtained. Copper is recovered from the enriched solution through iron powder replacement, and a high-purity lithium chloride solution is obtained through extraction and separation, while ferric chloride (recycled) and aluminum chloride crystals are also recovered. The leaching residue is treated with hydrochloric acid to obtain regenerated graphite, and the pH is adjusted with alkali to obtain high-purity iron phosphate. This method achieves full component recovery under mild conditions, reducing separation steps and chemical consumption through a "leaching-regeneration-leaching" cycle mechanism, thus achieving both environmental and economic benefits.

Description

Technical Field

[0001] This invention relates to the field of efficient recycling technology for waste lithium iron phosphate batteries, specifically to a method for recycling lithium iron phosphate mixed powder by leaching iron salts and recovering all components. Background Technology

[0002] Lithium iron phosphate (LFP) batteries are widely used due to their excellent safety performance, long cycle life, and low cost. However, with the rapid expansion of the new energy vehicle and energy storage markets, coupled with the early adoption of LFP batteries, a large number of batteries are reaching the end of their service life, ushering in the first wave of retirements. This has made the recycling of LFP batteries a focal point in the current field of used power battery recycling. However, used LFP batteries contain large amounts of lithium, iron, and phosphorus, which have high added value. Recycling these materials can not only reduce environmental pollution but also achieve resource utilization. Therefore, researching how to efficiently recycle all components of used LFP batteries is particularly important.

[0003] Existing lithium iron phosphate (LFP) leaching and recovery processes primarily target cathode powder. For example, CN115611252B employs an acid leaching-oxidation-crystallization technique to recover LFP, with the precipitate treated with triethanolamine for deep impurity removal to prepare lithium carbonate. However, this method suffers from drawbacks such as the inability to recycle the leaching agent, low lithium concentration in the leaching solution, and high costs due to the large amount of triethanolamine required. To address the issues of high hydrogen peroxide consumption and the inability to recycle the leaching agent in acid leaching, CN117165770A and CN117163928A disclose a method for recovering waste LFP materials, using Fe2(SO4)3 and FeCl3 respectively to recover cathode powder and achieve leaching agent recycling. However, the leaching agent regeneration process in this method requires complex separation processes such as oxidation, extraction, and back-extraction before regeneration can be achieved, and the Li in the leaching solution remains high. + The concentration remains low. CN110474123A discloses a comprehensive recycling method for cathode materials from waste lithium iron phosphate batteries, which also uses salt solutions such as Fe2(SO4)3 and FeCl3 to leach lithium ions. The leachate is first treated with alkali and an oxidant to obtain ferric hydroxide precipitate, which is then dissolved in dilute acid to obtain a regenerated leachate agent for recycling. The regeneration process still requires a relatively cumbersome separation process and consumes a large amount of acid and alkali solutions. Furthermore, during the mother liquor recycling process, alkali cations accumulate, significantly reducing the effectiveness of the leachate agent and leading to problems such as low purity in subsequent lithium salt products.

[0004] However, in reality, the sources of recycled waste batteries are complex and some are severely damaged, making it difficult to effectively separate the positive and negative electrodes. In industrial applications, to reduce costs and increase efficiency, the positive and negative electrodes are often mixed and crushed, leading to a significant increase in the copper, aluminum, and carbon content in the mixed black powder. Currently, the recycling of mixed powder mostly uses a combination of acid and oxidant for selective lithium extraction. However, the leaching agent in this method cannot be regenerated and recycled, significantly impacting the economic feasibility of the process. For example, CN116553510B discloses a method for recycling and regenerating waste lithium iron phosphate powder. First, alkaline leaching removes fluoride ions, followed by acid leaching and oxidation to extract lithium and remove aluminum, preparing lithium carbonate. Then, high-temperature roasting and hydrothermal phosphorus supplementation prepare iron phosphate. Finally, the recovered lithium carbonate and iron phosphate are calcined to achieve the final recycling of lithium iron phosphate. Although this method can achieve closed-loop regeneration of lithium iron phosphate, the graphite roasting removal process itself suffers from high energy consumption, high carbon emissions, and severe pollution. Furthermore, the pretreatment for fluoride removal and subsequent aluminum removal processes have problems such as high liquid alkali consumption and large waste volume, significantly weakening the green economics of this process. CN113912033A discloses a method for recycling mixed positive and negative electrode powder from spent lithium iron phosphate batteries through pre-lithiation lithium extraction. The method involves a series of processes including front-end alkali removal of aluminum, selective lithium extraction with sulfuric acid and oxidant, and high-concentration acid leaching of ferric phosphate and graphite mixed residue to recover high-concentration ferric phosphate dihydrate. However, the front-end alkali removal process involves excessive alkali consumption, and the oxidative acid leaching lithium extraction process consumes a large amount of oxidant, leading to high costs. Simultaneously, residual impurities such as iron and aluminum in the leachate are not effectively removed, severely affecting the refining quality of lithium carbonate. This results in the discharge of large amounts of waste acid and alkali solutions, causing secondary pollution and significantly increasing the economic cost of recycling and treatment. CN 113443640B discloses a method for preparing battery-grade lithium carbonate and battery-grade iron phosphate using waste powder from the positive and negative electrodes of lithium iron phosphate batteries. The method involves selectively extracting lithium using inorganic acids, oxidants, and modifiers (mainly soluble iron, calcium, and magnesium salts) to obtain a high-concentration lithium solution, which is then subjected to deep purification with sodium hydroxide to produce battery-grade lithium carbonate. A mixture of iron phosphate and graphite residue is converted into battery-grade iron phosphate after acid dissolution and deep purification. However, the deep purification process with sodium hydroxide not only consumes a large amount of alkali, but the excessively high sodium ion concentration in the solution also significantly increases the difficulty of preparing battery-grade lithium carbonate.

[0005] To address the issues of high oxidant consumption and difficulty in recycling leaching agents in acid leaching, CN119220819B discloses a method for extracting lithium, copper, and aluminum from spent lithium iron phosphate batteries, using ferric sulfate solution to leach the positive and negative electrode sheets. In the pretreatment stage of spent batteries, this method employs manual, meticulous dismantling instead of mechanical crushing, directly screening the positive and negative electrode sheets as raw materials for wet leaching, effectively reducing the copper and aluminum content in the raw materials. However, manual dismantling is inefficient and costly, and the material separation precision is insufficient in large-scale production, significantly hindering large-scale continuous production. Furthermore, the ferric sulfate leaching agent regeneration process requires first reacting with LiOH and the oxidant to form ferric hydroxide precipitate, followed by extensive acid dissolution before recycling. This cumbersome regeneration process not only lengthens the overall process but also significantly increases recycling costs due to the repeated consumption of large amounts of acid and alkali reagents.

[0006] In addition to the aforementioned wet recycling method, CN118291779B discloses a method for recovering lithium from lithium iron phosphate black powder. This method involves adding auxiliary materials, mixing and calcining with lithium iron phosphate black powder, acid dissolution, pH adjustment with alkali, and solid-liquid separation to obtain lithium slurry and iron phosphate slag. Although this process is short, the high-temperature calcination oxidation of lithium iron phosphate and decomposition of carbon powder result in extremely high energy consumption, low purity of the iron phosphate product, large carbon emissions, and potential environmental pollution. CN117954728B discloses a comprehensive recovery method for lithium, iron, and graphite from waste lithium iron phosphate battery black powder. This method involves mixing waste battery black powder with alkali, high-temperature melting, and water leaching to obtain water-leached slag rich in iron and graphite and water-leached liquid rich in lithium and phosphorus. The water-leached slag is further acid-dissolved to separate an iron-containing solution and recycled graphite, achieving integrated recovery of lithium, iron, and graphite from waste lithium iron phosphate battery black powder. However, the high-temperature melting process is energy-intensive and does not effectively separate impurities from the water-leached liquid and iron-containing solution. The aqueous leaching solution was mainly a mixture of sodium pyrophosphate and lithium hydroxide, and impurities such as copper and aluminum in the iron-containing solution were not effectively removed.

[0007] In summary, while some progress has been made in the efficient recycling of waste lithium iron phosphate positive and negative electrode mixed powder, existing processes generally face problems such as difficulty in regenerating and recycling the leaching agent or cumbersome and complex regeneration steps. Furthermore, there are shortcomings in leaching efficiency and the ability to recycle all battery components. Therefore, there is an urgent need to research a leaching agent that is easily regenerated and recycled to achieve a method for the efficient recycling of all components from waste lithium iron phosphate positive and negative electrode mixed powder. Summary of the Invention

[0008] The purpose of this invention is to overcome the shortcomings of existing technologies and provide a method for leaching lithium iron phosphate mixed powder using iron salts and recovering all components. This invention uses ferric chloride or ferric sulfate solution as the leaching agent, effectively recovering all components such as Li, Fe, P, Cu, Al, and graphite from waste lithium iron phosphate battery mixed powder under milder reaction conditions. By constructing a novel and efficient "leaching-regeneration-leaching" recycling mechanism, intermediate separation steps are significantly reduced, simplifying the process. Furthermore, by re-leaching the accumulated leaching residue, the leaching efficiency of lithium ions is effectively improved, resulting in good environmental and economic benefits, thus overcoming the shortcomings of existing technologies.

[0009] The technical solution of this invention to solve the above problems is as follows: a method for cyclic leaching of lithium iron phosphate mixed powder with iron salts and recovery of all components, the method comprising the following steps:

[0010] S1. Using an iron salt solution as a leaching agent, the waste lithium iron phosphate positive and negative electrode mixed powder is thoroughly mixed and stirred to carry out the leaching reaction, resulting in leaching solution and leaching residue; the iron salt is ferric chloride or ferric sulfate.

[0011] S2. The leachate is regenerated by acidification and oxidation. The resulting regenerated leachate is returned to S1 to leach the next batch of positive and negative electrode mixed powder. The acidification and oxidation regeneration and leaching are repeated until the regenerated leachate reaches the predetermined number of regenerations.

[0012] S3. The leaching residue obtained by combining S1 and S2 is divided into several groups. Each group of leaching residue is then subjected to a re-leaching reaction with the regenerating leaching agent after S2 has reached the predetermined number of regenerations, to obtain the final leachate and the final leaching residue.

[0013] S4. Add iron powder to the final leachate to replace copper ions, and obtain crude copper and lithium-containing mother liquor; add acid to the crude copper, and separate the solid and liquid to obtain industrial-grade copper powder and ferrous ion solution; oxidize the ferrous ion solution, and then distill it under reduced pressure to obtain an iron salt solution as a leaching agent.

[0014] S5. Add the extractant to the lithium-containing mother liquor from S4, and selectively extract iron and aluminum in multiple stages to obtain an extract phase loaded with iron ions, an extract phase loaded with aluminum ions, and a Li-containing extract phase. + Raffinate solution; the two types of extractable phases are back-extracted and post-treated to recover iron and aluminum;

[0015] S6. Add acid to the final leaching residue of S3 to dissolve ferric phosphate, and separate the solid and liquid to obtain crude graphite and ferric phosphate solution; the crude graphite is washed and calcined to obtain regenerated graphite; phosphorus or iron source is added to the ferric phosphate solution to adjust the iron-phosphorus material ratio, and then hydrogen peroxide is added to oxidize the ferrous ions, and then alkali solution is added to react fully to obtain crude ferric phosphate; the crude ferric phosphate is washed, aged and crystallized, and calcined to prepare anhydrous ferric phosphate.

[0016] Preferably, when the leaching agent is ferric chloride, the acid used in subsequent steps is hydrochloric acid; when the leaching agent is ferric sulfate, the acid used in subsequent steps is sulfuric acid.

[0017] The present invention provides a method for cyclic leaching of lithium iron phosphate mixed powder and recovery of all components using iron salts. It utilizes the isomorphic induction effect of FeCl3 or Fe2(SO4)3 to rapidly leach lithium iron phosphate material and separates the solid and liquid components to obtain graphite-containing FePO4 leaching residue and lithium chloride leaching solution containing iron, copper and aluminum.

[0018] The leachate is directly acidified and oxidized for regeneration. Through multiple rounds of leaching and re-leaching of the leaching residue, a concentrated solution containing lithium, iron, copper, and aluminum, and a graphite-containing ferric phosphate leaching residue are finally obtained. The final leachate is first replaced with iron powder to obtain industrial-grade copper powder, then P204 extraction is used for multi-stage stepwise extraction of iron and aluminum, and then multi-stage back-extraction is used to achieve efficient separation of iron and aluminum, resulting in a high-concentration lithium solution.

[0019] The FePO4 leaching residue containing graphite is first dissolved in dilute acid to obtain regenerated graphite. Then, the pH of the ferric phosphate solution is adjusted with alkali, followed by washing, aging, and calcination to obtain high-purity regenerated anhydrous ferric phosphate. Under milder reaction conditions, the complete recovery of Li, Fe, P, Cu, Al, and graphite components from waste lithium iron phosphate battery mixed powder is effectively achieved.

[0020] The beneficial effects of this invention are:

[0021] (1) This invention proposes a novel and efficient "leaching-regeneration-leaching" cycle system. Compared with the traditional "leaching-separation-regeneration-leaching" process, its core advantage lies in significantly reducing the dependence on and number of intermediate separation steps. This innovative design not only greatly simplifies the process flow and effectively reduces operational complexity and system energy consumption, but also brings significant economic and environmental benefits: hydrogen peroxide and acid consumption can be reduced by 30%-40%, and industrial water consumption can be reduced by 40%-50%. At the same time, wastewater volume is also significantly reduced, thereby significantly reducing the total operating cost and recycling cost. After leaching with iron salt solution, the single leaching rate of lithium ions is 98.12%. Ferrous ions in the lithium-containing solution are directly acidified and oxidized to regenerate the leaching agent, enabling continuous leaching up to 10 times, with the leaching rate still being 87.94% after the tenth leaching.

[0022] (2) The present invention proposes to regenerate the leachate after a predetermined number of cycles by acidification and oxidation, and to re-leach the accumulated leaching residue, which can significantly improve the lithium ion leaching efficiency. Before the re-leaching treatment, the total lithium ion leaching rate was 93.77% after ten cycles; after the re-leaching treatment, the total leaching rate can be further increased to 99.21%.

[0023] (3) In the key impurity removal process, this invention can not only efficiently remove impurity ions from the system, but also ensure that the separated impurity ions are ultimately recycled and transformed into valuable target products. A "iron powder replacement-dilute acid washing" process is used to obtain recycled copper powder with a purity of 99.8% (recovery rate 97%). The filtrate after dilute acid washing of crude copper contains a high concentration of ferrous solution, which can be used for the cyclic leaching of mixed powder after acidification, oxidation, and vacuum distillation. Using P204 multi-stage selective extraction, a relatively pure lithium-ion solution can be obtained. After multi-stage back-extraction, the extract phase yields iron salt, aluminum salt solutions, and an empty organic phase. The iron salt solution, after vacuum distillation, can be recycled for the leaching of mixed powder. The aluminum salt solution, after evaporation and crystallization, yields aluminum chloride hexahydrate crystals or aluminum sulfate crystals containing water of crystallization. The empty organic phase can be recycled for extraction. Compared with direct chemical precipitation, this method significantly avoids the problem of large-scale alkali consumption in traditional processes, significantly reduces wastewater treatment load, thereby greatly reducing production costs and improving the green and environmentally friendly level of the process.

[0024] (4) This invention develops a full-component recycling and regeneration technology with high economic added value and environmental friendliness for waste lithium iron phosphate battery mixed powder. This technology successfully regenerates iron phosphate (96% recovery rate, 98% purity) and graphite, and recovers valuable elements including Li, Fe, P, Cu, and Al, greatly optimizing resource utilization efficiency and significantly reducing operating costs. This fully demonstrates the great industrial application prospects of this invention. Attached Figure Description

[0025] Figure 1 This is a flowchart of the cyclic leaching step in the method of the present invention.

[0026] Figure 2 This is a flowchart of the final leachate separation and purification step in the method of the present invention.

[0027] Figure 3 This is a flowchart of the final leaching residue separation and regeneration step in the method of the present invention.

[0028] Figure 4 The graph shows the effect of the number of leaching cycles on the leaching effect of lithium and copper in Example 1.

[0029] Figure 5 The graph shows the effect of the number of leaching cycles on the leaching effect of lithium and copper in Example 2.

[0030] Figure 6 The graph shows the effect of the number of leaching cycles on the lithium leaching effect in Example 3. Detailed Implementation

[0031] The technical solution of the present invention will be further described below with reference to the accompanying drawings and specific embodiments.

[0032] The examples given are only for explaining the present invention and are not intended to limit the scope of the present invention. The technical features of each embodiment of the present invention can be combined accordingly without conflict with each other.

[0033] All embodiments obtained based on the embodiments of the present invention without making any inventive achievements are within the protection scope of the present invention.

[0034] The concentration of metals in the solution was determined using flame atomic absorption spectrometry (FAAS, Thermo Fisher) and inductively coupled plasma optical emission spectrometry (ICP-OES, Agilent 5800). The leaching rate of valuable metal elements was calculated using the following formula:

[0035]

[0036] Where C i V is the concentration of element i in the leachate; V is the volume of the leachate; m is the mass of the mixed powder; w i This represents the mass percentage of element i in the mixed powder.

[0037] The main reaction equations for the leaching process and the oxidative regeneration of the leaching agent are as follows:

[0038] ;

[0039] .

[0040] The present invention relates to a method for cyclic leaching of lithium iron phosphate mixed powder with iron salts and recovery of all components, which mainly includes three steps: a cyclic leaching step, a final leachate separation and purification step, and a final leachate residue separation and regeneration step. Figures 1 to 3 The flowcharts for the three steps are shown separately.

[0041] like Figure 1 As shown, the cyclic leaching process mainly includes:

[0042] S1. Using an iron salt solution as a leaching agent, a leaching reaction is carried out by thoroughly mixing and stirring the waste lithium iron phosphate positive and negative electrode mixed powder to obtain a leaching solution and leaching residue; the iron salt is ferric chloride or ferric sulfate. The waste lithium iron phosphate positive and negative electrode mixed powder mentioned in S1 is obtained by pretreatment of waste lithium iron phosphate batteries through discharge, mechanical dismantling, etc., and in addition to the main positive electrode material lithium iron phosphate and the negative electrode material graphite, it is allowed to contain one or more of the following: binder, current collector, and electrolyte.

[0043] Preferably, the waste lithium iron phosphate positive and negative electrode mixed powder is allowed to contain one or more impurities from positive and negative electrode aluminum foil and copper foil, and among the impurities, Cu element ≤15wt.% and Al element ≤10wt.%.

[0044] Preferably, in step S1, the ferric chloride and the Li in the lithium iron phosphate positive and negative electrode powders... + The molar ratio is 1.0-2.0; the solid-liquid ratio of lithium iron phosphate positive and negative electrode powder to leachate solution is 100-600 g / L; the reaction temperature is 20-60℃; and the reaction time is 20-60 min.

[0045] S2. The leachate is regenerated through acidification and oxidation; the resulting regenerated leachate is returned to S1 to leach the next batch of positive and negative electrode mixed powder, repeating the acidification, oxidation, regeneration, and leaching process until the regenerated leachate reaches the predetermined number of regenerations. The acid used for acidification in step S2 is hydrochloric acid or sulfuric acid, with a hydrogen ion molar amount equal to the Fe content in the leachate. 2+ The molar amount is 1.0-1.5 times; the oxidant used for oxidation is hydrogen peroxide, with a hydrogen peroxide mass concentration of 10%-30%, and the amount used is equal to the Fe content in the solution. 2+ The molar amount is 0.5-1.0 times, the oxidation time is 5-30 min, and the oxidation temperature is 20-40℃.

[0046] S3. The leaching residues obtained by combining S1 and S2 are divided into several groups. Each group of leaching residues is sequentially re-leached with the regenerating leaching agent after reaching the predetermined number of regenerations in S2 to obtain the final leachate and the final leaching residue. In S3, the regenerating leaching agent sequentially and continuously re-leaches each group of leaching residues without re-oxidizing the leachate. The solid-liquid ratio of each group of re-leaching reactions is 100-600 g / L, the reaction temperature is 20-60℃, and the reaction time is 20-60 min.

[0047] like Figure 2 As shown, the final leachate separation and purification steps mainly include:

[0048] S4. Add iron powder to the final leachate to displace copper ions, obtaining crude copper and a lithium-containing mother liquor; add acid to the crude copper, and perform solid-liquid separation to obtain industrial-grade copper powder and a ferrous ion solution; oxidize the ferrous ion solution, and then distill under reduced pressure to obtain an iron salt solution as the leaching agent; the amount of iron powder in S4 is 1.0-2.0 times the copper ion content in the solution; the hydrogen ion concentration in the acid is 0.5-5.0 mol / L; the oxidant used to oxidize the ferrous ion solution is hydrogen peroxide with a mass concentration of 10%-30%, and the amount used is equal to the Fe content in the solution. 2+ The molar amount is 0.5-1.0 times, the oxidation time is 5-30 min, and the oxidation temperature is 20-40℃.

[0049] S5. Add the extractant to the lithium-containing mother liquor from S4, and selectively extract iron and aluminum in multiple stages to obtain an extract phase loaded with iron ions, an extract phase loaded with aluminum ions, and a Li-containing extract phase. + Raffinate solution; the two types of extractable phases are back-extracted and post-treated to recover iron and aluminum;

[0050] In this process, the mother liquor containing lithium has a pH < 1. An extractant is added to it to obtain an extract phase loaded with iron ions and a raffinate phase containing aluminum and lithium salts. For the treatment of the raffinate phase containing aluminum and lithium salts, LiOH solution is first added to the extractant for saponification, controlling the degree of saponification to 40%-80%. The saponified extractant is then subjected to multi-stage extraction with the raffinate phase containing aluminum and lithium salts to obtain an extract phase loaded with aluminum ions and a raffinate phase containing Li... + Raffinate solution;

[0051] The extractants were all kerosene-diluted di(2-ethylhexyl) phosphate, with a volume fraction of 20%-40% in the extractants; the volume ratio of organic phase to aqueous phase in each extraction stage was (1-3):1.

[0052] In S5, the two types of extractable phases are back-extracted and post-treated to recover iron and aluminum, specifically including:

[0053] The iron-loaded extraction phase is subjected to acid multi-stage back-extraction to obtain an empty organic phase and an iron-containing aqueous phase. The empty organic phase is recovered as an extractant for cyclic extraction of iron, and the aqueous phase is subjected to vacuum distillation to obtain an iron salt leaching agent.

[0054] The aluminum-loaded extraction phase is subjected to acid multi-stage back-extraction to obtain an empty organic phase and an aluminum-containing aqueous phase. The empty organic phase is recovered as an extractant for cyclic extraction of aluminum, and the aqueous phase is evaporated and crystallized to obtain aluminum salt crystals.

[0055] The acids used in all back-extraction processes are hydrochloric acid or sulfuric acid, with hydrogen ion concentrations ranging from 2 to 8 mol / L. During the back-extraction process, the volume ratio of the organic phase to the aqueous phase is 1:(1-3).

[0056] like Figure 3 As shown, the final leaching residue separation and regeneration steps mainly include:

[0057] S6. Add acid to the final leaching residue obtained from the cyclic leaching step to dissolve ferric phosphate, and separate the solid and liquid to obtain crude graphite and ferric phosphate acid solution; wash the crude graphite with acid in small amounts and multiple times, and then calcine at high temperature under a nitrogen atmosphere to remove binder PVDF and other substances to obtain relatively pure regenerated graphite; add phosphorus or iron source to the ferric phosphate acid solution to adjust the iron-phosphorus material ratio, then add hydrogen peroxide to oxidize the ferrous ions, and then add alkaline solution to react fully to obtain crude ferric phosphate; the crude ferric phosphate is washed, aged and crystallized, and calcined to prepare anhydrous ferric phosphate.

[0058] The phosphorus source is one or more of ammonium dihydrogen phosphate, diammonium hydrogen phosphate, and phosphoric acid, with an iron-to-phosphorus molar ratio of 0.9-1.2; the iron source is ferric chloride or ferric sulfate; the alkaline solution is one or more of potassium hydroxide, sodium hydroxide, and ammonia water, with a concentration of 1.0-3.0 mol / L, and the pH is adjusted to 1.0-4.0 with the alkaline solution. The reaction temperature is 50℃-80℃, and the reaction time is 2-4 h. The crude graphite, after acid washing, is calcined in a nitrogen atmosphere at a temperature of 450℃-700℃ for 1-3 h. The crude ferric phosphate is washed with deionized water and aged and crystallized using dilute phosphoric acid at a concentration of 0.1-1.0 mol / L for 2-10 h; the calcination temperature of the ferric phosphate is 300℃-600℃, and the calcination time is 1-3 h.

[0059] The overall method will be described in detail below with reference to several specific embodiments.

[0060] Example 1

[0061] (1) Circulating leaching

[0062] Prepare a FeCl3 solution as a leaching agent and mix it with a mixture of positive and negative electrode powders from waste lithium iron phosphate batteries (by mass percentage, the main components and contents are: Li 3.05%, Fe 24.97%, Cu 7.54%, Al 1.24%, P 14.14%, C 13.68%). Stir thoroughly to complete the leaching reaction.

[0063] The lithium molar ratio in FeCl3 and the mixed powder was 1.50, and the solid-liquid ratio was 200 g·L⁻¹. -1 Under the conditions of reaction temperature 40 ℃ and reaction time 30 min, the single leaching rate of Li was 98.13%, and lithium-containing leachate and leaching residue were obtained by solid-liquid separation.

[0064] The amount of ferrous ions in lithium-containing leachate was determined using potassium dichromate titration. (Adding Fe...) 2+ Hydrochloric acid with a molar ratio of 1.0, and Fe 2+ Hydrogen peroxide with a molar ratio of 0.7 was used for oxidation at room temperature for 30 minutes, resulting in a near 100% regeneration rate of the leaching agent, yielding a lithium-containing regenerated leaching agent, FeCl3, which was then used to leach the next batch of material (a mixture of positive and negative electrode powder from waste lithium iron phosphate batteries). This process was repeated continuously; in this example, a total of ten batches of material were leached. The tenth leaching still maintained a Li leaching rate of 87.94%. The specific effect of the number of leaching cycles on the leaching effect of lithium and copper is as follows... Figure 4 As shown.

[0065] After ten rounds of leaching, the tenth leachate L was obtained. 10The ten batches of leaching residue were collected and divided into four groups: batches one, two, and three; batches four, five, and six; batches seven and eight; and batches nine and ten. The purpose of grouping the leaching residue was to control the solid-liquid ratio in the re-leaching reaction to a suitable ratio (100-600 g / L), avoiding the problem of limited lithium re-leaching caused by an excessively high solid-liquid ratio.

[0066] The tenth leaching solution was acidified and oxidized. 10 The four groups of leaching residues were leached sequentially without the need for further oxidation. Leaching and washing were completed at a leaching temperature of 40℃ and a leaching time of 30 min. The Li leaching rate increased to 99.21%, and the total Cu leaching rate was 78.36%. The final lithium-rich leachate and final leaching residue were obtained by filtration and separation. The total Li concentration in the lithium-rich leachate was 3.5 mol / L (per 100 mL), and the final leaching residue was a mixture of ferric phosphate and graphite.

[0067] (2) Final leachate separation and purification

[0068] The composition and content of the final leachate were determined using inductively coupled plasma atomic emission spectrometry (ICP-AES), as shown in Table 1.

[0069] Table 1 - Composition and content of the final leachate

[0070]

[0071] At room temperature, the final leachate was prepared with the following ratio: n(Fe):n(Cu). 2+ Reducing iron powder was added at a molar ratio of 1.05, and the displacement reaction was completed by continuous mechanical stirring for 30 min. After solid-liquid separation and analysis, the mother liquor contained lithium (L). M Copper ions were completely removed (displacement rate 100%). The crude copper was then acid-washed with 2 mol / L hydrochloric acid at a solid-liquid ratio of 100 g / L, followed by mechanical stirring and washing at room temperature for 1 h to remove excess iron powder. After the above process, copper powder with a purity of 99.8% was finally obtained, with a recovery rate of 96.9%.

[0072] Inductively coupled plasma atomic emission spectrometry was used to detect the lithium-containing mother liquor after copper removal. M The mass concentrations of each ion in the solution are shown in Table 2. The pH of the solution was measured to be 0.6 using a pH meter.

[0073] Table 2 - Lithium-containing mother liquor (L) F Ingredients and content

[0074]

[0075] An extractant was prepared by di(2-ethylhexyl) phosphate (P204) diluted with kerosene, wherein the volume fraction of P204 was 40%. At room temperature, the extractant was mixed with a lithium-containing mother liquor diluted with water. F The mixture (O / A = 1:1, lithium-containing mother liquor may not need dilution) is placed in a water bath constant temperature shaker for 30 min to obtain a mixed solution of iron-loaded extractant phase and aluminum chloride and lithium chloride (raffinate phase). At this point, the iron ion extraction rate reaches over 99%. Because the P204 extraction process is very sensitive to pH, at lower pH levels, iron ions have a stronger complexing ability than aluminum ions and are more easily extracted by P204. Therefore, under these pH conditions (pH < 1), aluminum ions are hardly extracted.

[0076] At room temperature, the iron-loaded extractant phase was mixed with 6 mol / L hydrochloric acid (O / A = 1:1) and back-extracted in a water bath with constant temperature shaker for 30 min. After three-stage back-extraction, an empty organic phase and an iron-enriched aqueous phase were obtained. The empty organic phase could be used as an extractant again to extract iron; the iron-enriched aqueous phase mainly consisted of ferric chloride and hydrochloric acid, which was distilled under reduced pressure to obtain FeCl3, which was recycled as the leaching agent in the leaching step.

[0077] The extractant was prepared again, using the same kerosene-diluted di(2-ethylhexyl) phosphate (P204), with a P204 volume fraction of 40%. Then, 0.5 mol / L LiOH solution was added to the extractant for saponification, achieving a saponification degree of 40%. At room temperature, an equal volume of the saponified extractant was mixed with the raffinate (a mixed solution of aluminum chloride and lithium chloride) diluted with water, and the mixture was placed in a water bath with a constant temperature shaker for 30 min. After three stages of extraction, an aluminum-loaded extractant and a relatively pure lithium chloride solution (raffinate) were obtained.

[0078] At room temperature, the aluminum-loaded extractant phase was mixed with 6 mol / L hydrochloric acid (O / A = 1:1) and back-extracted in a water bath with constant temperature shaker for 30 min. After multi-stage back-extraction, an empty organic phase and an aluminum-enriched aqueous phase were obtained. The empty organic phase could be used again to extract aluminum; the aluminum-enriched aqueous phase mainly consisted of aluminum chloride and hydrochloric acid, which could be evaporated and crystallized to obtain aluminum chloride hexahydrate crystals.

[0079] (3) Final leaching residue separation and regeneration

[0080] A 2 mol / L dilute hydrochloric acid solution was added to the final leaching residue at a solid-liquid ratio of 200 g / L, and the reaction was carried out at 60℃ for 1.5 h. Solid-liquid separation was then performed to obtain crude graphite and ferric phosphate solution. A The crude graphite was washed repeatedly with small amounts of 0.1 mol / L dilute hydrochloric acid, and then calcined at 600℃ under a nitrogen atmosphere to remove binders such as PVDF, resulting in relatively pure recycled graphite.

[0081] The above washing solution and ferric phosphate solution L A The mixture was then analyzed using potassium dichromate titration to determine the amount of ferrous ions in the lithium-containing leachate. Fe was added to the solution... 2+ Hydrogen peroxide with a molar ratio of 0.6 was reacted at room temperature for 20 min to completely convert it into Fe. 3+ The iron-phosphorus molar ratio was then adjusted to 1.0 using phosphoric acid.

[0082] Subsequently, 2 mol / L potassium hydroxide solution was added dropwise to precisely adjust the pH of the system to 2.0, and the reaction was continued at a constant temperature of 80℃ for 2 h. After the reaction was completed, crude ferric phosphate was obtained by solid-liquid separation. Then, the surface potassium chloride crystals were washed away with deionized water in small amounts several times to obtain amorphous ferric phosphate. It was placed in 0.5 mol / L phosphoric acid solution and aged at 80℃ for 4 h to prepare high-purity ferric phosphate dihydrate. Finally, it was calcined at 600℃ for 2 h to transform it into anhydrous ferric phosphate crystals with a purity of 98% and a recovery rate of 96%.

[0083] Example 2

[0084] (1) Circulating leaching

[0085] Prepare a FeCl3 solution as a leaching agent, mix it with the positive and negative electrode mixture powder of waste lithium iron phosphate batteries (the main components are the same as in Example 1), stir thoroughly, and complete the leaching reaction.

[0086] The lithium molar ratio in FeCl3 and the mixed powder was 1.30, and the solid-liquid ratio was 200 g·L⁻¹. -1 Under the conditions of reaction temperature of 40℃ and reaction time of 30min, the single leaching rate of Li was 98.03%, and the solid-liquid separation yielded lithium-containing leachate and leaching residue.

[0087] The amount of ferrous ions in lithium-containing leachate was determined using potassium dichromate titration. (Adding Fe...) 2+ Hydrochloric acid with a molar ratio of 1.0, and Fe 2+ Hydrogen peroxide with a molar ratio of 0.6 was used for oxidation at room temperature for 20 min, resulting in a near 100% regeneration rate of the leaching agent, yielding lithium-containing regenerated leaching agent FeCl3, which was then used to leach the next batch of material. This process was repeated for a total of four batches. The Li leaching rate was 89.28% in the fourth leaching. The specific effect of the number of leaching cycles on the lithium leaching effect is as follows: Figure 5 As shown.

[0088] Through four rounds of leaching, a fourth leaching solution (L4) and four batches of leaching residue were obtained. The four batches of leaching residue were then combined into two groups: the first and second batches together, and the third and fourth batches together. This step is crucial for controlling the appropriate solid-liquid ratio (100-600 g / L) to avoid limiting lithium re-leaching from the residue due to an excessively high solid-liquid ratio.

[0089] The two sets of leaching residues were leached sequentially using the fourth leaching solution L4, which was acidified and oxidized, without the need for further oxidation during the process. Leaching was completed at a leaching temperature of 40℃ and a leaching time of 30 min, followed by washing. The Li leaching rate was increased to 99.16%, and the total Cu leaching rate was 92.28%. The final lithium-rich leaching solution and the final leaching residue were obtained by filtration and separation. The total Li concentration in the lithium-rich leaching solution was 1.4 mol / L (per 100 mL), and the final leaching residue was a mixture of iron phosphate and graphite.

[0090] (2) Final leachate separation and purification

[0091] At room temperature, the final leachate was prepared with the following ratio: n(Fe):n(Cu). 2+ Reduced iron powder was added at a molar ratio of 1.1, and the displacement reaction was completed by continuous mechanical stirring for 30 min. Solid-liquid separation and subsequent analysis showed that the lithium-containing mother liquor (L) M Copper ions were completely removed (displacement rate 100%). The crude copper was then acid-washed with 2 mol / L hydrochloric acid at a solid-liquid ratio of 100 g / L, followed by mechanical stirring and washing at room temperature for 1 h to remove excess iron powder. After the above process, copper powder with a purity of 98.8% was finally obtained, with a recovery rate of 99%.

[0092] Inductively coupled plasma atomic emission spectrometry was used to detect the lithium-containing mother liquor after copper removal. M The mass concentration of each ion in the solution was determined, and the pH of the solution was measured to be 0.6 using a pH meter.

[0093] An extractant was prepared by di(2-ethylhexyl) phosphate (P204) diluted with kerosene, wherein the volume fraction of P204 was 30%. At room temperature, the extractant was mixed with a lithium-containing mother liquor diluted with water. M Mix (O / A=1:1) and place in a water bath constant temperature shaker for 30 min to obtain a mixed solution of iron-loaded extract phase and aluminum chloride and lithium chloride (raffinate phase), at which time the iron ion extraction rate reaches more than 99%.

[0094] At room temperature, the iron-loaded extractant phase was mixed with 6 mol / L hydrochloric acid (O / A = 1:1) and back-extracted in a water bath at 50°C for 30 min. After three-stage back-extraction, an empty organic phase and an iron-enriched aqueous phase were obtained. The empty organic phase could be reused for iron extraction; the iron-enriched aqueous phase mainly consisted of ferric chloride and hydrochloric acid, which was distilled under reduced pressure to obtain FeCl3, which was recycled as the leaching agent in the leaching step.

[0095] The extractant was prepared again, using the same kerosene-diluted di(2-ethylhexyl) phosphate (P204), with a P204 volume fraction of 30%. Then, 0.5 mol / L LiOH solution was added to the extractant for saponification, achieving a saponification degree of 40%. At room temperature, an equal volume of the saponified extractant was mixed with the raffinate (a mixed solution of aluminum chloride and lithium chloride) diluted with water, and the mixture was extracted for 30 min in a water bath with constant temperature shaking. After three stages of extraction, an aluminum-loaded extractant and a relatively pure lithium chloride solution (raffinate) were obtained.

[0096] At room temperature, the aluminum-enriched extractant phase was mixed with 6 mol / L hydrochloric acid (O / A = 1:1) and back-extracted in a water bath at 50°C for 30 min. After multiple back-extraction stages, an empty organic phase and an aluminum-enriched aqueous phase were obtained. The empty organic phase could be used again to extract aluminum; the aluminum-enriched aqueous phase mainly consisted of aluminum chloride and hydrochloric acid, which could be evaporated and crystallized to obtain aluminum chloride hexahydrate crystals.

[0097] (3) Final leaching residue separation and regeneration

[0098] A 3 mol / L dilute hydrochloric acid solution was added to the final leaching residue at a solid-liquid ratio of 100 g / L, and the reaction was carried out at 60℃ for 2 h. Solid-liquid separation was then performed to obtain crude graphite and ferric phosphate solution. A The crude graphite was washed repeatedly with small amounts of 0.1 mol / L dilute hydrochloric acid, and then calcined at 700℃ under a nitrogen atmosphere to remove binders such as PVDF, resulting in relatively pure recycled graphite.

[0099] The above washing solution and ferric phosphate solution L A The mixture was then analyzed using potassium dichromate titration to determine the amount of ferrous ions in the lithium-containing leachate. Fe was added to the solution... 2+ Hydrogen peroxide with a molar ratio of 0.6 was reacted at room temperature for 20 min to completely convert it into Fe. 3+ The iron-phosphorus ratio was then adjusted to 1.0 using phosphoric acid.

[0100] Subsequently, 2 mol / L potassium hydroxide solution was added dropwise to precisely adjust the pH of the system to 3.0, and the reaction was continued at a constant temperature of 80℃ for 2 h. After the reaction was completed, crude ferric phosphate was obtained by solid-liquid separation. Then, the surface potassium chloride crystals were washed away repeatedly with small amounts of deionized water to obtain amorphous ferric phosphate. This was then placed in 0.5 mol / L phosphoric acid solution and aged at 80℃ for 5 h to prepare high-purity ferric phosphate dihydrate. Finally, calcination at 600℃ for 2 h was performed to convert it into anhydrous ferric phosphate crystals with a purity of 96% and a recovery rate of 97%.

[0101] Example 3

[0102] (1) Circulating leaching

[0103] A Fe2(SO4)3 solution was prepared as a leaching agent and mixed with a mixture of positive and negative electrode powders from waste lithium iron phosphate batteries (the main components and contents are the same as in Example 1). The mixture was stirred thoroughly to complete the leaching reaction.

[0104] The lithium molar ratio in Fe2(SO4)3 and the mixed powder is 0.6, and the solid-liquid ratio is 100 g·L. -1 Under the conditions of reaction temperature 40 ℃ and reaction time 40 min, the single leaching rate of Li was 95.87%, and the solid-liquid separation yielded lithium-containing leachate and leaching residue.

[0105] The amount of ferrous ions in lithium-containing leachate was determined using potassium dichromate titration. (Adding Fe...) 2+ Sulfuric acid with a molar ratio of 0.5, and its reaction with Fe. 2+ Hydrogen peroxide with a molar ratio of 0.6 was used for oxidation at room temperature for 30 min, resulting in a near 100% regeneration rate of the leaching agent, yielding a lithium-containing regenerated leaching agent, Fe2(SO4)3, which was then used to leach the next batch of material. This process was repeated for a total of three batches. The Li leaching rate was 74.31% in the third leaching. The specific effect of the number of leaching cycles on the leaching efficiency of lithium and copper is as follows: Figure 6 As shown.

[0106] Three rounds of leaching were performed to obtain the third leaching solution L3 and three batches of leaching residue. The three batches of leaching residue were combined into one group. After acidifying and oxidizing the third leaching solution L3, the above-mentioned group of leaching residue was leached at a leaching temperature of 40℃ and a leaching time of 40 min. After washing, the Li leaching rate was increased to 98.26%. Filtration and separation yielded the final lithium-rich leaching solution and the final leaching residue. The total Li concentration in the lithium-rich leaching solution was 1.0 mol / L (based on 100 mL), and the final leaching residue was a mixture of ferric phosphate and graphite.

[0107] (2) Final leachate separation and purification

[0108] The composition and content of the final leachate were determined using inductively coupled plasma atomic emission spectrometry (ICP-AES). At room temperature, the final leachate was added in the following ratio: n(Fe):n(Cu). 2+ Reducing iron powder was added at a molar ratio of 1.05, and the displacement reaction was completed by continuous mechanical stirring for 30 min. After solid-liquid separation and analysis, the mother liquor contained lithium (L). M Copper ions were completely removed (displacement rate 100%). The resulting crude copper was then acid-washed with 1 mol / L sulfuric acid at a solid-liquid ratio of 100 g / L, followed by mechanical stirring and washing at room temperature for 1 h to remove excess iron powder. After the above process, copper powder with a purity of 99.7% was finally obtained, with a recovery rate of 96.1%.

[0109] Inductively coupled plasma atomic emission spectrometry was used to detect the lithium-containing mother liquor after copper removal. M The mass concentration of each ion in the solution was determined, and the pH of the solution was measured to be 0.6 using a pH meter.

[0110] An extractant was prepared by di(2-ethylhexyl) phosphate (P204) diluted with kerosene, wherein the volume fraction of P204 was 40%. At room temperature, the extractant was mixed with a lithium-containing mother liquor diluted with water. F Mix (O / A=1:1) and extract in a water bath constant temperature shaker for 30 min to obtain a mixed solution of iron-loaded extract phase and aluminum sulfate and lithium sulfate (raffinate phase). At this time, the iron ion extraction rate reaches more than 99%.

[0111] At room temperature, the iron-loaded extractant phase was mixed with 1.5 mol / L sulfuric acid (O / A = 1:1) and back-extracted in a water bath with constant temperature shaker for 30 min. After three-stage back-extraction, an empty organic phase and an iron-enriched aqueous phase were obtained. The empty organic phase could be reused for iron extraction; the iron-enriched aqueous phase, mainly composed of ferric sulfate, was recycled as the leaching agent in the leaching step.

[0112] The extractant was prepared again, using the same kerosene-diluted di(2-ethylhexyl) phosphate (P204), with a P204 volume fraction of 40%. Then, 0.5 mol / L LiOH solution was added to the extractant for saponification, achieving a saponification degree of 40%. At room temperature, an equal volume of the saponified organic extract phase was mixed with the raffinate phase diluted with water (a mixed solution of aluminum sulfate and lithium sulfate), and extracted in a water bath with a constant temperature shaker for 30 min. After three stages of extraction, an aluminum-loaded extract phase and a relatively pure lithium sulfate solution (raffinate phase) were obtained.

[0113] At room temperature, the aluminum-loaded extractant phase was mixed with 1.5 mol / L sulfuric acid (O / A = 1:1) and back-extracted in a water bath with constant temperature shaker for 30 min. After multi-stage back-extraction, an empty organic phase and an aluminum-enriched aqueous phase were obtained. The empty organic phase could be used again to extract aluminum; the aluminum-enriched aqueous phase mainly consisted of aluminum sulfate and sulfuric acid, which could be evaporated and crystallized to obtain aluminum sulfate crystals with water of crystallization.

[0114] (3) Final leaching residue separation and regeneration

[0115] A 1.5 mol / L dilute sulfuric acid solution was added to the final leaching residue at a solid-liquid ratio of 200 g / L, and the reaction was carried out at 60℃ for 2 h. Solid-liquid separation was then performed to obtain crude graphite and ferric phosphate solution. A The crude graphite was washed repeatedly with small amounts of 0.05 mol / L dilute sulfuric acid, and then calcined at 600°C under a nitrogen atmosphere to remove binders such as PVDF, resulting in relatively pure recycled graphite.

[0116] The above washing solution and ferric phosphate solution L A The mixture was then analyzed using potassium dichromate titration to determine the amount of ferrous ions in the lithium-containing leachate. Fe was added to the solution... 2+ Hydrogen peroxide with a molar ratio of 0.6 was reacted at room temperature for 20 min to completely convert it into Fe. 3+ The iron-phosphorus molar ratio was then adjusted to 1.0 using phosphoric acid.

[0117] Subsequently, 2 mol / L potassium hydroxide solution was added dropwise to precisely adjust the pH of the system to 2.0, and the reaction was continued at a constant temperature of 80℃ for 2 h. After the reaction was completed, crude ferric phosphate was obtained by solid-liquid separation. Then, the surface potassium sulfate crystals were washed away with deionized water in small amounts several times to obtain amorphous ferric phosphate. It was placed in 0.5 mol / L phosphoric acid solution and aged at 80℃ for 4 h to prepare high-purity ferric phosphate dihydrate. Finally, it was calcined at 600℃ for 2 h to transform it into anhydrous ferric phosphate crystals with a purity of 98% and a recovery rate of 96%.

[0118] The above-described embodiments are merely illustrative of several implementations of the present invention, and while the descriptions are specific and detailed, they should not be construed as limiting the scope of the present invention. Those skilled in the art can make various modifications and improvements without departing from the concept of the present invention, and these modifications and improvements all fall within the scope of protection of the present invention.

Claims

1. A method for cyclic leaching of lithium iron phosphate mixed powder with iron salts and recovery of all components, characterized in that, Includes the following steps: S1. Using an iron salt solution as a leaching agent, the waste lithium iron phosphate positive and negative electrode mixed powder is thoroughly mixed and stirred to carry out the leaching reaction, resulting in leaching solution and leaching residue; the iron salt is ferric chloride or ferric sulfate. S2. The leachate is regenerated by acidification and oxidation. The resulting regenerated leachate is returned to S1 to leach the next batch of positive and negative electrode mixed powder. The acidification and oxidation regeneration and leaching are repeated until the regenerated leachate reaches the predetermined number of regenerations. S3. The leaching residue obtained by combining S1 and S2 is divided into several groups. Each group of leaching residue is then subjected to a re-leaching reaction with the regenerating leaching agent after S2 has reached the predetermined number of regenerations, to obtain the final leachate and the final leaching residue. S4. Add iron powder to the final leachate to replace copper ions, and obtain crude copper and lithium-containing mother liquor; add acid to the crude copper, and separate the solid and liquid to obtain industrial-grade copper powder and ferrous ion solution; oxidize the ferrous ion solution, and then distill it under reduced pressure to obtain an iron salt solution as a leaching agent. S5. Oxidize the lithium-containing mother liquor from S4. Add an extractant to the oxidized lithium-containing mother liquor and perform multi-stage selective extraction of iron and aluminum to obtain an extract phase loaded with iron ions, an extract phase loaded with aluminum ions, and a Li-containing extract phase. + Raffinate solution; the two types of extractable phases are back-extracted and post-treated to recover iron and aluminum; S6. Add acid to the final leaching residue of S3 to dissolve ferric phosphate, and separate the solid and liquid to obtain crude graphite and ferric phosphate solution; the crude graphite is washed and calcined to obtain regenerated graphite; phosphorus or iron source is added to the ferric phosphate solution to adjust the iron-phosphorus material ratio, and then hydrogen peroxide is added to oxidize the ferrous ions, and then alkali solution is added to react fully to obtain crude ferric phosphate; the crude ferric phosphate is washed, aged and crystallized, and calcined to prepare anhydrous ferric phosphate.

2. The method according to claim 1, characterized in that: The waste lithium iron phosphate positive and negative electrode mixed powder is obtained by discharging and mechanically dismantling waste lithium iron phosphate batteries. In addition to lithium iron phosphate as the positive electrode material and graphite as the negative electrode material, it contains one or more of the following: binder, current collector, and electrolyte.

3. The method according to claim 1, characterized in that: In step S1, the iron ions in the leaching agent react with the Li in the lithium iron phosphate positive and negative electrode powders. + The molar ratio is 1.0-2.0; the solid-liquid ratio of lithium iron phosphate positive and negative electrode powder to leachate solution is 100-600 g / L; the reaction temperature is 20-60℃; and the reaction time is 20-60 min.

4. The method according to claim 1, characterized in that: The acid used for acidification in step S2 is hydrochloric acid or sulfuric acid, and its molar amount of hydrogen ions is equal to the amount of Fe in the leachate. 2+ The molar amount is 1.0-1.5 times; the oxidant used for oxidation is hydrogen peroxide, with a hydrogen peroxide mass concentration of 10%-30%, and the amount used is equal to the Fe content in the solution. 2+ The molar amount is 0.5-1.0 times, the oxidation time is 5-30 min, and the oxidation temperature is 20-40℃.

5. The method according to claim 1, characterized in that: In step S3, the regenerated leaching agent sequentially and continuously re-leaches each group of leaching residues without re-oxidizing the leaching solution. The solid-liquid ratio of each group of re-leaching reactions is 100-600 g / L, the reaction temperature is 20-60℃, and the reaction time is 20-60 min.

6. The method according to claim 1, characterized in that: In step S4, the amount of iron powder is 1.0-2.0 times the copper ion content in the solution; the hydrogen ion concentration in the acid is 0.5-5.0 mol / L; the oxidant used in the ferrous oxide solution is hydrogen peroxide with a mass concentration of 10%-30%, and the amount used is equal to the Fe content in the solution. 2+ The molar amount is 0.5-1.0 times, the oxidation time is 5-30 min, and the oxidation temperature is 20-40℃.

7. The method according to claim 1, characterized in that: In step S5, the mother liquor containing lithium has a pH < 1. An extractant is added to it to obtain an extract phase loaded with iron ions and a raffinate phase containing aluminum and lithium salts. For the treatment of the raffinate phase containing aluminum and lithium salts, LiOH solution is first added to the extractant for saponification, with a saponification degree of 40%-80%. The saponified extractant is then subjected to multi-stage extraction with the raffinate phase containing aluminum and lithium salts to obtain an extractant phase loaded with aluminum ions and a Li-containing raffinate phase. + Raffinate solution; The extractants were all kerosene-diluted di(2-ethylhexyl) phosphate, with a volume fraction of 20%-40% in the extractants; the volume ratio of organic phase to aqueous phase in each extraction stage was (1-3):

1.

8. The method according to claim 1, characterized in that: In step S5, the two types of extractable phases are back-extracted and post-treated to recover iron and aluminum, specifically including: The iron-loaded extraction phase is subjected to acid multi-stage back-extraction to obtain an empty organic phase and an iron-containing aqueous phase. The empty organic phase is recovered as an extractant for cyclic extraction of iron, and the aqueous phase is subjected to vacuum distillation to obtain an iron salt leaching agent. The aluminum-loaded extraction phase is subjected to acid multi-stage back-extraction to obtain an empty organic phase and an aluminum-containing aqueous phase. The empty organic phase is recovered as an extractant for cyclic extraction of aluminum, and the aqueous phase is evaporated and crystallized to obtain aluminum salt crystals. The acids used in all back-extraction processes are hydrochloric acid or sulfuric acid, with hydrogen ion concentrations ranging from 2 to 8 mol / L. During the back-extraction process, the volume ratio of the organic phase to the aqueous phase is 1:(1-3).

9. The method according to claim 1, characterized in that: In step S6, the phosphorus source is one or more of ammonium dihydrogen phosphate, diammonium hydrogen phosphate, and phosphoric acid, with an iron-phosphorus molar ratio of 0.9-1.2; the iron source is ferric chloride or ferric sulfate; the alkaline solution is one or more of potassium hydroxide, sodium hydroxide, and ammonia water, with a concentration of 1.0-3.0 mol / L, and the pH is adjusted to 1.0-4.0 with the alkaline solution. The reaction temperature is 50℃-80℃, and the reaction time is 2-4 h.

10. The method according to claim 1, characterized in that: In step S6, the crude graphite, after acid washing, is calcined in a nitrogen atmosphere at a temperature of 450℃-700℃ for 1-3 hours. The crude ferric phosphate was washed with deionized water and aged and crystallized using dilute phosphoric acid with a concentration of 0.1-1.0 mol / L for 2-10 h. The calcination temperature of the ferric phosphate was 300℃-600℃ and the calcination time was 1-3 h.

Images