A method for extracting lithium from lithium-containing aluminum electrolysis waste and preparing lithium dihydrogen phosphate

Abstract

This invention belongs to the field of metallurgical solid waste resource utilization and relates to a method for extracting lithium from lithium-containing aluminum electrolytic waste and preparing lithium dihydrogen phosphate. The lithium extraction from lithium-containing aluminum electrolytic waste includes the following steps: Step 1: Drying the lithium-containing aluminum electrolytic waste and then ball-milling it with a reactant; Step 2: Adding a solution to the calcined material, adjusting the pH to 1.5-2.5, and maintaining the temperature at 60-95℃ for 0.5-10 hours, followed by solid-liquid separation; Step 3: Adding hydrogen peroxide to filtrate A to obtain oxidized filtrate A; then, simultaneously and slowly introducing the oxidized filtrate A and lithium hydroxide into a reaction vessel, controlling the iron ion concentration in the reaction vessel to be below 1.0 g / L, the reaction pH to be 3-4, the reaction temperature to be 70-95℃, and maintaining the temperature for 0.5-2 hours, followed by solid-liquid separation to obtain filtrate B and filter residue B. This invention has the advantage of significantly reducing the amount of aluminum removal slag and reducing the overall lithium loss during the aluminum removal process.

Description

Technical Field

[0001] This invention belongs to the field of metallurgical solid waste resource utilization and relates to a method for extracting lithium from lithium-containing aluminum electrolysis waste and preparing lithium dihydrogen phosphate. Background Technology

[0002] Currently, the technology for resource utilization of lithium-containing aluminum electrolytes and overhaul residues is not yet mature. Existing technologies involve concentrated acid leaching or roasting-leaching of lithium-containing aluminum electrolytes and overhaul residues. While these methods dissolve as much lithium as possible from the raw materials into the leachate, elements such as iron, aluminum, and fluorine in the raw materials also enter the leachate. Most current impurity removal methods involve adding alkaline substances to the leachate to precipitate iron and aluminum elements. However, this does not solve the problem of lithium loss caused by lithium-iron and lithium-aluminum separation in lithium-containing leachates.

[0003] In the existing lithium-iron separation process, alkaline substances are added to the leachate to control the final pH to 3-6. Iron ions in the leachate are saturated and precipitated as Fe(OH)3. Fe(OH)3 exists in colloidal form, which has extremely poor filtration performance, seriously affecting industrial production efficiency. Furthermore, the iron-removed slag after solid-liquid separation has high moisture content and is difficult to wash subsequently, resulting in some lithium elements entering the iron-removed slag.

[0004] In existing lithium-aluminum separation processes, an alkaline substance is added to control the final pH to 7-10. Aluminum ions in the leachate precipitate as Al(OH)3, which has extremely small crystal particles (only 10-50 μm) and is a colloid with strong adsorption capacity. Most of the lithium enters the aluminum removal slag. Furthermore, when the leachate contains fluorine, the neutralization reaction forms insoluble lithium compounds such as LiNa2AlF6 or LiCaAlF6, causing some lithium to enter the aluminum removal slag. This insoluble lithium cannot be recovered through washing.

[0005] Existing technology CN116732348A discloses a method for recovering lithium salts from aluminum electrolytes. The method involves alkaline leaching of electrolyte powder, causing aluminum and fluorine in the aluminum electrolyte to enter the solution in ionic form, while lithium precipitates as LiF, etc. The precipitate is then acid-leached and impurities removed to obtain a lithium-containing solution. This method uses neutralization to remove aluminum and adds alkaline substances to adjust the pH to 4-6, but it does not solve the problem of lithium loss caused by Fe(OH)3 and Al(OH)3 colloids generated during the neutralization process.

[0006] CN116716487A discloses a method for extracting lithium from electrolytic aluminum waste. The method involves reacting electrolyte powder with concentrated sulfuric acid, followed by water leaching. The leachate is then neutralized and purified with hydrogen peroxide and an alkaline solution, evaporated and concentrated, and subjected to lithium precipitation with sodium carbonate and a carbonization reaction to obtain battery-grade lithium carbonate. This method uses direct leaching with concentrated sulfuric acid, which results in a low leaching rate and significant equipment corrosion. Furthermore, the purification process uses hydrogen peroxide to oxidize Fe. 2+ For Fe 3+ The process involves a one-step neutralization method to remove iron and aluminum. Alkali solution is added to adjust the pH to 7-10, while Fe(OH)3 and Al(OH)3 colloids are generated. A significant amount of lithium is carried out by the impurity removal residue, making it difficult to recover the lithium from the residue through washing with water.

[0007] CN116770097A discloses a method for synergistic lithium extraction using aluminum electrolytic cell overhaul residue. The method involves mixing raw materials including aluminum electrolytic cell overhaul residue, spodumene concentrate, a fluorine-fixing agent, and a binder, forming pellets, and then subjecting them to a roasting reaction. After roasting, leaching is performed using a sulfuric acid solution. This method utilizes the fluorine in the overhaul residue to achieve low-temperature activation and transformation of spodumene. However, iron, aluminum, and fluorine elements in the overhaul residue are dissolved by the sulfuric acid solution and enter the leaching solution. Furthermore, this method does not solve the problem of lithium loss caused by Fe(OH)3 and Al(OH)3 colloids generated during subsequent impurity removal.

[0008] CN 115198111 B relates to a method for extracting lithium from lithium-containing waste aluminum electrolyte, comprising the following steps: crushing the lithium-containing waste aluminum electrolyte to be treated to obtain electrolyte powder; mixing the electrolyte powder with a reactant, calcining at 600-1400℃ for 0.5-5 hours, cooling, and grinding to obtain a mixture powder; mixing the mixture powder with water, stirring and reacting, filtering to obtain filter residue and filtrate; using the filtrate for lithium precipitation to obtain lithium salt. This invention does not use acids or alkalis in its entire processing flow, the reactant is inexpensive and readily available, the production process is environmentally friendly, and the operating conditions are favorable. However, this invention requires a large amount of auxiliary materials, resulting in high auxiliary material costs, and also has a low lithium leaching rate and a high reaction temperature. Summary of the Invention

[0009] The purpose of this invention is to provide a method for extracting lithium from lithium-containing aluminum electrolytic waste and preparing lithium dihydrogen phosphate, which improves lithium extraction rate and reduces lithium loss.

[0010] To solve the above-mentioned technical problems, the technical solution of the present invention is as follows:

[0011] A method for extracting lithium from lithium-containing aluminum electrolysis waste includes the following steps:

[0012] Step 1: After drying the lithium-aluminum electrolytic waste, it is ball-milled and mixed with the reactant, and then calcined and ball-milled to obtain the calcined material;

[0013] Step 2: Add solvent to the calcined material, adjust the pH to 1.5-2.5, keep it at 60-95℃ for 0.5-10h, and then perform solid-liquid separation to obtain filtrate A and filter residue A;

[0014] Step 3: Add hydrogen peroxide to filtrate A to obtain oxidized filtrate A; then slowly pass oxidized filtrate A and lithium hydroxide into the reaction vessel simultaneously, controlling the concentration of iron ions in the reaction vessel to be below 1.0 g / L, the pH of the reaction to be 3-4, the reaction temperature to be 70-95℃, and after holding at the temperature for 0.5-2 h, separate the solid and liquid to obtain filtrate B and filter residue B;

[0015] Step 4: Lithium hydroxide is introduced into filtrate B, and the reaction temperature is controlled at 70-95℃, the pH at 7-10, and the introduction time is 2-10h; after the reaction is completed, solid-liquid separation is performed to obtain filtrate C and filter residue C.

[0016] Step 5: Adjust the pH of filtrate C to 12-13, then add trisodium phosphate, and react at 60-95℃ for 0.5-10 hours. Then perform solid-liquid separation to obtain filtrate D and filter residue D.

[0017] Step 6: Heat the filtrate D to 80-90°C, add trisodium phosphate, and continue to keep it warm for 0.5-3 hours. Then perform solid-liquid separation to obtain filtrate E and filter residue E. Filtrate E is the lithium precipitation mother liquor; filter residue E is lithium phosphate.

[0018] Lithium-containing aluminum electrolysis waste includes one or both of lithium-containing waste aluminum electrolyte and overhaul slag.

[0019] In one preferred embodiment, the reactant in step 1 is one or more of Ca(OH)2, CaCO3, CaO, lime, limestone and their hydrates.

[0020] In one preferred embodiment, the molar ratio of lithium to reactant in the lithium-containing aluminum electrolytic waste in step 1 is 1:2-5.

[0021] In one preferred embodiment, the mass ratio of lithium-aluminum electrolytic waste to reactant in step 1 is 1 to 5:1.

[0022] If the mass ratio of lithium-containing aluminum electrolytic waste to reactant is too small, it will affect subsequent leaching. If it is too large, not only will the production cost be too high, but it will also reduce the lithium content of the subsequent roasted material.

[0023] In one preferred embodiment, the calcination temperature in step 1 is 150–450°C.

[0024] If the calcination temperature in step 1 is too low, the reaction will be incomplete, resulting in a lower subsequent leaching rate. If the temperature is too high, sintering will occur, which will also reduce the extent of subsequent reactions and increase costs.

[0025] In one preferred embodiment, the solvent in step 2 is water or an alcohol solvent.

[0026] In one preferred embodiment, the mass ratio of the calcining material to the solution in step 2 is 1:1.5 to 10, preferably 1:1.5 to 3.

[0027] A high solid-liquid mass ratio will lead to a lower leaching rate. A low solid-liquid mass ratio will lead to a lower lithium concentration, which is not conducive to subsequent lithium precipitation.

[0028] In one preferred embodiment, step 2 further includes performing three reverse washings on filter residue A, with the washing liquid returned to step 2.

[0029] The filter residue from step 2 is a mixture of CaF2 and Na3AlF6, which can be returned to the aluminum plant after flotation treatment.

[0030] In one preferred embodiment, the hydrogen peroxide in step 3 is industrial 30% hydrogen peroxide. Hydrogen peroxide is added to filtrate A to remove Fe from filtrate A. 2+ All oxidized to Fe 3+ Simultaneously, cyanide ions are removed from the solution. The oxidized filtrate A is obtained.

[0031] In one preferred embodiment, in step 3, the oxidized filtrate A is reacted at a concentration of 0.5–1.5 mg / L. 3 / h rate (at 3m 3 The solution system (based on the standard) is introduced into the reaction vessel.

[0032] The oxidized filtrate A and lithium hydroxide were simultaneously and slowly introduced into the reactor. The rate at which the oxidized filtrate A was introduced was controlled to keep the iron ion concentration in the reactor below 1.0 g / L. Through multiple experiments, it was found that when the Fe concentration in the solution... 3+ When the concentration is greater than 1 g / L, Fe 3+ Fe(OH)3 precipitate begins to form. The filtrate A after oxidation is introduced at too high a rate, causing Fe... 3+ The rate at which it is added to the reactor is greater than that of Fe. 3+ The rate at which FeOOH is produced in the reaction leads to Fe 3+ As the ion concentration increases, when the Fe in the solution... 3+ When the concentration is greater than 1 g / L, Fe 3+ Fe(OH)3 precipitate begins to form.

[0033] The flow rate of the oxidized filtrate A is too slow, resulting in low production efficiency.

[0034] In one preferred embodiment, step 3 further includes: washing the filter residue B with water once, and then returning the washing liquid to step 3.

[0035] In step 3, the reaction temperature, time, and pH value after introducing lithium hydroxide into the oxidized filtrate A significantly affect the reaction process and efficiency. Specifically: Too low a reaction temperature will produce some Fe(OH)3 with strong adsorption capacity; too high a temperature will result in excessive energy consumption; too short a reaction time will lead to incomplete reaction; too long a reaction time will affect production efficiency; too low a pH value will reduce the iron removal effect and produce some Fe(OH)3 with strong adsorption capacity; too high a pH value will produce some Fe(OH)3 with strong adsorption capacity and also generate Al(OH)3 precipitate.

[0036] In one preferred embodiment, the lithium hydroxide in steps 3 and 4 is homemade.

[0037] In one preferred embodiment, the preparation steps of lithium hydroxide include: mixing water and calcium alkali with lithium carbonate to form a slurry, controlling the liquid-solid ratio to be 8-10:1, wherein the molar ratio of lithium carbonate to calcium alkali is 1.05-1.20, heating to 80-90℃, and holding the reaction at that temperature for 30 min; then performing solid-liquid separation, concentrating the filtrate to 40-50 g / L, and adding 0.05 g / L-1.0 g / L of surfactant to obtain self-made lithium hydroxide.

[0038] In one preferred embodiment, the lithium carbonate is raw lithium carbonate or substandard lithium carbonate.

[0039] In one preferred embodiment, the calcium base is calcium oxide, calcium carbonate, calcium bicarbonate, or calcium hydroxide.

[0040] During the reaction, if the liquid-to-solid ratio is too low, the reaction will be incomplete. If the liquid-to-solid ratio is too high, the concentration will decrease, increasing the cost of subsequent evaporation and concentration. If the holding time is too short, the reaction will be incomplete. If the holding time is too long, production efficiency will decrease.

[0041] In one preferred embodiment, the surfactant is one or more of sodium dodecylbenzenesulfonate, sodium fatty alcohol ether sulfate, sodium α-alkenyl sulfonate, and sodium secondary alkyl sulfonate.

[0042] The surfactant significantly increases the particle size of Al(OH)3, reduces its adsorption capacity, decreases slag volume, and reduces lithium loss during aluminum removal. Furthermore, in the preparation of lithium hydroxide, a suitable reaction temperature and a suitable solid-liquid ratio are beneficial for increasing particle size, and a reduced addition rate of the aluminum remover is also beneficial for increasing particle size. This results in aluminum hydroxide particles precipitated as an aluminum remover prepared by the method of the present invention having a larger particle size than when using commercially available lithium hydroxide.

[0043] In one preferred embodiment, the aluminum hydroxide precipitated using the lithium hydroxide as an aluminum remover has a particle size D50 of 25-35 micrometers.

[0044] Homemade lithium hydroxide requires no impurity removal, only reaction concentration, which is simple and inexpensive.

[0045] In step 4, the reaction temperature, time, and pH value of the lithium hydroxide influent to filtrate B significantly affect the reaction process and efficiency. Specifically: Too low a reaction temperature leads to a decrease in aluminum hydroxide particle size, increased adsorption, and increased slag volume. Higher temperatures result in higher energy consumption. Too low a pH value results in incomplete aluminum removal. Too high a pH value leads to changes in the form of aluminum hydroxide, making subsequent lithium recovery difficult.

[0046] In one preferred embodiment, in step 4, lithium hydroxide is added at a concentration of 0.03 mg / L. 3 ~0.15m 3 / h rate (at 3m 3 Add (based on the solution system).

[0047] The faster the lithium hydroxide passes through, the smaller the aluminum hydroxide particles become, resulting in greater adsorption capacity and increased slag volume, thus increasing the loss of aluminum and lithium. Conversely, a slower rate reduces production efficiency.

[0048] In one preferred embodiment, step 4 further includes washing the filter residue C once with water, and then returning the washing liquid to step 4; the washed aluminum slag is returned to step 1.

[0049] In step 4, the reaction temperature and time for adding trisodium phosphate to filtrate C significantly affect the reaction process and efficiency. Too low a temperature will affect calcium removal efficiency. Too high a temperature will result in excessive energy consumption. Too short a time will lead to incomplete calcium removal. Too long a time will reduce production efficiency.

[0050] In one preferred embodiment, step 6 further includes freezing the filtrate E to precipitate sodium, and then returning it to step 3 as wash water.

[0051] The temperature and time of lithium deposition in step 6 significantly affect the reaction process and efficiency. Too low a temperature will affect deposition efficiency. Too high a temperature will result in excessive energy consumption. Too short a time will lead to incomplete deposition. Too long a time will reduce production efficiency.

[0052] A method for preparing lithium dihydrogen phosphate from the above-mentioned lithium phosphate includes the following steps: adding water to filter residue E and then passing phosphoric acid through it, adjusting the final pH to 2-4, reacting for 0.5-2 hours, and then performing solid-liquid separation to obtain filtrate F; recrystallizing, concentrating, and drying filtrate F to obtain lithium dihydrogen phosphate.

[0053] During the reaction, the pH value needs to be below 4 for lithium phosphate to fully react into lithium dihydrogen phosphate, but too low a pH value will waste raw materials.

[0054] The present invention will be further explained below:

[0055] The principle of this invention is as follows:

[0056] Step 1: First, dry the lithium-aluminum electrolytic waste, then ball mill and mix it with the reactants. Next, feed the mixture into a rotary kiln for low-temperature calcination and roasting, cool, and ball mill to obtain the roasted material. The reaction formula is as follows:

[0057] C + O₂ = CO₂↑

[0058] 4CaCO3+2Na2LiAlF6=4CaF2+4NaF+Al2O3+2Li2O+4CO2↑

[0059] 4CaO+2Na2LiAlF6=4CaF2+4NaF+Al2O3+2Li2O

[0060] 4Ca(OH)2+2Na2LiAlF6=4CaF2+4NaF+Al2O3+2Li2O+4H2O↑

[0061] Step 2: Mix the roasted material with the solution at a liquid-to-solid ratio of 1.5 to 10. Adjust the final pH of the solution to 1.5 to 2.5 with concentrated acid and keep it warm for 0.5 to 10 hours. Separate the solid and liquid to obtain filtrate A (leaching solution) and filter residue A (leaching residue).

[0062] After three reverse washings of filter residue A, the washing liquid is returned to step 2. The filter residue from lithium-aluminum electrolysis is a mixture of CaF2 and Na3AlF6, which can be returned to the aluminum plant after flotation treatment.

[0063] The chemical reaction equations involved in this step are as follows:

[0064] Li₂O + H₂SO₄ = Li₂SO₄ + H₂O

[0065] 2LiF + H₂SO₄ = Li₂SO₄ + 2HF

[0066] 2NaF + H₂SO₄ = Na₂SO₄ + 2HF

[0067] Al₂O₃ + 3H₂SO₄ = Al₂(SO₄)₃ + 3H₂O

[0068] Fe + H₂SO₄ = FeSO₄ + H₂↑

[0069] Step 3: Add sufficient hydrogen peroxide to filtrate A, and remove Fe from filtrate A. 2+All oxidized to Fe 3+ Simultaneously, cyanide ions in the solution are broken down to obtain oxidized filtrate A. The hydrogen peroxide used is industrial-grade 30% hydrogen peroxide.

[0070] The oxidized filtrate A was diluted with 0.5–1.5 mg / L. 3 / h rate (at 3m 3 The solution (based on the standard) is introduced into the reactor, along with self-made lithium hydroxide. The pH is controlled at 3-4, and the reaction temperature is controlled at 70-95℃. After the oxidized filtrate A is completely introduced, it is kept at this temperature for 0.5-2 hours, followed by solid-liquid separation to obtain filtrate B (iron-free liquid) and filter residue B (iron-free slag). The neutralizing agent is the self-made lithium hydroxide solution, and the filter residue B is yellow goethite slag with excellent filtration performance.

[0071] After washing filter residue B once with water, the washing liquid is returned to step 3.

[0072] The chemical reaction equations involved in this step are as follows:

[0073] 2Fe2SO4+H2O2+H2SO4=Fe2(SO4)3+2H2O

[0074] HCN + H₂O₂ = HCNO + H₂O

[0075] 2Fe2(SO4)3+4H2O=4FeOOH↓+3H2SO4

[0076] H₂SO₄ + 2LiOH = Li₂SO₄ + 2H₂O

[0077] Step 4: Add all of the filtrate B into the reaction vessel, control the reaction temperature at 70–95°C, and add 50 g / L lithium hydroxide at a rate of 0.03 mg / L. 3 ~0.15m 3 / h rate (at 3m 3 The solution system is used as a reference. The solution is added to the filtrate B, and the pH of the solution is adjusted to 7-10. The addition time is 2-10 hours. After the reaction is completed, the solid and liquid are separated to obtain filtrate C (aluminum removal solution) and filter residue C (aluminum removal slag). After the filter residue C is washed with water once, the washing liquid is returned to step 4, and the water-washed aluminum slag is returned to step 1.

[0078] The chemical reaction equations involved in this step are as follows:

[0079] Al2(SO4)3+6LiOH→3Na2SO4+2Al(OH)3↓

[0080] Al(OH)3 + 6HF = H3AlF6 + 3H2O

[0081] 2H3AlF6+Li + +2Na+ =Na2LiAlF6↓+3H +

[0082] Step 5: Adjust the final pH of filtrate C to 12-13, then add an appropriate amount of trisodium phosphate, and react at 60-95℃ for 0.5-10 hours. Separate the solid and liquid to obtain filtrate D (purified liquid) and filter residue D (calcium and magnesium slag).

[0083] The chemical reaction equations involved in this step are as follows:

[0084] MgSO4+2NaOH=Mg(OH)2↓+Na2SO4

[0085] 3CaSO4+2Na3PO4=Ca3(PO4)2↓+3Na2SO4

[0086] Step 6: Heat the filtrate D to 80-90°C, add solid trisodium phosphate, and continue to keep it warm for 0.5-3 hours after addition. After solid-liquid separation and filtration, wash the filter cake multiple times to obtain filtrate E (lithium precipitation mother liquor) and filter residue E (lithium phosphate).

[0087] The filtrate E, after being frozen to precipitate sodium, is returned to steps 3 and 4 as wash water.

[0088] The chemical reaction equations involved in this step are as follows:

[0089] 3Li2SO4+2Na3PO4=2Li3PO4+3Na2SO4

[0090] Step 7: Add the filter residue E to the reactor, add water to adjust the liquid-to-solid ratio to 2:1, and stir at room temperature. Pass 85% phosphoric acid into the reactor, adjust the final pH to 2-4, and continue the reaction for 0.5-2 hours. Separate the solid and liquid components to obtain filtrate F (saturated lithium dihydrogen phosphate solution) and filter residue F (insoluble impurities).

[0091] The filtrate F was recrystallized, evaporated and concentrated, and dried to obtain the final battery-grade lithium dihydrogen phosphate product.

[0092] The chemical reaction equations involved in this step are as follows:

[0093] Li3PO4 + 2H3PO4 = 3LiH2PO4

[0094] The beneficial effects of this invention are as follows:

[0095] This invention proposes a method for extracting lithium from aluminum electrolysis waste to prepare lithium dihydrogen phosphate. The method employs a two-stage process to remove iron and aluminum. The first stage uses a goethite method to remove iron. By controlling the feeding rate and reaction temperature, a self-made lithium hydroxide solution (prepared from substandard lithium carbonate and lime slurry from the factory's battery-grade lithium carbonate production line, incurring only processing costs, with its high-value lithium element entering the system) is added as a neutralizing agent to control the reaction pH, resulting in the precipitation of FeOOH slag. Compared to traditional methods using calcium oxide, calcium carbonate, and magnesium carbonate, using a self-made lithium hydroxide solution as a neutralizing agent has the advantages of producing less iron-removing slag and having a simpler phase, facilitating resource recycling. Compared to sodium hydroxide, it has the advantages of lower cost and not introducing other impurities into the solution. The second stage uses a slow neutralization method with an aluminum removal agent. The self-made lithium hydroxide solution is used as the aluminum removal agent. By controlling the addition rate and reaction temperature of the aluminum removal agent, Al(OH)3 precipitate is precipitated, and soluble lithium is extracted from the aluminum-removing slag by water washing. A method for extracting insoluble lithium in the form of LiNa2AlF6 from water-washed aluminum slag was developed. Using self-made lithium hydroxide as a neutralizing agent, this method significantly reduces the amount of aluminum slag compared to traditional calcium oxide, thereby saving costs in subsequent insoluble lithium extraction, increasing lithium recovery from the aluminum slag, and reducing overall lithium loss during the aluminum removal process. This solves the problem of process lithium loss in existing lithium extraction processes and increases the profitability of lithium extraction from lithium-containing aluminum electrolytes and overhaul slag. It avoids the risk of accumulating lithium-containing electrolytes and fluorine-containing waste such as overhaul slag due to the enrichment of lithium and other alkali metals in the aluminum electrolysis industry, and also alleviates the lithium resource shortage caused by the rapid development of electric vehicles and energy storage industries, contributing to the development of the national new energy industry. Detailed Implementation

[0096] The present invention will be described in detail below with reference to embodiments. It should be noted that, unless otherwise specified, the embodiments and features in the embodiments of the present invention can be combined with each other.

[0097] Main raw material sources:

[0098] The lithium-containing aluminum electrolytic cell came from an aluminum plant in Henan Province.

[0099] The main components are shown in Table 1 below:

[0100] Table 1 Main components of lithium-aluminum electrolytic cells

[0101] element Li Na K Fe Al F other content(%) 1.856 23.180 2.367 0.758 14.135 52.98 4.724

[0102] The slag from the overhaul came from an aluminum plant in Hubei.

[0103] The main components are shown in Table 2 below:

[0104] Table 2 Main Components of Major Repair Slag

[0105] element Li Na K Fe Al F C other content(%) 1.519 16.770 0.636 1.394 6.547 18.57 32.89 21.674

[0106] Example 1

[0107] Preparation of lithium hydroxide:

[0108] Substandard lithium carbonate was mixed with water and calcium oxide to form a slurry with a liquid-to-solid ratio of 9:1, wherein the molar ratio of lithium carbonate to calcium hydroxide was 1.1:1. The mixture was heated to 85°C and kept at that temperature for 30 minutes. After solid-liquid separation, the filtrate was concentrated to 50 g / L, and 0.3 g / L sodium dodecylbenzenesulfonate was added to obtain the self-made lithium hydroxide.

[0109] Example 2

[0110] (1) One ton of lithium-aluminum electrolyte and 0.35 tons of calcium carbonate were placed in a ball mill, crushed and mixed, and passed through a 150-mesh sieve to obtain a mixture. The mixture was placed in a rotary kiln and calcined at 350°C for 1.5 hours. After cooling, it was ball-milled and passed through a 150-mesh sieve to obtain 1.14 tons of calcined material.

[0111] (2) The calcined material was mixed with water at a liquid-to-solid ratio of 3:1 and placed in a reaction vessel. The reaction temperature was controlled at 60°C. 98% concentrated sulfuric acid was added to adjust the pH of the solution to 3. The mixture was kept at this temperature for 2 hours for leaching. Solid-liquid separation was then performed to obtain 2.9 m³ of the solution. 3 Filtrate A (leaching solution) and filter residue A (leaching residue); Filter residue A was subjected to three countercurrent washings, with a liquid-to-solid ratio of 3:1, yielding 0.79 tons of leached residue. This residue was then subjected to flotation separation to obtain high-value fluorite and cryolite. The composition of filtrate A is shown in Table 3.

[0112] Table 3 Composition of Filtrate A

[0113] element Li Fe Al Na <![CDATA[SO4 2- ]]> Content (g / L) 6.2 3.7 10.2 52.9 196.6

[0114] The Li content in the three washing residues was 0.08%, and the calculated lithium leaching rate was 96.6%.

[0115] (3) Add 30 L of 30% H2O2 to the filtrate A, react at 60°C for half an hour, and remove Fe. 2+ Oxidized to Fe 3+ At the same time, it breaks down cyanide ions.

[0116] (4) Dissolve the oxidized filtrate A at 1.1m 3 A flow rate of 100 g / L was introduced into the reactor, controlling the reaction temperature at 85°C. Simultaneously, a 50 g / L self-made lithium hydroxide solution was introduced to maintain the pH at 3. After filtrate A was completely introduced, the reaction was continued at this temperature for 0.5 h. A yellow goethite precipitate with good filtration performance but poor adsorption capacity was formed in the solution. Solid-liquid separation was performed to obtain 3.0 mL of filtrate. 3Filtrate B (iron removal solution) and filter residue B (iron removal slag). Filter residue B (iron removal slag) was washed once, with a liquid-to-solid ratio of filter residue B (iron removal slag) to wash water of 2:1, yielding 20 kg of washed iron slag. Tests showed that filtrate B contained 0.02 g / L of Fe and washed iron slag contained 0.04% of Li. The calculated iron removal rate was 99.8%, and the iron and lithium loss rate was 0.03%.

[0117] (5) Add all of the filtrate B into the reaction vessel, control the reaction temperature at 85℃, and add 50g / L of self-made lithium hydroxide solution at 0.13m 3 The solution was added to the reactor at a rate of [value] / h to adjust the final pH to 8, with an addition time of 3 hours. After the addition was complete, the reaction was maintained at this temperature for another 0.5 hours. Solid-liquid separation was then performed to obtain 3.5m [amount of product]. 3 Filtrate C (aluminum removal solution) and filter residue C (aluminum removal slag). Filter residue C was washed once, with a liquid-to-solid ratio of filter residue C to wash water of 2:1, to remove soluble lithium. 62.3 kg of washed aluminum slag was obtained, and the Li content was tested to be 2.78%, of which insoluble lithium accounted for 9.93% of the total lithium content of the raw material. The washed aluminum slag was returned to step 1, and insoluble lithium was extracted by an auxiliary material roasting-leaching method. 74.6 kg of roasted aluminum slag was obtained, with Li content of 0.05%. The CA1 of the filtrate was 0.01 g / L, and the calculated aluminum removal rate was 99.9%, with a comprehensive lithium loss rate of 0.2%.

[0118] (6) Add all of the filtrate C to the reaction vessel, control the reaction temperature at 85℃, add solid sodium hydroxide, adjust the final pH to 13, add 1.1 times the theoretical amount of trisodium phosphate, react for 1 hour, separate the solid and liquid, and obtain 3.5m 3 Filtrate D (purified liquid) and 12 kg filter residue D (calcium and magnesium slag removed).

[0119] (7) Add all of the filtrate C to the reactor, control the reaction temperature at 90°C, add solid trisodium phosphate at 1 times the theoretical amount, and keep the reaction at this temperature for 1.5 hours. After solid-liquid separation, wash the filter cake multiple times to obtain filtrate E (lithium precipitation mother liquor) and 124 kg of filter residue E (lithium phosphate). After freezing to precipitate sodium, filtrate E is returned to step (3) and used as washing water in step (4).

[0120] (8) Add the filter residue E to the reactor, add water to adjust the liquid-to-solid ratio to 2:1, and stir at room temperature. Pass 85% phosphoric acid into the reactor and adjust the final pH to 2.5-3. Separate the solid and liquid to obtain filtrate F (lithium dihydrogen phosphate solution) and filter residue F (insoluble impurities). After drying filtrate F, add water to form a saturated lithium dihydrogen phosphate solution, and then concentrate and crystallize again. After two recrystallization operations, evaporate and concentrate, dry, crush, and package to obtain the final battery-grade lithium dihydrogen phosphate product.

[0121] After calculation, the overall process yields a lithium recovery rate of 96.09% based on the lithium content in the lithium-containing aluminum electrolyte.

[0122] Example 3

[0123] (1) Place 1 ton of overhaul slag and 0.25 tons of calcium oxide in a ball mill, crush and mix them, and pass them through a 200-mesh sieve to obtain a mixture. Place the mixture in a rotary kiln and calcine at 450°C for 2.5 hours. After cooling, ball mill the mixture and pass it through a 200-mesh sieve to obtain 0.96 tons of calcined material.

[0124] (2) The roasted material was mixed with water at a liquid-to-solid ratio of 2:1 and placed in a reaction vessel. The reaction temperature was controlled at 80°C. 98% concentrated sulfuric acid was added to adjust the pH of the solution to 2.5. The mixture was kept at this temperature for leaching for 4 hours. Solid-liquid separation was then performed to obtain 1.9 m³ of the solution. 3 Filtrate A (leaching solution) and filter residue A (leaching residue); filter residue A was subjected to three countercurrent washings, with a liquid-to-solid ratio of filter residue A to washing water of 2:1, yielding 0.67 tons of three-stage washing residue. The composition of filtrate A is shown in Table 4.

[0125] Table 4 Composition of Filtrate A

[0126] element Li Fe Al Na <![CDATA[SO4 2- ]]> Content (g / L) 7.8 10.8 7.5 31.17 119.63

[0127] The Li content in the three washing residues was 0.08%, and the calculated lithium leaching rate was 97.1%.

[0128] (3) Add 60 L of 30% H2O2 to the filtrate A, react at 60°C for half an hour, and remove Fe. 2+ Oxidized to Fe 3+ At the same time, it breaks down cyanide ions.

[0129] (4) Dissolve the oxidized filtrate A at 0.8m 3 A flow rate of 100 g / L was introduced into the reactor, controlling the reaction temperature at 90°C. Simultaneously, a 50 g / L self-made lithium hydroxide solution was introduced to maintain the pH at 3.5. After filtrate A was completely introduced, the reaction was continued at this temperature for 0.5 h. A yellow goethite precipitate with good filtration performance but poor adsorption capacity was formed in the solution. Solid-liquid separation yielded a 2.2 m³ precipitate. 3 Filtrate B (iron removal solution) and filter residue B (iron removal slag). Filter residue B (iron removal slag) was washed once, with a liquid-to-solid ratio of filter residue B (iron removal slag) to wash water of 2:1, yielding 72 kg of washed iron slag. Tests showed that filtrate B contained 0.03 g / L of Fe and washed iron slag contained 0.04% of Li. The calculated iron removal rate was 99.8%, and the iron and lithium loss rate was 0.19%.

[0130] (5) Add all of the filtrate B into the reaction vessel, control the reaction temperature at 80℃, and add 50g / L self-made lithium hydroxide at 0.09m 3The solution was added to the reactor at a rate of [value] / h to adjust the final pH to 8, with an addition time of 2 hours. After the addition was complete, the reaction was maintained at this temperature for another 0.5 hours. Solid-liquid separation was performed to obtain 2.5m [unit of liquid]. 3 Filtrate C (aluminum removal solution) and filter residue C (aluminum removal slag). Filter residue C was washed once, with a liquid-to-solid ratio of filter residue C to wash water of 2:1, to remove soluble lithium. 32 kg of washed aluminum slag was obtained, and the Li content of the washed aluminum slag was tested to be 2.88%, of which insoluble lithium accounted for 6.07% of the total lithium content of the raw material. The washed aluminum slag was returned to step 1 for calcination roasting-leaching to remove insoluble lithium. 38.1 kg of roasted aluminum slag was obtained, with Li content of 0.06%, and the Al content of filtrate C was tested to be 0.01 g / L. The calculated aluminum removal rate was 99.9%, and the overall lithium loss rate after aluminum removal was 0.1%.

[0131] (6) Add all of the filtrate C to the reaction vessel, control the reaction temperature at 70℃, add solid sodium hydroxide, adjust the final pH to 12, add 1.2 times the theoretical amount of trisodium phosphate, react for 1 hour, separate the solid and liquid, and obtain 2.7m 3 Filtrate D (purified liquid) and 11 kg of filter residue D (calcium and magnesium slag removed).

[0132] (7) Add all of the filtrate C to the reactor, control the reaction temperature at 80°C, add 1.05 times the theoretical amount of solid trisodium phosphate, and keep the reaction at this temperature for 2.5 hours. After solid-liquid separation, wash the filter cake multiple times to obtain filtrate E (lithium precipitation mother liquor) and 104 kg of filter residue E (lithium phosphate). After freezing to precipitate sodium, filtrate E is returned to step (3) and used as washing water in step (4).

[0133] (8) Add the filter residue E to the reactor, add water to adjust the liquid-to-solid ratio to 3:1, and stir at room temperature. Pass 85% phosphoric acid into the reactor and adjust the final pH to 3-3.5. Separate the solid and liquid to obtain filtrate F (lithium dihydrogen phosphate solution) and filter residue F (insoluble impurities). After drying filtrate F, add water to form a saturated lithium dihydrogen phosphate solution, and then concentrate and crystallize again. After two recrystallization operations, evaporate and concentrate, dry, crush, and package to obtain the final battery-grade lithium dihydrogen phosphate product.

[0134] After calculation, the overall process yields a lithium recovery rate of 96.68% based on the lithium content in the lithium-containing aluminum electrolyte.

[0135] The performance of battery-grade lithium dihydrogen phosphate products was tested, and the results are shown in Table 5.

[0136] Table 5 Performance test results of battery-grade lithium dihydrogen phosphate products

[0137]

[0138] Comparative Example 1

[0139] This comparative example provides a method for extracting lithium using a lithium-containing aluminum electrolyte, the raw materials and steps of which are basically the same as those in Example 2.

[0140] The difference lies in the fact that the feeding speed is increased from 0.13 m / s to 0.39 m / s in step 5. 3 / h.

[0141] As a result, the amount of aluminum slag increased, with the amount of water-washed aluminum slag increasing from 62.3 kg in Example 1 to 138.6 kg. The water-washed aluminum slag showed a Li content of 2.47%, of which insoluble lithium accounted for 19.62% of the total lithium content. This led to a significant decrease in the lithium concentration in the solution and increased the cost of subsequent calcination-leaching for lithium recovery. Furthermore, after removing insoluble lithium through the calcination-calcination-leaching method, the overall lithium loss rate after aluminum removal increased from 0.2% in Example 1 to 0.45%.

[0142] Comparative Example 2

[0143] This comparative example provides a method for extracting lithium using a lithium-containing aluminum electrolyte, the raw materials and steps of which are basically the same as those in Example 2.

[0144] The difference is that in step 5, the aluminum removal temperature is reduced from 85°C to 60°C.

[0145] As a result, the amount of aluminum slag increased, with the amount of water-washed aluminum slag increasing from 62.3 kg in Example 1 to 238.6 kg. Tests showed that the water-washed aluminum slag contained 2.09% Li, of which insoluble lithium accounted for 28.59% of the total lithium content. This led to a significant decrease in the lithium concentration in the solution and increased the cost of subsequent calcination-leaching for lithium recovery. Furthermore, after removing insoluble lithium through the calcination-calcination-leaching method, the overall lithium loss rate after aluminum removal increased from 0.2% in Example 1 to 0.77%.

[0146] Comparative Example 3

[0147] This comparative example provides a method for extracting lithium from lithium-containing aluminum electrolytes. The raw materials and steps are basically the same as those in Example 2, except that calcium oxide is used as an aluminum removal agent in step 5, and the feeding time is the same as in Example 2.

[0148] Comparative Example 4

[0149] This comparative example provides a method for extracting lithium using a lithium-containing aluminum electrolyte. The raw materials and steps are basically the same as those in Example 2, except that calcium carbonate is used as an aluminum removal agent in step 5, and the feeding time is the same as in Example 2.

[0150] Comparative Example 5

[0151] This comparative example provides a method for extracting lithium from lithium-containing aluminum electrolytes. The raw materials and steps are basically the same as those in Example 2, except that: in step 5, purchased lithium hydroxide is used as an aluminum removal agent, and the feeding time is the same as in Example 2.

[0152] The results of comparative examples 2 and 3, 4 and 5 are shown in Table 6.

[0153] Table 6 Comparison results of Example 2 and Comparative Examples 3, 4, and 5

[0154]

[0155] The results show that using the self-made lithium hydroxide solution as the aluminum removal agent requires less slag than using traditional calcium oxide, calcium carbonate, or purchased lithium hydroxide. This is because using calcium oxide or calcium hydroxide as aluminum removal agents produces precipitates such as calcium sulfate, increasing the slag volume. The aluminum hydroxide particles precipitated using the self-made lithium hydroxide solution are larger than those precipitated using purchased lithium hydroxide. The D50 of the aluminum hydroxide particles precipitated from the self-made lithium hydroxide solution is 28.79 micrometers, while the D50 of the aluminum hydroxide particles precipitated from the purchased lithium hydroxide solution is 17.32 micrometers. Furthermore, insoluble lithium is easier to recover because insoluble lithium exists in the form of LiCaAlF6 when using calcium oxide or calcium hydroxide as aluminum removal agents, while it exists in the form of LiNa2AlF6 when using the self-made lithium hydroxide solution. The lithium element in LiNa2AlF6 is easier to recover using the calcination roasting method compared to LiCaAlF6. Overall, the aluminum removal loss rate and cost are the lowest.

[0156] Comparative Example 6

[0157] This comparative example provides a method for extracting lithium from lithium-containing aluminum electrolytes. The raw materials and steps are basically the same as in Example 2, except that in step 5, the calcination-leaching method is not used to treat the washed aluminum slag; instead, the washed aluminum slag is directly discarded. The tested Li content in the washed aluminum slag was 2.78%. The calculated overall lithium loss rate after aluminum removal is 9.7%.

[0158] Comparative Example 7

[0159] This comparative example provides a method for extracting lithium from aluminum electrolyte containing lithium. The raw materials and steps are basically the same as in Example 2, except that in step 5, instead of using the calcination-leaching method to treat the washed aluminum slag, the washed aluminum slag is added to a reaction vessel, followed by 124.6 L of water and 6.2 kg of calcium oxide, and stirred until homogeneous. The temperature is raised to 85°C and held for one hour. Solid-liquid separation yields 103.2 kg of alkaline-washed aluminum slag, with a Li content of 1.3%. The overall lithium loss rate after aluminum removal is calculated to be 7.5%.

[0160] Comparative Example 8

[0161] This comparative example provides a method for extracting lithium from lithium-containing aluminum electrolytes. The raw materials and steps are basically the same as in Example 2, except that in step 5, instead of using the calcination-leaching method to treat the washed aluminum slag, the washed aluminum slag is fed into a rotary kiln without any auxiliary materials and roasted at 400°C for 2 hours, yielding 53.0 kg of roasted material. This roasted material is mixed with water at a liquid-to-solid ratio of 2:1 and placed in a reaction vessel. The reaction temperature is controlled at 80°C, and 98% concentrated sulfuric acid is added to adjust the pH of the solution to 2.5. The mixture is then kept at this temperature for 4 hours for leaching. After solid-liquid separation, 50.6 kg of roasted slag is obtained. The Li content of the roasted slag is tested to be 2.72%. The calculated overall lithium loss rate after aluminum removal is 7.7%.

[0162] The results of comparative examples 2 and 6, 7 and 8 are shown in Table 7.

[0163] Table 7 Comparison results between Example 2 and Comparative Examples 6, 7, and 8

[0164]

[0165] The results show that the calcination-leaching method can recover most of the insoluble lithium from the washed aluminum slag and return it to the system, with a recovery efficiency far higher than other methods.

[0166] Comparative Example 9

[0167] This comparative example provides a method for extracting lithium from overhaul slag. The raw materials and steps are basically the same as in Example 3, except that in step 4, the rate at which the oxidized filtrate A is added is changed from 0.8 m / s² to 0.8 m / s². 3 / h increased to 2.4m 3 / h.

[0168] As a result, some Fe(OH)3 colloids with poor filtration performance and strong adsorption capacity were generated in the solution. The solution had poor filtration performance, making solid-liquid separation difficult. After solid-liquid separation, iron-removed slag was obtained. After one washing of the iron-removed slag, 173.2 kg of water-washed iron slag was obtained. The Li content of the water-washed iron slag was measured to be 1.05%. The calculated lithium loss rate was 12.04%.

[0169] Comparative Example 10

[0170] This comparative example provides a method for extracting lithium from overhaul slag. The raw materials and steps are basically the same as those in Example 3, except that in step 4, the reaction temperature is reduced from 90°C to 50°C.

[0171] As a result, no goethite was produced in the solution, and iron ions precipitated as Fe(OH)3. The solution had poor filtration performance, making solid-liquid separation difficult. After solid-liquid separation, iron-removed slag was obtained. After one washing of the iron-removed slag, 165.6 kg of water-washed iron slag was obtained. The Li content of the water-washed iron slag was measured to be 1.04%. The calculated lithium loss rate was 11.4%.

[0172] Comparative Example 11

[0173] This comparative example provides a method for extracting lithium from overhaul slag. The raw materials and steps are basically the same as those in Example 3, except that in step 4, the reaction control pH is changed from 3.5 to 5.

[0174] As a result, some Fe(OH)3 colloids with poor filtration performance and strong adsorption capacity were formed in the solution, and some Al(OH)3 precipitates were also formed. The solution had poor filtration performance, making solid-liquid separation difficult. After solid-liquid separation, iron-aluminum slag was obtained. After one washing of the iron-aluminum slag, 196.8 kg of washed iron-aluminum slag was obtained. The Li content of the washed iron-aluminum slag was measured to be 1.54%. The calculated iron-lithium loss rate was 20.1%.

[0175] Comparative Example 12

[0176] This comparative example provides a method for extracting lithium from overhaul slag. The raw materials and steps are basically the same as those in Example 3, except that in step 4, 10% calcium oxide is used as a neutralizing agent.

[0177] Comparative Example 13

[0178] This comparative example provides a method for extracting lithium from overhaul slag. The raw materials and steps are basically the same as those in Example 3, except that in step 4, 30% sodium hydroxide is used as a neutralizing agent.

[0179] Comparative Example 14

[0180] This comparative example provides a method for extracting lithium from overhaul slag. The raw materials and steps are basically the same as those in Example 3, except that magnesium carbonate is used as a neutralizing agent in step 4.

[0181] The results of comparative examples 3 and 12, 13, and 14 are shown in Table 8.

[0182] Table 8 shows the comparison results between Example 3 and Comparative Examples 12, 13, and 14.

[0183]

[0184]

[0185] The results show that under suitable addition rates, reaction temperatures, and process pH, various neutralizing agents can produce goethite slag with good filtration performance, in which lithium is easily washed out. Using a self-made lithium hydroxide solution as a neutralizing agent, the amount of slag to be treated and the iron removal loss rate are less than those of traditional calcium oxide and magnesium carbonate, and slightly lower than those of sodium hydroxide. Furthermore, the self-made lithium hydroxide has the advantages of lower cost and no introduction of impurities into the solution compared to sodium hydroxide.

[0186] Comparative Example 15

[0187] This comparative example provides a method for extracting lithium using a lithium-containing aluminum electrolyte, the raw materials and steps of which are basically the same as those in Example 3.

[0188] The difference lies in the following: In step 4, the self-made lithium hydroxide is added to the oxidized filtrate A directly until the endpoint pH = 8, and the reaction temperature and addition rate are consistent with those in Example 3. After one washing of the iron-aluminum slag, the water-washed iron-aluminum slag is returned to step 1 for calcification roasting-leaching.

[0189] As a result, iron and aluminum ions in the solution simultaneously precipitated as two colloidal precipitates, Fe(OH)3 and Al(OH)3, resulting in poor filtration performance. After solid-liquid separation, iron-aluminum slag was obtained. After one washing of the iron-aluminum slag, 216.2 kg of water-washed iron-aluminum slag was obtained. The Li content of the water-washed iron-aluminum slag was measured to be 1.63%. After calcination roasting and leaching of the water-washed iron-aluminum slag, 247.9 kg of roasted iron-aluminum slag was obtained. The Li content of the roasted iron-aluminum slag was measured to be 0.07%. The overall iron-aluminum-lithium loss rate was calculated to be 1.1%, which is higher than the 0.29% in Example 3.

[0190] Comparative Example 15 used a one-step neutralization method to precipitate iron and aluminum simultaneously in the leachate, resulting in a large amount of slag, difficult filtration, reduced production efficiency, and increased costs for subsequent recovery of insoluble lithium. Furthermore, the overall aluminum loss during iron, aluminum, and lithium removal was higher than in Example 3. Therefore, a two-step method for separate iron and aluminum removal is superior to a one-step neutralization method for iron and aluminum removal.

Claims

1. A method of extracting lithium from lithium-containing aluminum electrolytic scrap, characterized by, The method comprises the following steps: Step 1: dry the lithium-containing aluminum electrolytic waste and mix it with a reagent by ball milling, then calcine and ball mill to obtain a calcined material; Step 2: add a solution to the calcined material, adjust the pH to 1.5-2.5, and then incubate at 60-95℃ for 0.5-10h, and then perform solid-liquid separation to obtain filtrate A and residue A; Step 3: add hydrogen peroxide to the filtrate A to obtain oxidized filtrate A; then slowly pass the oxidized filtrate A and lithium hydroxide into a reaction kettle at the same time, control the concentration of iron ions in the reaction kettle to be less than 1.0g / L, the pH of the reaction is 3-4, the reaction temperature is 70-95℃, and incubate for 0.5-2h, then perform solid-liquid separation to obtain filtrate B and residue B; Step 4: pass lithium hydroxide into the filtrate B, control the reaction temperature to be 70-95℃, the pH to be 7-10, and the passing time to be 2-10h; after the reaction is completed, perform solid-liquid separation to obtain filtrate C and residue C; Step 5: adjust the pH of the filtrate C to 12-13, then add trisodium phosphate, and react at 60-95℃ for 0.5-10h, then perform solid-liquid separation to obtain filtrate D and residue D; Step 6: warm the filtrate D to 80-90℃, then add trisodium phosphate, continue to incubate for 0.5-3h, then perform solid-liquid separation to obtain filtrate E and residue E, and the filtrate E is the lithium precipitation mother liquor; the residue E is lithium phosphate; The reagent in step 1 is one or more of Ca(OH)2, CaCO3, CaO, lime, limestone and hydrates thereof; The solution in step 2 is water or an alcohol solvent; In step 4, lithium hydroxide is added at a rate of 0.03m³-0.15m³ / h; The lithium hydroxide in steps 3 and 4 is self-made; The preparation steps of the self-made lithium hydroxide include: mixing water and calcium base into a slurry in lithium carbonate, controlling the liquid-solid ratio to be 8-10:1, wherein the molar ratio of lithium carbonate to calcium base is 1.05-1.20, warming to 80-90℃, and incubating for 30min; then perform solid-liquid separation, concentrate the filtrate to 40-50g / L, and add 0.05g / L-1.0g / L of a surfactant to obtain self-made lithium hydroxide; The calcium base is calcium oxide, calcium carbonate, calcium bicarbonate or calcium hydroxide; In step 3, based on a system of 3m³ of solution, the oxidized filtrate A is passed into the reaction kettle at a rate of 0.5-1.5m³ / h.

2. The method of claim 1, wherein, The molar ratio of lithium in the lithium-containing aluminum electrolytic waste to the reagent in step 1 is 1:2-5.

3. The method of claim 1, wherein, The solid-liquid mass ratio of the calcined material to the solution in step 2 is 1:1.5-10.

4. The method of claim 3, wherein, The solid-liquid mass ratio of the calcined material to the solution in step 2 is 1:1.5-3.

5. The method of claim 1, wherein, The lithium carbonate is raw ore containing lithium carbonate or unqualified lithium carbonate.

6. The method of claim 1, wherein, The surfactant is one or more of sodium dodecylbenzenesulfonate, sodium fatty alcohol ether sulfate, sodium alpha-alkenyl sulfonate, and sodium secondary alkyl sulfonate.

7. The method of claim 1, wherein, In step 3, the residue B is also washed once, and the washing liquid is returned to step 3.

8. The method of claim 7, wherein, In step 6, the filtrate E is also frozen to precipitate sodium, and then returned to step 3 as washing water.

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