Method for short-process recovery of battery-grade lithium carbonate and valuable metals from waste ternary lithium battery black powder

Abstract

The invention belongs to the technical field of waste battery recovery, and particularly relates to a short-process method for recovering battery-grade lithium carbonate and valuable metals from waste ternary lithium battery black powder. According to the method, a lithium closed-loop technical system of blank roasting, cyclic leaching, synergistic impurity removal, lithium phosphate aluminum removal, directional extraction and split-flow extraction of lithium, nickel and magnesium and seed crystal induction parallel flow lithium precipitation is constructed, a roasting aid, a reducing agent and an aluminum removal agent do not need to be added in the whole process, and the redundant processes of traditional lithium carbonate purification such as carbonization, resin purification and pyrolysis are omitted; the lithium liquid concentration cost is saved, no hazardous waste exists, green and clean production is achieved, and the quality of battery-grade products is guaranteed through collaborative impurity removal and directional extraction.

Description

Technical Field

[0001] The present invention belongs to the technical field of waste battery recycling, and specifically relates to a method for recovering battery-grade lithium carbonate and valuable metals from black powder of waste ternary lithium batteries in a short process. Background Art

[0002] Waste ternary lithium batteries are rich in valuable metals such as lithium, cobalt, nickel, copper, and manganese. Efficiently recycling these metals and producing high-purity battery-grade products is crucial for reducing industry costs and minimizing resource waste. However, existing waste ternary lithium battery recycling processes still have many shortcomings in terms of process optimization, cost control, environmental benefits, and product quality, and breakthroughs are urgently needed.

[0003] Traditional roasting processes for waste ternary lithium battery black powder mainly include sulfate roasting and sodium carbonate roasting. The sulfate roasting process uses concentrated sulfuric acid, which easily causes agglomeration during the mixing process and produces a large amount of sulfur dioxide. The sodium carbonate roasting process requires the addition of additives such as sodium carbonate to promote lithium conversion. The roasting temperature is high (800-1000°C), which is not only costly but also introduces sodium impurities, requiring a complex subsequent sodium removal process.

[0004] The traditional crude lithium carbonate purification process involves a series of steps: crude lithium carbonate → carbonization → resin removal of calcium and magnesium → pyrolysis. Calcium and magnesium ions in the carbonized liquid must be removed by resin adsorption, requiring frequent resin regeneration, increasing costs and reducing production efficiency. This complex process is associated with high lithium loss rates and energy consumption (pyrolysis temperatures reach 80-100°C). Furthermore, impurities such as sodium and sulfate are difficult to remove (e.g., sodium content exceeding 0.05%), leading to significant fluctuations in product quality.

[0005] Existing impurity removal processes are mostly step-by-step, resulting in high reagent consumption and high metal loss. For example, aluminum sulfate is commonly used for fluoride ion removal, while heavy metal removal relies on sodium sulfide. The use of multiple reagents significantly increases costs. High-pH aluminum removal (such as using sodium carbonate to adjust the pH to 5-6) can easily lead to the loss of valuable metals such as cobalt, nickel, and manganese as hydroxide precipitates. The aluminum removal slag has a high moisture content and poor filterability, increasing the difficulty and cost of solid waste treatment and resulting in resource waste. Furthermore, conventional roasting slag for lithium extraction with sulfuric acid leaching often requires the addition of reducing agents such as sodium metabisulfite to reduce the high-valent metals, generating sulfur dioxide pollution and high environmental treatment costs.

[0006] During the recovery process, lithium is dispersed in the leachate, lithium precipitation mother liquor, raffinate and other links. The existing process lacks effective closed-loop enrichment means, resulting in a total lithium recovery rate of less than 90%; and the lithium concentration in the lithium extraction leachate and raffinate is often less than 5g / L. Traditional evaporation and concentration treatment has high energy consumption, large equipment investment and high cost, making it difficult for enterprises to recycle economically.

[0007] The recovery of metals such as cobalt, nickel, copper, manganese, and magnesium often requires additional purification to reach battery-grade standards due to poorly optimized extraction processes. In traditional extraction processes, improper selection of extractants and improper process parameter settings can easily lead to excessive impurity levels in the product (e.g., nickel content exceeding 0.05% in cobalt sulfate), limiting its application in high-end battery production and reducing the added value of the recycled product.

[0008] In view of the various problems existing in the existing technology for the recycling of waste ternary lithium batteries, there is an urgent need to develop a method for recycling battery-grade lithium carbonate and valuable metals from waste ternary lithium battery black powder in a short process. Summary of the Invention

[0009] The purpose of the present invention is to provide a method for recovering battery-grade lithium carbonate and valuable metals from waste ternary lithium battery black powder in a short process. The method provided by the present invention has a short process and high production efficiency.

[0010] In order to achieve the above object, the present invention provides the following technical solutions: The present invention provides a method for recovering battery-grade lithium carbonate and valuable metals from waste ternary lithium battery black powder in a short process, comprising the following steps: (1) Calcination of waste ternary lithium battery black powder in a natural gas atmosphere to obtain calcined slag; (2) subjecting the roasted slag to low-acid circulating leaching and filter pressing using a dilute sulfuric acid solution to obtain a lithium-rich circulating leachate and low-acid leaching slag; (3) When the lithium concentration in the lithium-rich circulating leachate is ≥20 g / L, a portion of the lithium-rich circulating leachate is diverted and mixed with calcium oxide, and then subjected to defluorination and deweighting and filter pressing in sequence to obtain a defluorination and deweighting liquid; (4) extracting and removing calcium from the defluorination and deheavy-weight removal liquid to obtain a decalcified liquid; (5) adding zirconium hydroxide to the decalcification liquid to deeply remove silicon, phosphorus and fluorine to obtain a silicon, phosphorus and fluorine-removing liquid; (6) Using pure water as the base liquid, adding battery-grade lithium carbonate as a seed crystal, adding the silicon, phosphorus and fluorine removal liquid and sodium carbonate solution into the synthesis kettle in a parallel flow form to induce lithium precipitation, thereby obtaining lithium carbonate and lithium precipitation mother liquor; (7) adding trisodium phosphate to the lithium precipitation mother liquor for synthesis to obtain crude lithium phosphate; (8) subjecting the low-acid leaching residue to high-acid leaching and filter pressing in sequence to obtain high-acid leaching residue and high-acid leaching liquid; (9) adding manganese dioxide to the high acid leaching solution for oxidation, adding the obtained oxidized solution to the crude lithium phosphate product for iron and aluminum removal and filter pressing to obtain an iron and aluminum removal solution and an iron and aluminum removal slag; (10) extracting copper from the iron and aluminum-removed liquid to obtain copper raffinate and battery-grade copper sulfate; (11) extracting the copper raffinate to obtain a decontaminated solution and a manganese solution; the obtained manganese solution is subjected to deep purification in sequence to obtain battery-grade manganese sulfate; the deep purification is carried out in sequence by sulfidation to remove weight, remove calcium and remove fluoride; (12) extracting cobalt from the impurity-removed liquid to obtain battery-grade cobalt sulfate and a cobalt raffinate containing low-concentration lithium, nickel, and magnesium; extracting lithium, nickel, and magnesium from the cobalt raffinate containing low-concentration lithium, nickel, and magnesium, and re-entering the cobalt extraction section for extraction and separation to obtain a raffinate containing high-concentration lithium, nickel, and magnesium; (13) extracting magnesium from the raffinate containing high concentrations of lithium, nickel and magnesium to obtain magnesium raffinate and battery-grade magnesium sulfate; (14) The magnesium raffinate is subjected to nickel extraction to obtain lithium-rich solution and battery-grade nickel sulfate; the lithium-rich solution is reused for fluorine removal and weight removal in step (3).

[0011] Preferably, the blank is calcined at a temperature of 450-650° C. and for a time of 1-3 hours.

[0012] Preferably, during the low-acid circulation leaching, the leachate obtained from each low-acid leaching is acidified and then reused for low-acid circulation leaching, and the pH of the dilute sulfuric acid solution or the leachate after acid adjustment is 1-3; during each low-acid leaching, the liquid-to-solid ratio is 10-20 ml / g, and the time of each low-acid leaching is 1-4 hours.

[0013] Preferably, the time for fluorine removal and weight removal is 1 to 4 hours; the temperature is 25 to 50° C.; the ratio of the added molar amount of calcium oxide to the molar amount of fluorine in the partial lithium-rich circulating leachate is 1.1 to 1.3:1; the molar amount of fluorine is the sum of 1 / 2 of the molar amount of fluoride ions and the molar amount of heavy metal ions; the heavy metals include one or more of Co, Ni, Mn and Cu.

[0014] Preferably, the extractant for extracting and removing calcium is di(2-ethylhexyl) phosphate; the amount of zirconium hydroxide added is 20 to 60 times the total mass of silicon, phosphorus and fluorine in the decalcification solution; and the time for deep removal of silicon, phosphorus and fluorine is 1 to 3 hours.

[0015] Preferably, in step (6), the molar ratio of sodium carbonate in the sodium carbonate solution to the Li element in the silicon, phosphorus and fluorine removal solution is 1.1-1.2:2, the concentration of the sodium carbonate solution is 250-350 g / L; the amount of battery-grade lithium carbonate added is 5-15 wt.% of the mass of the synthesized lithium carbonate, and the temperature for inducing lithium precipitation is 80-100°C.

[0016] Preferably, the ratio of the added molar amount of the trisodium phosphate to the molar amount of the Li element in the lithium precipitation mother liquor is 1.1-1.2:3, the synthesis temperature of the crude lithium phosphate is 80-100° C., and the synthesis time is 1-3 hours.

[0017] Preferably, the initial concentration of the sulfuric acid solution for high acid leaching is 4-7 mol / L; the liquid-solid ratio of high acid leaching is 3-5 mL / g; the temperature of high acid leaching is 60-90° C., and the time of high acid leaching is 3-6 h.

[0018] Preferably, the ratio of the molar amount of manganese dioxide added to the molar amount of Fe element in the high-acid leaching solution is 1.1~1.2:2, the oxidation time is 1~2h, and the ratio of the molar amount of the added crude lithium phosphate to the total molar amount of iron ions and aluminum ions in the oxidized solution is 1~1.2:1; when removing iron and aluminum, the pH value of the system is 3~4, and the iron and aluminum removal reaction time is 1~3h.

[0019] Preferably, the extractant for copper extraction is a mixture of 5-dodecyl salicylaldehyde oxime and 2-hydroxy-5-nonylacetophenone oxime; the extractant for impurity extraction is di(2-ethylhexyl) phosphate; the extractant for cobalt extraction is 2-ethylhexyl mono-2-ethylhexyl phosphate; the extractant for magnesium extraction is di(2,4,4-trimethylpentyl)phosphinic acid; and the extractant for nickel extraction is 2-ethylhexyl mono-2-ethylhexyl phosphate.

[0020] Compared with the prior art, the beneficial effects of the technical solution of the present invention include at least: (1) The blank roasting process of waste ternary lithium battery black powder does not require roasting aids such as sodium carbonate or sulfuric acid. Instead, the carbon dioxide generated by the sufficient combustion of natural gas is used to efficiently convert the lithium oxide in the black powder into lithium carbonate at high temperature. The lithium carbonate yield is higher than 90%, while avoiding the introduction of sodium impurities and saving the subsequent sodium removal process. The process is simplified, the cost is reduced, and the energy consumption is significantly reduced.

[0021] (2) The present invention optimizes the lithium carbonate synthesis process by adopting a parallel flow lithium precipitation + seed induction + hot water washing method to avoid the formation of lithium sodium sulfate complex salt impurities during the lithium precipitation process, while also avoiding local supersaturation of lithium carbonate. This effectively reduces the sodium and sulfate content in the synthesized lithium carbonate, reduces the difficulty of washing, and improves filtration performance. The lithium carbonate after hot water washing meets battery-grade product standards. The lithium precipitation process eliminates the traditional redundant process of "carbonization-resin calcium and magnesium removal-pyrolysis" and directly produces battery-grade lithium carbonate in a short process, greatly shortening the lithium carbonate preparation cycle, reducing costs and energy consumption, and improving production efficiency.

[0022] (3) Low-acid circulation leaching and diversion of blank roasted slag achieve lithium solution concentration enrichment, solve the problem of low solubility of lithium carbonate, and enrich the lithium concentration in the leachate to more than 20g / L. At the same time, the cobalt extraction process adds a multi-stage collaborative lithium, nickel and magnesium extraction tank in front of the cobalt extraction section, so that the raffinate containing low concentration of lithium, nickel and magnesium in the cobalt extraction section is partially diverted into the lithium, nickel and magnesium extraction section for lithium, nickel and magnesium extraction, and the organic phase loaded with lithium, nickel and magnesium is returned to the cobalt extraction section for extraction and separation to obtain a raffinate containing high concentration of lithium, nickel and magnesium. The lithium concentration of the raffinate containing high concentration of lithium, nickel and magnesium can be enriched to more than 10g / L. The lithium solution enrichment process of low-acid circulation leaching and cobalt extraction and diversion extraction of lithium, nickel and magnesium of the present invention solves the problem of high cost of low-concentration lithium concentration, significantly improves lithium precipitation efficiency, reduces lithium recovery cost, and reduces energy consumption of subsequent evaporation and concentration.

[0023] (4) Calcium oxide is used to simultaneously remove fluorine and heavy metal impurities, achieving the removal of multiple impurities in one step; combined with the graded purification process of calcium extraction and zirconium hydroxide adsorption of silicon, phosphorus and fluorine, impurities such as calcium, silicon, phosphorus and fluorine are efficiently removed, laying the foundation for the preparation of high-purity battery-grade lithium carbonate products. Low-acid leaching residue acid leaching uses a high-temperature acidic environment to cause the metal oxides to self-oxidize and dissolve, eliminating the need for reducing agents, reducing sulfur dioxide pollution, and the high-acid leaching residue meets the discharge standards after washing; lithium phosphate prepared from lithium precipitation mother liquor is used in the high-acid leaching liquid iron and aluminum removal process, removing iron and aluminum under acidic conditions (pH 3~4), with good iron and aluminum removal effect, eliminating the need for the addition of traditional aluminum removal agents such as sodium carbonate and calcium carbonate. After washing, the cobalt, nickel, manganese, copper and lithium metal entrainment in the iron and aluminum removal slag is less than 0.1%, realizing the resource utilization of the mother liquor, significantly reducing the aluminum removal cost, improving the filterability of the aluminum removal slag, increasing the aluminum removal capacity, and achieving the iron and aluminum removal slag meeting the discharge standards; (5) Through the directional solvent extraction process, the copper extraction, impurity extraction and manganese solution deep purification, cobalt extraction, magnesium extraction, nickel extraction and other processes are optimized to prepare battery-grade products such as copper sulfate, manganese sulfate, cobalt sulfate, magnesium sulfate, and nickel sulfate, respectively. The product purity is improved (≥99.9%), the metal recovery rate of cobalt, nickel, manganese, copper, and magnesium is increased, the process flow is shortened, and production efficiency is improved. The lithium-rich solution obtained from the P507 nickel extraction is returned to the defluorination and deweighting process in a closed loop and enters the lithium precipitation system, thereby improving the lithium recovery rate and realizing the efficient recycling of lithium resources.

[0024] In summary, the present invention has a short process, simple technology, high production efficiency, low energy consumption, green and clean production, no hazardous waste generation, and excellent product quality. During the recycling process, the recovery rate of valuable metals such as cobalt, nickel, copper, manganese, lithium, and magnesium reaches more than 99%, and the valuable metal products all meet battery-grade standards, with obvious economic and environmental benefits. BRIEF DESCRIPTION OF THE DRAWINGS

[0025] In order to more clearly illustrate the embodiments of the present invention or the technical solutions in the prior art, the following briefly introduces the drawings required for use in the embodiments. Obviously, the drawings described below are only some embodiments of the present invention. For ordinary technicians in this field, other drawings can be obtained based on these drawings without paying any creative work.

[0026] Figure 1 The present invention provides a flow chart for the short-process recovery of battery-grade lithium carbonate and valuable metals from waste ternary lithium battery black powder. DETAILED DESCRIPTION

[0027] The present invention provides a method for recovering battery-grade lithium carbonate and valuable metals from waste ternary lithium battery black powder in a short process, comprising the following steps: (1) Calcination of waste ternary lithium battery black powder in a natural gas atmosphere to obtain calcined slag; (2) subjecting the roasted slag to low-acid circulating leaching and filter pressing using a dilute sulfuric acid solution to obtain a lithium-rich circulating leachate and low-acid leaching slag; (3) When the lithium concentration in the lithium-rich circulating leachate is ≥20 g / L, a portion of the lithium-rich circulating leachate is diverted and mixed with calcium oxide, and then subjected to defluorination and deweighting and filter pressing in sequence to obtain a defluorination and deweighting liquid; (4) extracting and removing calcium from the defluorination and deheavy-weight removal liquid to obtain a decalcified liquid; (5) adding zirconium hydroxide to the decalcification liquid to deeply remove silicon, phosphorus and fluorine to obtain a silicon, phosphorus and fluorine-removing liquid; (6) Using pure water as the base liquid, adding battery-grade lithium carbonate as a seed crystal, adding the silicon, phosphorus and fluorine removal liquid and sodium carbonate solution into the synthesis kettle in a parallel flow form to induce lithium precipitation, thereby obtaining lithium carbonate and lithium precipitation mother liquor; (7) adding trisodium phosphate to the lithium precipitation mother liquor for synthesis to obtain crude lithium phosphate; (8) subjecting the low-acid leaching residue to high-acid leaching and filter pressing in sequence to obtain high-acid leaching residue and high-acid leaching liquid; (9) adding manganese dioxide to the high acid leaching solution for oxidation, adding the obtained oxidized solution to the crude lithium phosphate product for iron and aluminum removal and filter pressing to obtain an iron and aluminum removal solution and an iron and aluminum removal slag; (10) extracting copper from the iron and aluminum-removed liquid to obtain copper raffinate and battery-grade copper sulfate; (11) extracting the copper raffinate to obtain a decontaminated solution and a manganese solution; the obtained manganese solution is subjected to deep purification in sequence to obtain battery-grade manganese sulfate; the deep purification is carried out in sequence by sulfidation to remove weight, remove calcium and remove fluoride; (12) extracting cobalt from the impurity-removed liquid to obtain battery-grade cobalt sulfate and a cobalt raffinate containing low-concentration lithium, nickel, and magnesium; extracting lithium, nickel, and magnesium from the cobalt raffinate containing low-concentration lithium, nickel, and magnesium, and re-entering the cobalt extraction section for extraction and separation to obtain a raffinate containing high-concentration lithium, nickel, and magnesium; (13) extracting magnesium from the raffinate containing high concentrations of lithium, nickel and magnesium to obtain magnesium raffinate and battery-grade magnesium sulfate; (14) The magnesium raffinate is subjected to nickel extraction to obtain lithium-rich solution and battery-grade nickel sulfate; the lithium-rich solution is reused for fluorine removal and weight removal in step (3).

[0028] The invention performs blank roasting on waste ternary lithium battery black powder in a natural gas atmosphere to obtain roasted slag.

[0029] As an embodiment of the present invention, the blank calcination temperature can be 450~650℃, specifically 450℃, 500℃, 550℃, 600℃ or 650℃; the blank calcination time is 1~3h, specifically 1h, 2h or 3h.

[0030] As an embodiment of the present invention, blank roasting can efficiently convert lithium into lithium carbonate.

[0031] The invention uses a dilute sulfuric acid solution to carry out low-acid circulation leaching and filter pressing on the roasted slag to obtain lithium-rich circulation leaching solution and low-acid leaching slag.

[0032] As an embodiment of the present invention, during low-acid cyclic leaching, the leachate obtained by each low-acid leaching needs to be acid-adjusted and then returned to the circulation leaching of new roasted slag, and the pH of the dilute sulfuric acid solution and the leachate after acid adjustment are both 1-3, specifically 2; during each low-acid leaching, the liquid-to-solid ratio is 10-20 mL / g, and the leaching time can be 1-4 hours, specifically 1 hour, 2 hours, 3 hours or 4 hours.

[0033] When the lithium concentration in the lithium-rich circulating leachate is greater than or equal to 20 g / L, the present invention diverts part of the lithium-rich circulating leachate and mixes it with calcium oxide, and sequentially performs defluorination and deweighting and pressure filtration to obtain a defluorination and deweighting liquid.

[0034] As an embodiment of the present invention, the reaction occurring during the defluorination and weight removal process is: MeSO4+CaO+H2O=Me(OH)2+CaSO4 (Me: one or more of Co, Ni, Mn, Cu); 2HF+ CaO=CaF2↓+H2O.

[0035] As an embodiment of the present invention, the ratio of the added molar amount of calcium oxide to the molar amount of fluorine in the partial lithium-rich circulating leachate is 1.1~1.3:1; the molar amount of fluorine is the sum of 1 / 2 of the molar amount of fluoride ions and the molar amount of heavy metal ions, specifically 1.1:1, 1.2:1 or 1.3:1; the time for defluorination and weight removal can be 1~4h, specifically 1h, 2h, 3h or 4h; the temperature can be 25~50°C, specifically 25°C, 35°C, 45°C or 50°C.

[0036] After obtaining the defluorination and deheaving liquid, the present invention extracts and removes calcium from the defluorination and deheaving liquid to obtain a decalcification liquid.

[0037] As an embodiment of the present invention, the extractant for extracting and removing calcium may be di(2-ethylhexyl) phosphate (P204).

[0038] As an embodiment of the present invention, during the extraction and calcium removal, the extraction phase ratio can be 4:1.

[0039] As an embodiment of the present invention, the extraction and calcium removal further obtains a calcium-loaded organic phase, and the calcium-loaded organic phase is back-extracted with a 1-3 mol / L hydrochloric acid solution.

[0040] After obtaining the decalcification liquid, the present invention adds zirconium hydroxide to the decalcification liquid to deeply remove silicon, phosphorus and fluorine to obtain a desiliconized, phosphorus and fluorine-removing liquid.

[0041] As an embodiment of the present invention, the amount of zirconium hydroxide added is 20 to 60 times the total mass of silicon, phosphorus and fluorine in the decalcification liquid, specifically 30 to 40 times; the time for deep removal of silicon, phosphorus and fluorine can be 1 to 3 hours, specifically 1 hour; the method of deep removal is adsorption; as an embodiment of the present invention, the zirconium hydroxide is desorbed and regenerated with 5 to 10 wt.% liquid alkali after adsorption saturation.

[0042] After obtaining the silicon, phosphorus and fluorine removal liquid, the present invention uses pure water as the base liquid, adds battery-grade lithium carbonate as a seed crystal, and adds the silicon, phosphorus and fluorine removal liquid and sodium carbonate solution into a synthesis kettle in a parallel flow form to induce lithium precipitation (synthesize lithium carbonate) to obtain battery-grade lithium carbonate and lithium precipitation mother liquor.

[0043] As an embodiment of the present invention, the reaction occurring during the induction lithium precipitation process is: Li2SO4+Na2CO3=Li2CO3+Na2SO4.

[0044] As an embodiment of the present invention, the molar ratio of sodium carbonate in the sodium carbonate solution and the Li element in the silicon, phosphorus and fluorine removal liquid is 1.1~1.2:2, and the concentration of the sodium carbonate solution can be 250~350g / L, specifically 250g / L, 300g / L or 350g / L; the amount of lithium carbonate seed added is 5~15wt.% of the mass of the synthesized lithium carbonate, and the temperature for inducing lithium precipitation (synthetic lithium carbonate) can be 80~100℃, specifically 80℃, 90℃ or 100℃; as an embodiment of the present invention, after inducing lithium precipitation to obtain lithium carbonate, the synthesized lithium carbonate is further washed with hot water three times, the washing temperature is 80~100℃, and the washing time for each washing is 1~2h.

[0045] After obtaining the lithium precipitation mother liquor, the present invention adds the lithium precipitation mother liquor into trisodium phosphate for synthesis to prepare crude lithium phosphate.

[0046] As an embodiment of the present invention, the reaction occurring during the synthesis of crude lithium phosphate is: 3Li2SO4+2Na3PO4=2Li3PO4+3Na2SO4.

[0047] As an embodiment of the present invention, the ratio of the added molar amount of trisodium phosphate to the molar amount of Li element in the lithium precipitation mother liquor is 1.1~1.2:3, the synthesis temperature of the crude lithium phosphate can be 80~100°C, specifically 80°C, 90°C or 100°C, and the synthesis time can be 1~3h, specifically 1h, 1.5h, 2h or 3h.

[0048] The invention sequentially performs high-acid leaching and filter pressing on the low-acid leaching residue to obtain high-acid leaching residue and high-acid leaching liquid.

[0049] As an embodiment of the present invention, the initial concentration of the sulfuric acid solution for high acid leaching is 4-7 mol / L, specifically 4 mol / L, 5 mol / L, 6 mol / L and 7 mol / L; the liquid-to-solid ratio of high acid leaching can be 3-5 mL / g, specifically 3 mL / g, 4 mL / g or 5 mL / g; the temperature of high acid leaching can be 60-90°C, specifically 60°C, 70°C, 80°C or 90°C, and the time of high acid leaching can be 3-6 h, specifically 3 h, 4 h, 5 h or 6 h.

[0050] After obtaining the high-acid leaching solution, the present invention adds manganese dioxide to the high-acid leaching solution for oxidation, and adds the crude lithium phosphate to the obtained oxidation solution to remove iron and aluminum and filter press to obtain a solution after removing iron and aluminum.

[0051] As an embodiment of the present invention, the reaction occurring during the oxidation process is: 2Fe 2+ +MnO2+4H+ = 2Fe 3+ +Mn 2+ +2H2O.

[0052] As an embodiment of the present invention, the reaction occurring during the iron and aluminum removal process is: Fe2(SO4)3+2Li3PO4=2FePO4↓+3Li2SO4; Al2(SO4)3+2Li3PO4=2AlPO4↓+3Li2SO4.

[0053] As an embodiment of the present invention, the ratio of the molar amount of manganese dioxide added to the molar amount of Fe element in the high-acid leaching solution is 1.1~1.2:2, the oxidation time can be 1~2h, and the ratio of the molar amount of the added crude lithium phosphate to the total molar amount of iron ions and aluminum ions in the oxidized liquid is 1~1.2:1; during the iron and aluminum removal, the pH value of the system can be 3~4, and the iron and aluminum removal reaction time can be 1~3h, specifically 1h, 2h or 3h.

[0054] After obtaining the iron- and aluminum-removed liquid, the present invention sequentially extracts copper from the iron- and aluminum-removed liquid to obtain copper raffinate and battery-grade copper sulfate.

[0055] As an embodiment of the present invention, the extraction agent for copper extraction can be a mixture of 5-dodecyl salicylaldehyde oxime and 2-hydroxy-5-nonylacetophenone oxime (LIX 984N); as an embodiment of the present invention, during the copper extraction, the extraction phase ratio can be 1-5:1, and the copper-loaded organic obtained by the copper extraction is back-extracted with a 2-4 mol / L sulfuric acid solution to obtain battery-grade copper sulfate.

[0056] After obtaining the copper raffinate, the present invention extracts impurities from the copper raffinate to obtain a de-impurified solution and a manganese solution; the obtained manganese solution is deeply purified to obtain battery-grade manganese sulfate; the deep purification is sequentially performed by sulfidation to remove weight, remove calcium, and remove fluoride; As an embodiment of the present invention, during the sulfidation weight removal, the ratio of the molar amount of Na2S added to the molar amount of Me (Me is one or more of Fe, Al, Cu, and Co) in the manganese solution is 1 to 1.5:1; As an embodiment of the present invention, during the calcium removal, the ratio of the molar amount of NaF added to the molar amount of Ca in the solution after sulfidation weight removal is 2~6:1; during the fluorine removal, the mass of zirconium hydroxide is 50~100 times the mass of fluoride ions in the solution after calcium removal.

[0057] As an embodiment of the present invention, the extraction agent for the extraction can be di(2-ethylhexyl) phosphate (P204); as an embodiment of the present invention, during the extraction, the extraction ratio can be 1~5:1, and the organic layer loaded with metal impurities obtained by the extraction is back-extracted with 2~4 mol / L sulfuric acid solution.

[0058] After obtaining the impurity-removed liquid, the present invention extracts cobalt from the impurity-removed liquid to obtain battery-grade cobalt sulfate and a cobalt raffinate containing low-concentration lithium, nickel and magnesium; the obtained cobalt raffinate containing low-concentration lithium, nickel and magnesium is diverted into a lithium, nickel and magnesium extraction section to extract lithium, nickel and magnesium, and the obtained lithium, nickel and magnesium-loaded organic phase enters the cobalt extraction section again for extraction and separation to obtain a raffinate containing high-concentration lithium, nickel and magnesium.

[0059] As an embodiment of the present invention, the extractant for extracting cobalt can be 2-ethylhexyl mono-2-ethylhexyl phosphate (P507); as an embodiment of the present invention, during the cobalt extraction, the extraction phase ratio can be 1-5:1, and the cobalt-loaded organic obtained by the cobalt extraction is back-extracted with a 1-3 mol / L sulfuric acid solution to obtain battery-grade cobalt sulfate.

[0060] As one embodiment of the present invention, the cobalt extraction and the diversion extraction of lithium, nickel and magnesium are carried out in the cobalt extraction section and the lithium, nickel and magnesium extraction section respectively; as one embodiment of the present invention, a multi-stage lithium, nickel and magnesium extraction section can be added before the cobalt extraction section, and the extractant of the lithium, nickel and magnesium extraction section is P507; the number of stages of the lithium, nickel and magnesium extraction section is 2 to 6. As one embodiment of the present invention, during the extraction process, the diversion portion of the raffinate containing low concentration of lithium, nickel and magnesium in the cobalt extraction section enters the lithium, nickel and magnesium extraction section for extraction of lithium, nickel and magnesium, and the obtained organic phase loaded with lithium, nickel and magnesium is returned to the cobalt extraction section for extraction and separation to obtain a raffinate containing high concentration of lithium, nickel and magnesium.

[0061] After obtaining the raffinate containing high concentrations of lithium, nickel and magnesium, the present invention extracts magnesium from the raffinate containing high concentrations of lithium, nickel and magnesium to obtain magnesium raffinate and battery-grade magnesium sulfate.

[0062] As an embodiment of the present invention, the extractant for magnesium extraction can be di(2,4,4-trimethylpentyl)phosphinic acid (C272); the extraction phase ratio of the magnesium extraction can be 1~5:1, and the magnesium-loaded organic obtained by the magnesium extraction is back-extracted with 2~4 mol / L sulfuric acid solution to obtain battery-grade magnesium sulfate.

[0063] After obtaining the magnesium raffinate, the present invention extracts nickel from the magnesium raffinate to obtain lithium-rich solution and battery-grade nickel sulfate; the lithium-rich solution is reused for step (3) of removing fluorine and weight.

[0064] As an embodiment of the present invention, the extractant for nickel extraction can be P507; during the nickel extraction, the extraction phase ratio can be 1~5:1, specifically 1:1, 2:1, 3:1, 4:1 or 5:1, and the nickel-loaded organic phase obtained by the nickel extraction is back-extracted with 2~4 mol / L sulfuric acid solution to obtain battery-grade nickel sulfate.

[0065] The present invention constructs a lithium closed-loop technology system of "blank roasting-circulating leaching-synergistic impurity removal-lithium phosphate aluminum removal-directional extraction and diversion extraction of lithium nickel and magnesium-seed crystal induction and parallel flow sedimentation of lithium". The whole process does not require the addition of roasting aids, reducing agents and aluminum removal agents, and eliminates redundant processes of traditional lithium carbonate purification such as carbonization, resin purification, and pyrolysis. It also saves the cost of lithium liquid concentration, achieves green and clean production without hazardous waste, and ensures high product quality through synergistic impurity removal and directional extraction. The recovery rates of cobalt, nickel, copper, manganese, magnesium, and lithium in the whole process are all over 99%, and the cobalt, nickel, copper, manganese, magnesium, and lithium products all meet battery-grade standards. The preparation process of battery-grade lithium carbonate is shortened by more than 50% compared with the traditional process, and the processing cost per ton of lithium is reduced by more than 40%. The process is short, the process is simple, the efficiency is high, the energy consumption is low, and the product quality is excellent, providing an efficient solution for the green and high-value recovery of waste lithium batteries.

[0066] In order to further illustrate the present invention, the scheme of the present invention is described in detail below with reference to the accompanying drawings and embodiments, but they should not be understood as limiting the scope of protection of the present invention.

[0067] Example 1 The waste ternary lithium batteries were crushed and sieved to obtain waste ternary lithium battery black powder, the main element content of which is shown in Table 1.

[0068] Table 1 Content of main elements in black powder of waste ternary lithium batteries (unit: %)

[0069] Reaction process: (1) The black powder of waste ternary lithium batteries was introduced into natural gas for blank roasting at a temperature of 550 °C and a roasting time of 2 h to obtain roasted slag.

[0070] (2) The roasted slag was circulated and leached with a dilute sulfuric acid solution at a pH of 2 to enrich lithium. When the lithium concentration of the lithium-rich circulating leachate was ≥20 g / L, part of the leachate was diverted to the defluorination and deweighting process. The undiverted lithium-rich circulating leachate was acidified to a pH of 2 and then returned to the circulation leaching of new roasted slag. The liquid-to-solid ratio of each low-acid leaching was 10 mL / g, and the low-acid leaching time was 1 h. The leachate obtained from each low-acid leaching was acidified to a pH of 2 and then returned to the circulation leaching of new roasted slag. The low-acid circulating leaching process obtained a diverted lithium-rich leachate and a low-acid leaching slag. The concentrations of some elements in the diverted low-acid leachate and the contents of some elements in the low-acid leaching slag are shown in Tables 2 and 3.

[0071] Table 2 Element concentrations of split lithium-rich leachate (g / L)

[0072] Table 3 Content of some elements in low acid leaching residue (%)

[0073] (3) According to n(CaO):[n(Me)+n(F - ) / 2]=1.2:1 (Me: Co, Ni, Mn, Cu elements), calcium oxide was added to the split lithium-rich leachate to remove fluorine and weight. The reaction time was 2 h and the reaction temperature was 30 ° C to obtain a defluorination and weight removal liquid. Table 4 shows the concentrations of some elements in the defluorination and weight removal liquid.

[0074] Table 4 Concentration of some elements in defluorination and heavy liquid (g / L)

[0075] (4) The defluorination and deheaving liquid was subjected to P204 calcium extraction treatment with an extraction ratio of 3:1 to obtain a decalcified liquid. Table 5 shows the concentrations of some elements in the decalcified liquid.

[0076] Table 5 Concentration of some elements in decalcification solution (g / L)

[0077] (5) Zirconium hydroxide in an amount 30 times the total mass of silicon, phosphorus and fluorine in the solution was added to the decalcification solution for deep removal of silicon, phosphorus and fluorine. The treatment time was 1 hour to obtain a silicon, phosphorus and fluorine-free solution. Table 6 shows the concentrations of some elements in the silicon, phosphorus and fluorine-free solution.

[0078] Table 6 Concentration of some elements in silicon, phosphorus and fluorine removal liquid (g / L)

[0079] (6) The silicon, phosphorus, and fluorine removal solution and the sodium carbonate solution were added to the synthesis reactor in parallel. During the synthesis process, battery-grade lithium carbonate was added as a seed crystal to induce lithium precipitation. Sodium carbonate was added according to the ratio of n(Na2CO3):n(Li) = 1.2:2. The concentration of the sodium carbonate solution was 300 g / L. The amount of seed crystal added was 10 wt.% of the mass of the synthesized lithium carbonate. The synthesis temperature was 100°C. After synthesis, the lithium carbonate and lithium precipitation mother liquor were obtained by centrifugation. The lithium carbonate was washed with hot water three times at 80°C for 1 hour to obtain battery-grade lithium carbonate. The concentration of some elements in the lithium precipitation mother liquor and the purity of the battery-grade lithium carbonate were analyzed. The results are shown in Tables 7 and 8.

[0080] Table 7 Concentration of some elements in lithium precipitation mother liquor (g / L)

[0081] Table 8 Content of battery-grade lithium carbonate (%)

[0082] (7) Trisodium phosphate was added to the lithium precipitation mother liquor to synthesize crude lithium phosphate. Trisodium phosphate was added according to the ratio of n(Na3PO4):n(Li)=1.1:3. The synthesis temperature was 80℃ and the synthesis time was 1.5h. Table 9 shows the content of some elements in the crude lithium phosphate.

[0083] Table 9 Content of some elements in crude lithium phosphate (%)

[0084] (8) The low-acid leaching residue was added to a 6 mol / L sulfuric acid solution for high-acid leaching. The high-acid leaching liquid-to-solid ratio was 5 mL / g, the leaching temperature was 80°C, and the leaching time was 5 h. High-acid leaching residue and high-acid leaching solution were obtained. The high-acid leaching residue was subjected to four countercurrent washes to obtain washed high-acid leaching residue. The concentrations of some elements in the high-acid leaching solution and the contents of some elements in the washed high-acid leaching residue were analyzed. The results are shown in Tables 10 and 11.

[0085] Table 10 Concentrations of some elements in high acid leaching solution (g / L)

[0086] Table 11 Content of some elements in high acid leaching residue after washing (%)

[0087] (9) Add manganese dioxide to the high acid leaching solution according to n(MnO2):n(Fe)=1.1:2 for oxidation, and the oxidation time is 1h. 3+ )+n(Al 3+ )] = 1.05:1, crude lithium phosphate was added to the oxidized liquid to remove iron and aluminum. The pH of the deoxidized liquid was controlled at 3.5, and the reaction time was 2 hours. This produced a deoxidized liquid and deoxidized aluminum slag. The deoxidized aluminum slag was then subjected to four countercurrent washes to obtain washed deoxidized aluminum slag. The concentrations of certain elements in the deoxidized liquid and the contents of certain elements in the washed deoxidized aluminum slag were analyzed, and the results are shown in Tables 12 and 13.

[0088] Table 12 Element concentrations in the liquid after iron and aluminum removal (g / L)

[0089] Table 13 Content of some elements in iron and aluminum slag after washing (%)

[0090] (10) After the removal of iron and aluminum, the liquid was subjected to LIX984N copper extraction, P204 impurity extraction and deep purification of manganese liquid, P507 cobalt extraction and split-stream extraction of lithium, nickel and magnesium, C272 magnesium extraction, and P507 nickel extraction in sequence. The extraction ratios were all 3:1. The number of stages of the lithium, nickel and magnesium extraction section of the P507 cobalt extraction process was 4. The back-extraction acids of LIX984N copper extraction, P204 impurity extraction, P507 cobalt extraction, C272 magnesium extraction, and P507 nickel extraction were 3 mol / L sulfuric acid, 3 mol / L sulfuric acid, 2 mol / L sulfuric acid, 2 mol / L sulfuric acid, and 3 mol / L sulfuric acid, respectively, to obtain battery-grade copper sulfate, manganese liquid, battery-grade cobalt sulfate, battery-grade magnesium sulfate, and battery-grade nickel sulfate, respectively. The manganese solution undergoes extensive impurity removal through sulfidation to remove weight (the molar ratio of added Na2S to the molar amount of Me in the solution is 1.5:1, where Me represents Fe, Al, Cu, and Co), decalcification (the molar ratio of added NaF to the molar amount of Ca in the solution after sulfidation to remove weight is 4:1), and defluorination (the mass of zirconium hydroxide is 50 times the mass of fluorine in the solution after decalcification) to produce battery-grade manganese sulfate. The raffinate from the lithium, nickel, and magnesium extraction section of the P507 cobalt extraction process, which contains high concentrations of lithium, nickel, and magnesium, is extracted with C272 for magnesium and then with P507 for nickel to produce a lithium-rich solution. The lithium-rich solution is then returned to the defluorination and deweighting process for closed-loop lithium recovery. The purity of battery-grade products such as copper sulfate, manganese sulfate, cobalt sulfate, magnesium sulfate, and nickel sulfate, as well as the content of certain elements in the lithium-rich solution, are analyzed. The results are shown in Tables 14 and 15.

[0091] Table 14 Purity of battery-grade products such as copper sulfate, manganese sulfate, cobalt sulfate, magnesium sulfate, and nickel sulfate (g / L)

[0092] Table 15 Concentrations of some elements in lithium-rich solution (g / L)

[0093] The recoveries of cobalt, nickel, copper, manganese, magnesium and lithium in Example 1 are shown in Table 16.

[0094] Table 16 Recovery rates of cobalt, nickel, copper, manganese, magnesium and lithium (%)

[0095] Example 2 The waste ternary lithium batteries were crushed and sieved to obtain waste ternary lithium battery black powder, the contents of its main elements are shown in Table 17.

[0096] Table 17 Content of main elements in black powder of used ternary lithium batteries (unit: %)

[0097] (1) The black powder of waste ternary lithium batteries was roasted in a blank process by introducing natural gas at a temperature of 600 °C for 1.5 h to obtain roasted slag.

[0098] (2) The roasted slag was circulated and leached with a dilute sulfuric acid solution at a pH of 2.5 to enrich lithium. When the lithium concentration of the lithium-rich circulating leachate was ≥20 g / L, part of the solution was diverted to the defluorination and deweighting process. The undiverted lithium-rich leachate was acidified to a pH of 2.5 and then returned to the circulation leaching of new roasted slag. The liquid-to-solid ratio of each low-acid leaching was controlled to 15 mL / g, and the low-acid leaching time was 1 h. The leachate obtained from each low-acid leaching was acidified to a pH of 2.5 and then returned to the circulation leaching of new roasted slag. The low-acid circulating leaching process obtained a diverted lithium-rich leachate and a low-acid leachate slag. The concentration of some elements in the diverted lithium-rich leachate and the content of some elements in the low-acid leaching slag were analyzed, and the results are shown in Tables 18 and 19.

[0099] Table 18 Element concentrations of split lithium-rich leachate (g / L)

[0100] Table 19 Content of some elements in low acid leaching residue (%)

[0101] (3) According to n(CaO):[n(Me)+n(F - ) / 2] = 1.15:1 (Me: Co, Ni, Mn, Cu elements). Calcium oxide was added to the split lithium-rich leachate to remove fluorine and heavy metals. The reaction time was 1.5 hours, resulting in a defluorinated and heavy metals-free solution. The concentrations of certain elements in the defluorinated and heavy metals-free solution were analyzed, and the results are shown in Table 20.

[0102] Table 20 Concentration of some elements in defluorination and heavy liquid (g / L)

[0103] (4) The defluorination and deheaving liquid was treated with P204 for calcium extraction at an extraction ratio of 4:1 to obtain a decalcified liquid. The concentrations of some elements in the decalcified liquid were analyzed, and the results are shown in Table 21.

[0104] Table 21 Concentration of some elements in decalcification solution (g / L)

[0105] (5) Zirconium hydroxide (40 times the total mass of silicon, phosphorus, and fluorine in the solution) was added to the decalcification solution for deep silicon, phosphorus, and fluorine removal. The treatment time was 1 hour, resulting in a silicon, phosphorus, and fluorine-free solution. The concentrations of some elements in the silicon, phosphorus, and fluorine-free solution were analyzed, and the results are shown in Table 22.

[0106] Table 22 Concentration of some elements in silicon, phosphorus and fluorine removal liquid (g / L)

[0107] (6) The silicon, phosphorus, and fluorine removal solution and sodium carbonate solution were added to the synthesis reactor in parallel. During the synthesis process, battery-grade lithium carbonate was added as a seed crystal to induce lithium precipitation. Sodium carbonate was added at a ratio of n(Na2CO3):n(Li) = 1.15:2. The concentration of the sodium carbonate solution was 330 g / L. The amount of seed crystal added was 15% of the mass of the synthesized lithium carbonate. The synthesis temperature was 100°C. After synthesis, the lithium carbonate and lithium precipitation mother liquor were obtained by centrifugation. The lithium carbonate was washed with hot water three times at 80°C for 1 hour to obtain battery-grade lithium carbonate. The concentration of some elements in the lithium precipitation mother liquor and the purity of the battery-grade lithium carbonate were analyzed. The results are shown in Tables 23 and 24.

[0108] Table 23 Concentration of some elements in lithium precipitation mother liquor (g / L)

[0109] Table 24 Battery-grade lithium carbonate content (%)

[0110] (7) Trisodium phosphate was added to the lithium precipitation mother liquor to synthesize crude lithium phosphate. Trisodium phosphate was added at a ratio of n(Na3PO4):n(Li) = 1.15:3. The synthesis temperature was 80°C and the synthesis time was 2 h. The crude lithium phosphate was analyzed for the content of some elements. The results are shown in Table 25.

[0111] Table 25 Content of some elements in crude lithium phosphate (%)

[0112] (8) The low-acid leaching residue was added to a 7 mol / L sulfuric acid solution for high-acid leaching. The leaching liquid-to-solid ratio was 5:1, the leaching temperature was 80°C, and the leaching time was 6 h. This produced high-acid leaching residue and high-acid leaching liquid. The high-acid leaching residue was then subjected to four countercurrent washes to produce washed high-acid leaching residue. The concentrations of some elements in the high-acid leaching liquid and the contents of some elements in the washed high-acid leaching residue were analyzed. The results are shown in Tables 26 and 27.

[0113] Table 26 Concentrations of some elements in high acid leaching solution (g / L)

[0114] Table 27 Content of some elements in high acid leaching residue after washing (%)

[0115] (9) Manganese dioxide was added to the high-acid leachate at a ratio of n(MnO2):n(Fe) = 1.1:2 for oxidation, and the oxidation time was 1 h. Crude lithium phosphate was added to the oxidized solution at a ratio of n(Li3PO4):[n(Fe)+n(Al)] = 1.05:1 for de-ironing and de-aluminization. The pH of the de-ironing solution was controlled at 3.7, and the reaction time was 2 h. De-ironing solution and de-ironing slag were obtained. The de-ironing slag was washed with countercurrent four times to obtain washed de-ironing slag. The concentrations of some elements in the de-ironing solution and the contents of some elements in the washed de-ironing slag were analyzed, and the results are shown in Tables 28 and 29.

[0116] Table 28 Element concentrations in the liquid after iron and aluminum removal (g / L)

[0117] Table 29 Content of some elements in iron and aluminum slag after washing (%)

[0118] (10) After the removal of iron and aluminum, the liquid was subjected to LIX984N copper extraction, P204 impurity extraction and deep purification of manganese liquid, P507 cobalt extraction and split extraction of lithium, nickel and magnesium, C272 magnesium extraction, and P507 nickel extraction in sequence. The extraction ratios were all 4:1. The split stage number of the split lithium enrichment section of the P507 cobalt extraction process was 5. The back extraction acids of LIX984N copper extraction, P204 impurity extraction, P507 cobalt extraction, C272 magnesium extraction, and P507 nickel extraction were 3 mol / L sulfuric acid, 3 mol / L sulfuric acid, 2 mol / L sulfuric acid, 2 mol / L sulfuric acid, and 3 mol / L sulfuric acid, respectively, to obtain battery-grade copper sulfate, manganese liquid, battery-grade cobalt sulfate, battery-grade magnesium sulfate, and battery-grade nickel sulfate, respectively. The manganese solution is deeply purified through sulfidation and degravity removal (the ratio of the added molar amount of Na2S to the molar amount of Me in the manganese solution is 1.5:1, where Me is Fe, Al, Cu, or Co), decalcification (the ratio of the added molar amount of NaF to the molar amount of Ca in the solution after sulfidation and degravity removal is 4:1), and defluoridation (the mass of zirconium hydroxide is 50 times the mass of fluorine in the solution after decalcification) to produce battery-grade manganese sulfate. The raffinate from the P507 cobalt extraction process, which contains high concentrations of lithium, nickel, and magnesium, is extracted through C272 for magnesium and P507 for nickel extraction to produce a lithium-rich solution. The lithium-rich solution is then returned to the defluoridation and degravity removal process for closed-loop lithium recovery. The purity of battery-grade products such as copper sulfate, manganese sulfate, cobalt sulfate, magnesium sulfate, and nickel sulfate, as well as the content of certain elements in the lithium-rich solution, are analyzed. The results are shown in Tables 30 and 31.

[0119] Table 30 Purity of battery-grade products such as copper sulfate, manganese sulfate, cobalt sulfate, magnesium sulfate, and nickel sulfate (g / L)

[0120] Table 31 Concentrations of some elements in lithium-rich solution (g / L)

[0121] The recovery rates of cobalt, nickel, copper, manganese, magnesium and lithium in Example 2 are shown in Table 32.

[0122] Table 32 Recovery rates of cobalt, nickel, copper, manganese, magnesium and lithium (%)

[0123] Although the above embodiment provides a detailed description of the present invention, it is only a part of the embodiments of the present invention, not all of the embodiments. Other embodiments can be obtained based on this embodiment without creativity, and these embodiments all fall within the scope of protection of the present invention.

Claims

1. A method for recovering battery-grade lithium carbonate and valuable metals from waste ternary lithium battery black powder in a short process, characterized in that: The following steps are involved: (1) Calcination of waste ternary lithium battery black powder in a natural gas atmosphere to obtain calcined slag; (2) subjecting the roasted slag to low-acid circulating leaching and filter pressing using a dilute sulfuric acid solution to obtain a lithium-rich circulating leachate and low-acid leaching slag; (3) When the lithium concentration in the lithium-rich circulating leachate is ≥20 g / L, a portion of the lithium-rich circulating leachate is diverted and mixed with calcium oxide, and then subjected to defluorination and deweighting and filter pressing in sequence to obtain a defluorination and deweighting liquid; (4) extracting and removing calcium from the defluorination and deheavy-weight removal liquid to obtain a decalcified liquid; (5) adding zirconium hydroxide to the decalcification liquid to deeply remove silicon, phosphorus and fluorine to obtain a silicon, phosphorus and fluorine-removing liquid; (6) Using pure water as the base liquid, adding battery-grade lithium carbonate as a seed crystal, adding the silicon, phosphorus and fluorine removal liquid and sodium carbonate solution into the synthesis kettle in a parallel flow form to induce lithium precipitation, thereby obtaining lithium carbonate and lithium precipitation mother liquor; (7) adding trisodium phosphate to the lithium precipitation mother liquor for synthesis to obtain crude lithium phosphate; (8) subjecting the low-acid leaching residue to high-acid leaching and filter pressing in sequence to obtain high-acid leaching residue and high-acid leaching liquid; (9) adding manganese dioxide to the high acid leaching solution for oxidation, adding the obtained oxidized solution to the crude lithium phosphate product for iron and aluminum removal and filter pressing to obtain an iron and aluminum removal solution and an iron and aluminum removal slag; (10) extracting copper from the iron and aluminum-removed liquid to obtain copper raffinate and battery-grade copper sulfate; (11) extracting the copper raffinate to obtain a de-impurity solution and a manganese solution; The obtained manganese liquid is subjected to deep purification in sequence to obtain battery-grade manganese sulfate; the deep purification is carried out in sequence by sulfidation to remove weight, remove calcium and remove fluoride; (12) extracting cobalt from the impurity-removed liquid to obtain battery-grade cobalt sulfate and a cobalt raffinate containing low-concentration lithium, nickel, and magnesium; extracting lithium, nickel, and magnesium from the cobalt raffinate containing low-concentration lithium, nickel, and magnesium, and re-entering the cobalt extraction section for extraction and separation to obtain a raffinate containing high-concentration lithium, nickel, and magnesium; (13) extracting magnesium from the raffinate containing high concentrations of lithium, nickel and magnesium to obtain magnesium raffinate and battery-grade magnesium sulfate; (14) The magnesium raffinate is subjected to nickel extraction to obtain lithium-rich solution and battery-grade nickel sulfate; the lithium-rich solution is reused for fluorine removal and weight removal in step (3).

2. The method according to claim 1, wherein The blank is calcined at a temperature of 450-650° C. and for a time of 1-3 hours.

3. The method according to claim 1, wherein During the low-acid circulation leaching, the leachate obtained from each low-acid leaching is acidified and then reused for low-acid circulation leaching, and the pH of the dilute sulfuric acid solution or the leachate after acid adjustment is 1-3; during each low-acid leaching, the liquid-to-solid ratio is 10-20 ml / g, and the time of each low-acid leaching is 1-4 hours.

4. The method according to claim 1, wherein The time for defluorination and deweighting is 1 to 4 hours; the temperature is 25 to 50° C.; the ratio of the added molar amount of calcium oxide to the molar amount of fluorine in the partial lithium-rich circulating leachate is 1.1 to 1.3:1; the molar amount of fluorine is the sum of 1 / 2 of the molar amount of fluoride ions and the molar amount of heavy metal ions; the heavy metals include one or more of Co, Ni, Mn and Cu.

5. The method according to claim 1, wherein The extractant for extracting and removing calcium is di(2-ethylhexyl) phosphate; the amount of zirconium hydroxide added is 20 to 60 times the total mass of silicon, phosphorus and fluorine in the decalcification solution; and the time for deep removal of silicon, phosphorus and fluorine is 1 to 3 hours.

6. The method according to claim 1, wherein In step (6), the molar ratio of sodium carbonate in the sodium carbonate solution to the Li element in the silicon, phosphorus and fluorine removal solution is 1.1-1.2:2, the concentration of the sodium carbonate solution is 250-350 g / L; the amount of battery-grade lithium carbonate added is 5-15 wt.% of the mass of the synthesized lithium carbonate, and the temperature for inducing lithium precipitation is 80-100°C.

7. The method according to claim 1, wherein The ratio of the added molar amount of the trisodium phosphate to the molar amount of the Li element in the lithium precipitation mother liquor is 1.1-1.2:3, the synthesis temperature of the crude lithium phosphate is 80-100° C., and the synthesis time is 1-3 hours.

8. The method according to claim 1, wherein The initial concentration of the sulfuric acid solution for high acid leaching is 4-7 mol / L; the liquid-to-solid ratio of high acid leaching is 3-5 mL / g; the temperature of high acid leaching is 60-90° C., and the time of high acid leaching is 3-6 hours.

9. The method according to claim 1, wherein The ratio of the molar amount of manganese dioxide added to the molar amount of Fe element in the high-acid leaching solution is 1.1-1.2:2, the oxidation time is 1-2 hours, and the ratio of the molar amount of the added crude lithium phosphate to the total molar amount of iron ions and aluminum ions in the oxidized solution is 1-1.2:1; during the iron and aluminum removal, the pH value of the system is 3-4, and the iron and aluminum removal reaction time is 1-3 hours.

10. The method according to claim 1, wherein The extractant for copper extraction is a mixture of 5-dodecyl salicylaldehyde oxime and 2-hydroxy-5-nonylacetophenone oxime; the extractant for impurity extraction is di(2-ethylhexyl) phosphate; the extractant for cobalt extraction is 2-ethylhexyl mono-2-ethylhexyl phosphate; the extractant for magnesium extraction is di(2,4,4-trimethylpentyl)phosphinic acid; and the extractant for nickel extraction is 2-ethylhexyl mono-2-ethylhexyl phosphate.

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