Method for treating aluminum electrolysis solid waste

Abstract

The present application relates to a kind of aluminium electrolysis solid waste processing method, comprising the following steps: providing aluminium electrolysis solid waste powder;Wherein, the aluminium electrolysis solid waste powder includes overhaul slag and waste slag, the overhaul slag includes one or several of waste refractory material and waste cathode, the waste slag includes one or two of waste carbon residue and waste electrolyte, and the mass ratio of overhaul slag and waste slag is 6-8:1-4;The aluminium electrolysis solid waste powder is mixed with water at 1:3-5 solid-liquid mass ratio, after reaction at 80-95 DEG C, solid-liquid separation, obtains leaching residue and leaching solution;After acid leaching to the leaching residue, solid-liquid separation, obtains acid leaching residue and lithium-rich acid leaching solution.The processing method of the present application can simultaneously realize to aluminium electrolysis waste refractory material, waste slag and other various aluminium electrolysis solid waste.The processing method of the present application can realize the full quantization recovery of carbon, lithium, cryolite and other substances in aluminium electrolysis solid waste, and the economic benefit is good.

Description

Technical Field

[0001] This invention relates to a method for treating solid waste from aluminum electrolysis, and more particularly to a method for the resource-based treatment of solid waste from aluminum electrolysis, belonging to the technical field of harmless and resource-based treatment of solid waste and hazardous waste. Background Technology

[0002] Currently, the Hall-Eruth molten salt electrolysis process is the main technology for aluminum smelting. During the electrolytic aluminum production process, the penetration and corrosion of the molten high-temperature electrolyte cause deformation and cracking of the electrolytic cell lining. Molten aluminum and electrolyte leak from the cracks to the bottom of the electrolytic cell, rendering it unusable. Therefore, on average, major overhauls of aluminum electrolytic cells are required every 5-8 years, generating large amounts of waste electrolyte, waste cathode, waste carbon slag, and waste refractory materials. Statistics show that for every ton of electrolytic aluminum produced, 10-15 kg of waste cathode, 6-16 kg of waste carbon slag, approximately 14 kg of waste electrolyte, and 10-15 kg of waste refractory materials are generated. Therefore, it is necessary to further treat these aluminum electrolysis solid wastes.

[0003] The "heart" of new energy vehicles is the lithium battery, and the core raw material for manufacturing lithium batteries is lithium carbonate. With the rapid development of new energy vehicles, the demand for lithium carbonate has increased dramatically, growing annually. Aluminum electrolysis solid waste contains 0.43-2% metallic lithium, with a grade close to that of lepidolite ore. Furthermore, the waste cathode contains over 55% graphitized carbon, the waste slag contains over 22% carbon, the waste electrolyte contains over 60% cryolite, and the waste refractory materials also contain 0.45-1.5% lithium. Therefore, aluminum electrolysis solid waste has enormous comprehensive recycling value.

[0004] Against this backdrop, on the basis of harmless treatment, further green and high-value utilization of aluminum electrolysis solid waste can be carried out to extract valuable elements such as lithium carbonate, graphite and cryolite, which is a win-win move that turns waste into treasure and benefits the country and the people.

[0005] Waste carbon slag and waste cathodes mainly consist of carbon and some cryolite, possessing certain recycling value. Currently, the primary technology for industrial-scale resource recovery of carbon slag / waste cathodes is the hydrometallurgical process. The hydrometallurgical process utilizes the differences in the physical and chemical properties of different components in the carbon slag, employing crushing, ball milling, and flotation to recover and recycle fluorine-containing electrolytes and carbon components from the carbon slag / waste cathodes. While the hydrometallurgical process essentially achieves primary resource utilization of carbon slag, the quality of the byproducts, carbon powder and electrolytes, does not meet the requirements for industrial-scale production and requires further processing. Many researchers have also conducted extensive studies on pyrometallurgical recovery processes for carbon slag. Pyrometallurgical processes generally employ chemical reactors, integrating oxidation combustion, catalytic conversion, gasification and impurity removal, conditioning and homogenization, and flue gas treatment. Heating methods include electricity or gas, primarily utilizing high-temperature combustion of carbon in the carbon slag / waste cathodes, with the product being recycled cryolite. The technical disadvantages of the pyrometallurgical process lie in the lifespan of the combustion chamber and the handling of fluorine-containing gases. The process also has high energy consumption and cannot recover carbon from the raw materials, resulting in significant waste. Its advantage is that the recycled cryolite produced has relatively high purity and quality. In China, this process is used for electrolyte purification in the wet flotation of carbon slag / waste cathodes, but its industrial application in the treatment of carbon slag / waste cathodes from electrolytic aluminum plants is extremely rare.

[0006] With the long-term operation of aluminum electrolytic cells, impurities and additives in the raw materials accumulate in the electrolyte, altering its properties and leading to consequences such as low cell temperature, poor alumina dissolution, and excessive sediment at the bottom of the cell. To ensure the normal and stable operation of the electrolytic cells, the current practice is to periodically extract a portion of the high-impurity electrolyte from the producing electrolytic cells and replace it with new electrolyte to ensure the total amount of impurities remains within the process requirements. The replaced electrolyte contains a high percentage of valuable components and can generally be used for the initial start-up of electrolytic cells in newly built aluminum plants after simple treatment. However, with the domestic aluminum production capacity stabilizing in recent years, relying on the initial start-up cells of newly built aluminum plants to utilize this portion of electrolyte is no longer feasible. The replaced waste electrolyte cannot be directly returned to the electrolytic cells for reuse and can only be stockpiled in large quantities, resulting in resource waste and new environmental hazards. Currently, research has developed resource-based treatment methods for waste electrolytes using wet or pyrometallurgical processes, or a combination of wet and pyrometallurgical methods, but these methods have limitations in raw materials and high reagent consumption. Summary of the Invention

[0007] In view of the shortcomings of the existing technology, the purpose of this invention is to provide a method for treating aluminum electrolysis solid waste with lower reagent consumption.

[0008] To solve the above-mentioned technical problems, the technical solution of the present invention is as follows:

[0009] A method for treating solid waste from aluminum electrolysis includes the following steps:

[0010] S1. Provide aluminum electrolysis solid waste powder;

[0011] The aluminum electrolysis solid waste powder includes overhaul slag and waste slag. The overhaul slag includes one or more of waste refractory materials and waste cathodes. The waste slag includes one or two of waste carbon slag and waste aluminum electrolyte. The mass ratio of overhaul slag to waste slag is 6-8:1-4.

[0012] S2. The aluminum electrolyte solid waste powder is mixed with water at a solid-liquid mass ratio of 1:3-5, reacted at 80-95℃, and then the solid and liquid are separated to obtain leaching residue and leaching solution.

[0013] S3. After acid leaching the leaching residue, solid-liquid separation is performed to obtain acid leaching residue and lithium-rich acid leaching solution.

[0014] Therefore, by mixing overhaul slag and waste slag in a certain proportion, aluminum and fluorine in the aluminum electrolysis solid waste can be introduced into the solution during the S2 water leaching process, and lithium-containing substances such as lithium fluoride embedded in the aluminum electrolysis solid waste can be exposed and dissociated. Subsequently, through simple acid leaching, lithium can be dissolved and enriched in the acid leaching solution. The acid leaching solution can be used to prepare lithium salt products through conventional impurity removal and lithium precipitation, realizing resource utilization. The whole process is simple and efficient, has stronger applicability to raw materials, and does not require the addition of alkali to dissolve aluminum and fluorine in aluminum electrolysis solid waste, thus saving reagents and reducing costs. At the same time, it also realizes the recovery and enrichment of lithium in materials with low lithium content, such as overhaul slag.

[0015] Furthermore, in S1, the particle size of the aluminum electrolysis solid waste powder is 200-300 mesh.

[0016] Furthermore, in S1, the aluminum electrolysis solid waste powder includes waste refractory materials, waste cathodes, and waste slag, with a mass ratio of waste refractory materials, waste cathodes, and waste slag of 5:1-3:1-4.

[0017] Furthermore, the waste refractory material contains 5-14 wt% CaF2 and 0.5-10% NaF; furthermore, the waste refractory material contains 0.75-10% C, 0.45-1.5% Li, 30-50 wt% Al2O3, and 25-45 wt% SiO2.

[0018] Furthermore, the waste cathode contains 10-20% NaF; the waste cathode contains 40-60% C, 0.45-1.2% Li, and 0.5-5% cryolite.

[0019] Furthermore, the waste carbon residue contains 15-40% C, 1-1.5% Li, and 20-45% cryolite, and generally, the NaF content is close to 0.

[0020] Furthermore, in the waste aluminum electrolyte, the C content is 0.1-2%, the cryolite content is 40-75%, the Li content is 1-2%, and generally, the NaF content is close to 0. Furthermore, in S2, the mixing time is 0.5-1 h, and the reaction time is 2-4 h.

[0021] Optionally, in S2, during solid-liquid separation, a plate and frame filter press or a belt filter is used. Preferably, a belt filter with higher separation efficiency and the ability to wash wet materials online is used.

[0022] Optionally, in S2, the leaching residue is washed with water, and the amount of washing water is 0.5-1.5 times the dry weight of the leaching residue.

[0023] Further, in S3, acid leaching is carried out at 80-95℃; preferably, during acid leaching, the initial liquid-solid mass ratio is controlled to be 3-5:1, more preferably 3.5-4.5:1; preferably, the acid leaching time is 2-4h, more preferably 2.5-3.5h.

[0024] Optionally, in S3, during solid-liquid separation, a plate and frame filter press or a belt filter is used. Preferably, a filter press with a pressing function is used, and a plate and frame filter press with a lower moisture content in the filter residue is used. This can further reduce the lithium element entrained in the leaching residue and improve the lithium recovery rate.

[0025] Furthermore, the acid used in the acid leaching is hydrochloric acid and / or sulfuric acid, more preferably concentrated sulfuric acid with a concentration of 93-98 wt%, and the amount of concentrated sulfuric acid added is 20-50 wt% of the dry weight of the leaching residue.

[0026] Further, after S3, the acid leaching residue is slurryed to form a slurry with a liquid-to-solid mass ratio of 3-4:1; foaming agent, inhibitor and collector are added to the slurry, and flotation is performed to obtain the flotation top stream; then the flotation top stream is subjected to solid-liquid separation to obtain filtrate and carbon;

[0027] The foaming agent is added at a rate of 0.9-1.1 wt% of the dry weight of the acid leaching residue; the inhibitor is added at a rate of 0.04-0.06 wt% of the dry weight of the acid leaching residue; and the collector is added at a rate of 0.8-1.2 wt% of the dry weight of the acid leaching residue. The collector is composed of gasoline and kerosene in a mass ratio of 65-85:15-35.

[0028] Chemical leaching disrupts the embedded structure of aluminum electrolysis solid waste, allowing carbon in the raw material to be released or dissociated from the structure. Given the low density of carbon, it is expected to be effectively separated by flotation.

[0029] Through in-depth research, the applicant discovered that by using gasoline and kerosene in the above proportion to form a compound collector, and by controlling the addition of foaming agent, inhibitor, and collector, good flotation separation of carbon in acid leaching residue can be achieved, while obtaining high carbon recovery rate and carbon grade.

[0030] Optionally, when performing solid-liquid separation on the flotation top stream, a plate and frame filter press or a belt filter can be used. Preferably, a belt filter with higher separation efficiency and the ability to wash wet material online should be used.

[0031] Furthermore, the foaming agent is pine oil; the inhibitor is water glass, and the modulus of the water glass is 2.4-2.6.

[0032] Further, after S3, after removing impurities from the lithium-rich acid leaching solution, water-soluble carbonate is added, lithium is precipitated, solid-liquid separation is performed, and the solution is dried to obtain the lithium carbonate product.

[0033] Optionally, when removing impurities from the lithium-rich acid leaching solution, calcium hydroxide or calcium oxide is added to the acid leaching solution to convert fluoride ions, aluminum ions, and iron ions into precipitates, thereby achieving solid-liquid separation. Then, an appropriate amount of sodium carbonate and sodium hydroxide is added to remove calcium and magnesium ions from the filtrate. Subsequently, ion exchange is performed for deep impurity removal, where the chelating groups on the ion exchange resin selectively adsorb calcium and magnesium ions in the solution, further reducing the calcium and magnesium ion content in the filtrate.

[0034] Optionally, after impurity removal and before lithium precipitation, the acid leaching solution is evaporated and concentrated. Optionally, the concentration of lithium sulfate, sodium sulfate, and potassium sulfate in the concentrated solution obtained after evaporation and concentration is 350-380 g / L.

[0035] Optionally, the water-soluble carbonate is sodium carbonate. Optionally, the lithium precipitation reaction temperature is 90-95℃; optionally, the lithium precipitation reaction time is 4-6 hours.

[0036] Optionally, the water content of the lithium carbonate product is ≤0.2wt%.

[0037] Optionally, after lithium precipitation, the remaining liquid obtained from solid-liquid separation is neutralized to remove excess carbonate ions, and then evaporated and crystallized to recover sulfate (such as sodium sulfate).

[0038] Further, after step S2, acid is added to the leachate at 40-60℃, and the reaction continues until the pH of the leachate is constant at 8.8-9.2. Solid-liquid separation is then performed to obtain cryolite product. Thus, cryolite can be recovered simply by adjusting the acid content, further improving processing efficiency. Subsequently, the remaining leachate can be treated to remove cyanide and impurities, followed by evaporation and crystallization to recover sulfates, chlorides, and other substances.

[0039] Compared with the prior art, the beneficial effects of the present invention are as follows:

[0040] (1) The processing method of the present invention can simultaneously achieve the synergistic treatment of various aluminum electrolysis solid wastes such as aluminum electrolysis waste refractory materials, waste cathodes, and waste slag, and helps to save reagent usage, making it more green and environmentally friendly.

[0041] (2) The treatment method of the present invention can realize the full-quantity recovery of carbon, lithium, cryolite and other substances in aluminum electrolysis solid waste, with good economic benefits.

[0042] (3) The present invention has a high recovery rate of valuable elements, high purity of the obtained product, and high selling value. Detailed Implementation

[0043] 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 described in the embodiments of the present invention can be combined with each other. Unless otherwise specified, the relevant percentages refer to mass percentages.

[0044] Example 1

[0045] In this embodiment, the method for processing aluminum electrolysis solid waste to produce battery-grade lithium carbonate and recovering carbon and cryolite includes the following steps:

[0046] (1) Raw material crushing and mixing

[0047] Waste refractory materials, waste cathodes, waste carbon slag, and waste electrolytes from aluminum electrolysis are crushed and ground to obtain various powders of 250 mesh.

[0048] A screw weighing scale was used to measure the materials, and a twin-shaft mixer was used to mix the above four types of ground raw materials. The mass ratio of waste refractory materials, waste cathodes, waste carbon slag, and waste electrolyte was 5:2:2:1 to obtain aluminum electrolysis solid waste powder.

[0049] The waste refractory material contains 10.17% C, 1.37% Li, 4.5% NaF, 40.3% Al2O3, 38.6% SiO2, and 8.14% CaF2.

[0050] The waste cathode contains 55% C, 1.13% Li, 3.72% CaF2, and 19.2% NaF.

[0051] The waste carbon residue contains 20% C, 1.2% Li, and 44.85% cryolite.

[0052] The waste aluminum electrolyte contains 0.2% C, 1.47% Li, and 71.56% cryolite.

[0053] (2) Water immersion, acid immersion

[0054] The aluminum electrolysis solid waste powder was mixed with water at a liquid-solid mass ratio of 4:1, slurried at 85°C for 1 hour, and then leached for 3 hours. After filtration through a belt filter, leaching residue and leaching solution were obtained. The leaching residue was washed with industrial water at a volume equal to the residue volume and then set aside for later use.

[0055] The leaching residue was leached using 98% concentrated sulfuric acid. The reaction temperature was controlled at 80℃, the initial liquid-to-solid mass ratio was 3:1, the slurrying time was 1 hour, the leaching reaction time was 3 hours, and the amount of concentrated sulfuric acid added was 40% of the dry weight of the leaching residue. After the reaction, the residue was filtered through a plate and frame filter press to obtain filter residue and acid leaching solution. The filter residue was washed with industrial water of one volume to obtain acid leaching residue for later use.

[0056] (3) Flotation

[0057] The acid leaching residue was pulped with a liquid-to-solid mass ratio of 3:1 for 0.5 hours. The foaming agent was pine oil, added at 1% of the acid leaching residue (dry basis). The inhibitor was water glass, added at 0.05% of the acid leaching residue (dry basis), with a water glass modulus of 2.5. A compound collector of gasoline and diesel oil was used, with gasoline content of 70% and kerosene content of 30%, added at 1% of the acid leaching residue (dry basis).

[0058] (4) Purification and concentration of acid leaching solution

[0059] After primary impurity removal of the acid leaching solution using calcium hydroxide slurry, solid-liquid separation is performed. Specifically, the temperature is controlled at 60℃, the liquid-to-solid mass ratio of the calcium hydroxide slurry is 3:1, the reaction time is 3 hours, the pH at the reaction endpoint is controlled at 8, and the amount of washing water for the primary impurity removal residue is twice the amount of residue (dry basis). The washing liquid is then returned to the acid leaching process.

[0060] Subsequently, the acid leaching solution is subjected to secondary impurity removal using a mixed slurry of sodium carbonate and sodium hydroxide, followed by solid-liquid separation. Specifically, the temperature is controlled at 85℃, the mass ratio of sodium carbonate to sodium hydroxide in the mixed slurry is 5:1, the liquid-solid mass ratio of solvent to solute in the mixed slurry is 3:1, the reaction time is 3 hours, the pH at the reaction endpoint is controlled at 12, the washing water volume of the secondary impurity removal residue is twice the residue volume (dry basis), and the washing liquid should be returned to the acid leaching process.

[0061] Then, the acid leaching solution is thoroughly purified using ion exchange resin (resin model CH-93, product of Lanxiao Company) to ensure that the total concentration of calcium and magnesium ions in the acid leaching solution is ≤5ppm.

[0062] The acid leaching solution was evaporated and concentrated using an MVR complete set of equipment, so that after evaporation and concentration, the lithium ion concentration in the concentrated solution was 15 g / L, and the sum of the concentrations of lithium sulfate, sodium sulfate and potassium sulfate was 365 g / L.

[0063] (5) Lithium precipitation

[0064] Sodium carbonate powder was measured using a screw weigher. Sodium carbonate and water were mixed at a mass ratio of 3:1 to obtain a slurry. The slurry was heated to 75°C.

[0065] Lithium precipitation is performed using a reverse addition method, with the reaction temperature controlled at 95℃ and the reaction time controlled at 4 hours. The feed rate of the concentrated liquid is controlled at 4 m / s. 3 After the reaction, solid and liquid separation was performed to obtain crude lithium carbonate and lithium precipitation mother liquor.

[0066] Crude lithium carbonate is subjected to slurry washing, in which pure water is used for slurry washing, the liquid-to-solid ratio of the slurry should be controlled at 3:1, and the washing should be carried out twice.

[0067] The washed crude lithium carbonate is slurryed to a liquid-to-solid ratio of 25:1, and then pumped to a carbonization tower for carbonization pyrolysis.

[0068] During carbonization and pyrolysis, the CO2 gas supply should be controlled at 0.2t / t lithium carbonate, the CO2 gas flow rate should be controlled at 4L / min, the pressure inside the carbonization tower should be controlled at 0.15MPa, the carbonization temperature should be controlled at 25℃, the carbonization slurry solid ratio should be controlled at 25:1, and the pyrolysis temperature should be controlled at 95℃. The pyrolysis mother liquor should be recycled, with 1 / 5 of it opened each time and returned to prepare sodium carbonate slurry.

[0069] Lithium carbonate is dried in an electric kiln at a temperature of 500℃, and the moisture content of the dried lithium carbonate product is controlled to be ≤0.2%.

[0070] (6) Recovery of lithium precipitation mother liquor

[0071] The mother liquor from lithium precipitation was decarbonized by adding sulfuric acid dropwise, and the pH at the end of the reaction was controlled to be 7 to obtain the decarbonized mother liquor.

[0072] Then, the decarbonization mother liquor was evaporated and crystallized using an MVR complete set of equipment. 40% of the solution volume was evaporated, and the evaporation concentration ratio was 3.25 times to obtain sodium sulfate. After drying, anhydrous sodium sulfate product that can be sold externally can be obtained. (7) Cryolite recovery

[0073] Sulfuric acid was added to the leachate to induce cryolite precipitation, with the temperature controlled at 50℃, the final pH controlled at 9, and the reaction time at 2.5 h. After the reaction, the solid and liquid were separated, washed, and dried to obtain the cryolite product.

[0074] Testing revealed that the obtained lithium carbonate product met the requirements of the "Battery Grade Lithium Carbonate" (YS / T582—2023) standard Li2CO3-D1 for battery grade lithium carbonate, with a comprehensive lithium recovery rate (calculated based on the lithium obtained from the lithium carbonate product) of 76.47%; a total cryolite recovery rate of 75%; and a total carbon recovery rate of 85% (with a grade of 90%), resulting in significant economic benefits.

[0075] Comparative Example 1

[0076] Repeat Example 1, except that in step (1), the mass ratio of waste refractory material, waste cathode, waste carbon slag and waste electrolyte is 5:2:2:3.

[0077] Testing revealed a lithium recovery rate of 58% and a cryolite recovery rate of 53%. The likely reasons are that the relative content of waste refractory materials and waste cathodes decreased, and the lithium-containing substances in waste carbon slag and waste electrolyte were not thoroughly dissociated and exposed. Some lithium and cryolite entered the leaching residue and could not be recovered, leading to a lower overall lithium and cryolite recovery rate.

[0078] Example 2

[0079] Repeat Example 1, except that in step (1), the mass ratio of waste refractory material, waste cathode, waste carbon slag and waste electrolyte is 5:2:2:2.

[0080] Tests showed that the overall lithium recovery rate was 68% and the total cryolite recovery rate was 61%.

[0081] Comparative Example 2

[0082] Repeat Example 1, except that the collector is gasoline.

[0083] Tests showed that the char recovery rate was 75% and the char grade was 95%.

[0084] Comparative Example 3

[0085] Example 1 was repeated, except that the collector was kerosene.

[0086] Tests showed that the char recovery rate was 90% and the char grade was 66%.

[0087] Comparative Example 4

[0088] Example 1 was repeated, except that the compound collector contained 60% gasoline and 40% kerosene.

[0089] The test results showed that the char recovery rate was 88% and the char grade was 80%.

[0090] Example 3

[0091] Example 1 was repeated, except that the gasoline content in the compound collector was 65% and the kerosene content was 35%.

[0092] Tests showed that the char recovery rate was 86% and the char grade was 86%.

[0093] Example 4

[0094] Example 1 was repeated, except that the compound collector contained 80% gasoline and 20% kerosene.

[0095] Tests showed that the char recovery rate was 80% and the char grade was 92%.

[0096] Example 5

[0097] Example 1 was repeated, except that the gasoline content in the compound collector was 85% and the kerosene content was 15%.

[0098] The test results showed that the char recovery rate was 78% and the char grade was 93%.

[0099] Comparative Example 5

[0100] Example 1 was repeated, except that the compound collector contained 90% gasoline and 10% kerosene.

[0101] The test results showed that the char recovery rate was 76% and the char grade was 94%.

[0102] Example 6

[0103] Example 1 was repeated, except that the amount of the compound collector added was 0.8% of the amount of the acid leaching residue (dry basis).

[0104] The carbon recovery rate was found to be 85%.

[0105] Example 7

[0106] Example 1 was repeated, except that the amount of the compound collector added was 0.9% of the amount of the acid leaching residue (dry basis).

[0107] The carbon recovery rate was found to be 88%.

[0108] Example 8

[0109] Example 1 was repeated, except that the amount of the compound collector added was 1.1% of the amount of the acid leaching residue (dry basis).

[0110] Tests showed that the carbon recovery rate was 90%.

[0111] Example 9

[0112] Example 1 was repeated, except that the amount of the compound collector added was 1.2% of the amount of the acid leaching residue (dry basis).

[0113] Tests showed that the carbon recovery rate was 90%.

[0114] The above embodiments should be understood as being used only to illustrate the present invention more clearly, and not to limit the scope of the present invention. After reading the present invention, any modifications of the present invention in various equivalent forms by those skilled in the art fall within the scope defined by the appended claims.

Claims

1. A method for treating aluminum electrolytic solid waste, characterized by, The method comprises the following steps: S1, providing aluminum electrolysis solid waste powder; The aluminum electrolysis solid waste powder comprises overhaul slag and waste slag, the overhaul slag is composed of waste refractory material and waste cathode, and the waste slag is composed of waste carbon slag and waste electrolyte, and the mass ratio of the overhaul slag to the waste slag is 6-8:1-4; S2, uniformly mixing the aluminum electrolysis solid waste powder with water at a solid-liquid mass ratio of 1:3-5, and then performing reaction at 80-95°C, and then performing solid-liquid separation to obtain leaching residue and leaching solution; S3, performing acid leaching on the leaching residue, and then performing solid-liquid separation to obtain acid leaching residue and lithium-rich acid leaching solution; After S2, adding acid to the leaching solution at 40-60°C, and then performing reaction until the pH value of the leaching solution is constant at 8.8-9.2, and then performing solid-liquid separation to obtain cryolite product; After S3, slurry is formed from the acid leaching residue at a liquid-solid mass ratio of 3-4:1, a foaming agent, an inhibitor and a collector are added to the slurry, and then flotation is performed to obtain flotation top flow, and then solid-liquid separation is performed on the flotation top flow to obtain filtrate and carbon; The addition amount of the foaming agent is 0.9-1.1wt% of the dry base weight of the acid leaching residue, the addition amount of the inhibitor is 0.04-0.06wt% of the dry base weight of the acid leaching residue, and the addition amount of the collector is 0.8-1.2wt% of the dry base weight of the acid leaching residue, and the collector is composed of gasoline and kerosene at a mass ratio of 65-85:15-35.

2. The method of treatment of aluminum electrolysis solid waste according to claim 1, characterized in that, In S1, the particle size of the aluminum electrolysis solid waste powder is 200-300 meshes.

3. The method of treatment of aluminum electrolysis solid waste according to claim 1, characterized in that, In S1, the aluminum electrolysis solid waste powder comprises waste refractory material, waste cathode and waste slag, and the mass ratio of the waste refractory material, the waste cathode and the waste slag is 5:1-3:1-4.

4. The method of treatment of aluminum electrolysis solid waste according to claim 1, characterized in that, In S2, the mixing time is 0.5-1h, and the reaction time is 2-4h.

5. The method of treatment of aluminum electrolysis solid waste according to any one of claims 1-3, characterized in that, In S3, acid leaching is performed at 80-95°C.

6. The method of treatment of aluminum electrolysis solid waste according to claim 5, characterized in that, During acid leaching, the initial liquid-solid mass ratio is controlled to be 3-5:1, and the acid leaching time is 2-4h.

7. The method of treatment of aluminum electrolysis solid waste according to claim 5, characterized in that, The acid used in acid leaching is hydrochloric acid and / or sulfuric acid.

8. The method of treatment of aluminum electrolysis solid waste according to claim 7, characterized in that, The acid used in acid leaching is concentrated sulfuric acid with a concentration of 93-98wt%, and the addition amount of the concentrated sulfuric acid is 20-50wt% of the dry base weight of the leaching residue.

9. The method of treatment of aluminum electrolysis solid waste according to claim 1, characterized in that, The foaming agent is pine oil, and the inhibitor is water glass with a modulus of 2.4-2.

6.

10. The method of treatment of aluminum electrolysis solid waste according to any one of claims 1-3, characterized in that, After S3, impurities are removed from the lithium-rich acid leaching solution, a water-soluble carbonate is added, lithium is precipitated, solid-liquid separation is performed, and drying is performed to obtain lithium carbonate product.