Recycling method for complex electrolyte hazardous waste rich in aluminum oxide

Abstract

The invention belongs to the technical field of aluminum smelting solid hazardous waste recycling, and particularly relates to a recycling method of aluminum oxide-rich complex electrolyte hazardous waste. According to the method, the complex aluminum electrolyte hazardous waste, aluminum salt and silicon-containing oxide are mixed and then subjected to segmented roasting-leaching treatment, fluorine in the electrolyte is converted into aluminum fluoride firstly and then converted into silicon tetrafluoride gas, separation of fluorine and aluminum and separation of gas and solid are achieved, and sodium and lithium elements are converted into soluble salt and further separated and recycled. According to the process, collaborative recovery and high-value utilization of various valuable elements in the complex aluminum electrolyte hazardous waste are achieved. The whole technological process is environment-friendly, toxic and harmful gas emission is avoided, secondary solid waste residues are not generated, the method has the advantages of being simple in process, high in recovery rate, high in product additional value and the like, and a feasible technical scheme is provided for resource utilization of the aluminum electrolysis solid waste.

Description

Technical Field

[0001] This invention belongs to the field of solid hazardous waste recycling technology in aluminum smelting, specifically relating to a resource utilization method for complex electrolyte hazardous waste rich in alumina. Background Technology

[0002] The production of metallic aluminum or aluminum-silicon alloys using lithium-rich alumina or aluminum-silicon oxide involves a series of serious process problems. During electrolysis, impurities from the raw materials accumulate in the electrolyte system as fluorides, leading to a significant increase in the concentration of these impurities. This accumulation significantly alters the physicochemical properties of the electrolyte, disrupting the thermal balance of the electrolytic cell. Changes in the electrolyte's chemical composition affect its conductivity and alumina-dissolving capacity, resulting in decreased stability during electrolysis. More seriously, impurities such as lithium and potassium accumulate in the alumina-rich, complex electrolyte waste—including sediments at the bottom of the electrolytic cell, anode covering materials, and spent electrolytes—creating a vicious cycle that ultimately prevents the reuse of these electrolyte wastes in the electrolysis process. Currently, the main industrial method for treating this complex alumina-rich electrolyte waste is stockpiling. This not only occupies vast amounts of land resources and causes severe environmental pollution but also results in a huge waste of valuable resources such as aluminum, fluorine, sodium, and lithium. Therefore, developing efficient and environmentally friendly technologies for the resource recovery of complex electrolyte waste is of great strategic significance for promoting the green and low-carbon development of my country's aluminum industry.

[0003] Patent CN116768246A uses a segmented roasting method with aluminum sulfate to convert fluorine into aluminum fluoride and sodium, lithium, and potassium into soluble salts. However, the roasting process generates a large amount of sulfur-containing flue gas and is difficult to effectively convert the inert alumina in the electrolyte, resulting in insufficient purity of the aluminum fluoride product and limiting its application range. Patent CN118479499A proposes a calcination roasting-acid-alkali combined leaching process to achieve aluminum-fluorine separation, but the high-temperature acid leaching consumes a lot of energy, and the subsequent alkali neutralization process wastes resources. Patents CN105293536A and CN102079534A use concentrated sulfuric acid to treat fluorine-containing waste residue to produce cryolite, but the strong acid environment will release highly corrosive hydrogen fluoride gas, which will seriously damage the equipment. The calcium hydroxide-hydrochloric acid combined recovery process proposed by patent CN116732348A faces the dilemma of large reagent consumption and large amount of secondary solid waste generation.

[0004] In summary, for the aluminum salt roasting method, the high alumina content (up to 40%) in the furnace bottom sediment and anode covering material makes effective alumina conversion difficult. This results in unreacted alumina being mixed into the aluminum fluoride product, significantly reducing product purity and limiting its application range. While the acid-base combined method can achieve elemental separation, it suffers from high acid and alkali consumption, severe equipment corrosion, and high processing costs. Therefore, developing a novel technology capable of co-extracting and high-value-added utilization of aluminum, fluorine, sodium, and lithium in complex electrolyte hazardous waste with high alumina content has become a critical technological bottleneck that urgently needs to be overcome in this resource utilization field, and is of great significance for promoting the sustainable development of the aluminum industry. Summary of the Invention

[0005] To address the shortcomings of existing technologies, this invention proposes a resource recovery method for complex electrolyte hazardous waste rich in alumina, enabling the recycling of valuable elements in the complex electrolyte hazardous waste while simultaneously producing high-value-added silicon tetrafluoride gas.

[0006] The method of the present invention mainly includes the following:

[0007] The crushed, complex electrolyte waste rich in alumina is mixed evenly with aluminum salts and silicon oxides, and then subjected to a staged roasting-leaching treatment to obtain alumina products (pure alumina or a mixture of alumina and silicon dioxide) and silicon tetrafluoride gas. There are two methods for this staged roasting-leaching treatment: Method 1: one-stage roasting-leaching-two-stage roasting, and Method 2: one-stage roasting-two-stage roasting-leaching.

[0008] Method 1 includes the following: complex electrolyte hazardous waste is mixed evenly with aluminum salt, and after a first stage of roasting, a first stage roasting product is obtained. The first stage roasting product is leached to obtain a first stage filter residue and a first stage filtrate. The first stage filter residue is dried at 60℃-150℃ and then roasted in a second stage to obtain alumina product and silicon tetrafluoride gas.

[0009] Method 2 includes the following: complex electrolyte hazardous waste is mixed evenly with aluminum salt and silicon oxide, and then roasted in a first stage followed by a second stage to obtain the second stage roasting product and silicon tetrafluoride gas. The second stage roasting product is then leached to obtain the second stage filter residue, i.e., alumina product and the second stage filtrate.

[0010] The first-stage roasting temperature is 300℃-600℃, and the first-stage roasting time is 60min-480min. This avoids the pyrolysis of aluminum salts and the generation of corrosive gases under high-temperature (>600℃) conditions, while also reducing energy consumption and damage to equipment. The second-stage roasting temperature is 600℃-1200℃, and the second-stage roasting time is 30min-480min. The alumina product has a purity ≥98.6% and can be used for aluminum electrolysis or aluminum-silicon alloy electrolysis; the silicon tetrafluoride gas has a purity ≥99.80%.

[0011] Complex electrolyte hazardous waste mainly contains alumina and elements such as fluorine, lithium, sodium, potassium, magnesium, and calcium. After a first-stage roasting reaction, the fluorine in the complex electrolyte hazardous waste is converted into aluminum fluoride, and the lithium, sodium, potassium, magnesium, and calcium elements are converted into soluble salts. Alumina does not participate in the reaction, and all these elements form the first-stage roasting product. After a second-stage roasting, all the fluorine is converted into silicon tetrafluoride gas, achieving highly efficient separation between fluorine and aluminum, and between gas and solid. No other impurity gases are generated during the second-stage roasting reaction, greatly reducing the purification cost of silicon tetrafluoride gas.

[0012] Complex electrolyte hazardous waste rich in alumina consists of one or more of the following: furnace bottom sediment, anode covering material, and waste electrolyte from the electrolytic production of aluminum or aluminum alloys; aluminum salts include aluminum chloride, aluminum sulfate, and aluminum nitrate; silicon oxides consist of one or more of the following: mullite, fly ash, coal gangue, and other solid waste impurities, as well as silicon dioxide. The mass ratio of complex electrolyte hazardous waste to aluminum salts is 1:0.25-1:4; the molar ratio of fluorine in complex electrolyte hazardous waste to silicon in silicon oxides is 4:(1-10).

[0013] The aluminum-silicon oxide contains impurities of ≤0.2% iron oxide, ≤0.1% calcium oxide, and ≤0.02% magnesium oxide, with the remainder being aluminum oxide and silicon dioxide. When the molar ratio of fluorine to silicon in the complex electrolyte hazardous waste is 4:1, all silicon and fluorine are converted to silicon tetrafluoride, and all aluminum is converted to aluminum oxide, yielding a pure alumina product with a purity ≥98.6% (SiO2 <0.1%), which can be used for aluminum electrolysis.

[0014] The leaching agent for the leaching treatment is one or more of water, salt solution, and dilute acid solution. The salt solution is one or more of sodium salt solution and ammonium salt solution; the dilute acid solution is one or more of dilute hydrochloric acid, dilute sulfuric acid, and dilute nitric acid; the liquid-solid ratio of the leaching agent to the first-stage or second-stage roasted product is (4-10):1. The leaching temperature is 20℃-95℃, and the leaching time is 10min-480min. After leaching treatment, the soluble salts in the roasted product enter the filtrate (first-stage or second-stage filtrate), while aluminum fluoride and alumina enter the first-stage filter residue (method one) or alumina enters the second-stage filter residue (method two).

[0015] The first-stage or second-stage filtrate can be processed as follows:

[0016] Sodium carbonate or sodium carbonate aqueous solution is added to the first or second stage filtrate and heated and stirred at a temperature of 50℃-100℃. The lithium carbonate product and the delithiation filtrate are obtained by filtration. The purity of the lithium carbonate product is ≥98.0%. The delithiation filtrate is evaporated and crystallized at 50℃-120℃ to obtain the sodium salt product.

[0017] This invention addresses the challenge of resource recovery from complex electrolyte waste with high alumina content by innovatively developing a green and efficient synergistic extraction method. Using complex electrolyte waste and aluminum salts as raw materials, a roasting-leaching-filtration process system achieves selective separation and efficient recovery of elements such as aluminum, fluorine, sodium, and lithium. Key advancements include the development of aluminum salt roasting technology for efficient conversion of complex electrolyte waste; the establishment of a multi-stage separation and purification mechanism to ultimately obtain alumina products and silicon tetrafluoride gas; and the design of a solution-based step-by-step treatment process for the separate preparation of lithium carbonate and sodium salts. Compared with existing technologies, the advantages of this invention are mainly reflected in: environmental friendliness, with no toxic byproduct emissions throughout the process and waste gas and wastewater meeting treatment standards; economic benefits, with inexpensive and readily available raw materials, avoiding high-temperature and high-pressure leaching operations under strong acid or strong alkali conditions; and industrial feasibility, with a simple process route, conventional equipment, and stable and controllable process parameters. Pure alumina can be returned to the aluminum electrolytic cell for the production of metallic aluminum; a mixture of alumina and silicon dioxide can be fed into the electrolytic cell to produce aluminum-silicon alloys; silicon tetrafluoride gas can be used to produce high-purity aluminum fluoride or, after further purification, can be used as a raw material for the production of high-purity silicon; lithium carbonate, after further purification, can be used as a raw material for lithium batteries. This method provides a practical solution for the resource utilization of complex electrolyte hazardous waste, not only realizing the high-value utilization of waste, but also having significant environmental and economic benefits for promoting the green and sustainable development of the electrolytic aluminum industry. Attached Figure Description

[0018] Figure 1 This is a flowchart of Embodiment 1 of the present invention.

[0019] Figure 2 The image shows the XRD pattern of the pure alumina product obtained in Example 1 of this invention.

[0020] Figure 3 The image shows the XRD pattern of the pure alumina product obtained in Example 4 of this invention. Detailed Implementation

[0021] The technical solution of the present invention will be clearly and completely described below with reference to the embodiments and accompanying drawings. It should be noted that the embodiments described in this invention are only for further explanation and illustration, and not for limiting their application scope. Based on this invention, all other embodiments obtained by those skilled in the art without inventive effort are within the protection scope of this invention.

[0022] Example 1

[0023] This embodiment uses method one for processing, and the process flow is as follows: Figure 1 As shown.

[0024] S1: The crushed bottom precipitate of the electrolytic aluminum furnace is mixed evenly with aluminum chloride and subjected to a first-stage roasting reaction to obtain a first-stage roasting product. The mass ratio of bottom precipitate to aluminum chloride is 1:0.25, the roasting temperature is 300℃, and the roasting time is 480min.

[0025] S2: The first stage of roasted product is leached, and filtered to obtain a first stage of residue and a first stage of filtrate. The leaching agent is a sodium chloride solution with a concentration of 1 mol / L-5 mol / L; in this example, 2.5 mol / L is used. The liquid-to-solid ratio of the leaching agent to the first stage of roasted product is 4:1, the leaching temperature is 95℃, and the leaching time is 10 min.

[0026] S3: After drying the filter residue at 60℃, it is mixed evenly with silica (the molar ratio of fluorine in the furnace bottom precipitate to silicon in the silica is 4:1), and a two-stage calcination reaction is carried out at 600℃ for 480 min to obtain pure alumina product and silicon tetrafluoride gas. The XRD pattern of the obtained pure alumina product is shown below. Figure 2 As shown, the purity of the pure alumina product is 99.5%. The purity of the silicon tetrafluoride gas is 99.88%.

[0027] S4: Add sodium carbonate to the first filtrate and heat and stir at 50°C. Filter to obtain lithium carbonate product and delithiation filtrate. The lithium carbonate product has a purity of 99.0%. Evaporate the delithiation filtrate at 50°C to obtain sodium chloride product.

[0028] Example 2

[0029] This embodiment uses method one for processing.

[0030] S1: The crushed electrolytic aluminum anode covering material is mixed evenly with aluminum sulfate and subjected to a first-stage roasting reaction to obtain a first-stage roasting product. The mass ratio of anode covering material to aluminum sulfate is 1:1, the roasting temperature is 450℃, and the roasting time is 240 min.

[0031] S2: The first stage of roasted product is leached, and filtered to obtain a first stage of residue and a first stage of filtrate. The leaching agent is water. The liquid-solid ratio of the leaching agent to the first stage of roasted product is 6:1, the leaching temperature is 60℃, and the leaching time is 60min.

[0032] S3: After drying the first stage of filter residue at 120℃, it is mixed evenly with the aluminum-silicon oxide obtained after removing impurities from mullite (the molar ratio of fluorine in the anode covering material to silicon in the mullite is 4:1), and a two-stage calcination reaction is carried out at 800℃ for 240 minutes to obtain pure alumina product and silicon tetrafluoride gas. The purity of the pure alumina product is 98.8%. The purity of the silicon tetrafluoride gas is 99.86%.

[0033] S4: Add sodium carbonate solution to the first filtrate and heat and stir at 80℃. Filter to obtain lithium carbonate product and delithiation filtrate. The lithium carbonate product has a purity of 98.8%. Evaporate the delithiation filtrate at 80℃ to obtain sodium sulfate product.

[0034] Example 3

[0035] This embodiment uses method one for processing.

[0036] S1: The crushed waste electrolyte from electrolytic aluminum is mixed evenly with aluminum nitrate and subjected to a first-stage roasting reaction to obtain a first-stage roasting product. The mass ratio of waste electrolyte to aluminum nitrate is 1:4, the roasting temperature is 600℃, and the roasting time is 60min.

[0037] S2: The first stage of roasted product is subjected to acid leaching and water leaching treatments sequentially, and filtered to obtain a first stage of filter residue and a first stage of filtrate. The leaching agent used for acid leaching is dilute nitric acid with a concentration of 1 mol / L-4 mol / L; in this embodiment, 2 mol / L is used. The liquid-solid ratio of the leaching agent to the first stage of roasted product is 10:1, the leaching temperature is 20℃, and the leaching time is 480 min. The filter residue after acid leaching is then subjected to water leaching treatment, and the water leaching process is carried out until the pH of the washing solution reaches 7.

[0038] S3: After drying the first stage of filter residue at 150℃, it is mixed evenly with the aluminum-silicon oxide obtained after removing impurities from fly ash (the molar ratio of fluorine in the waste electrolyte to silicon in the fly ash is 4:10), and a two-stage calcination reaction is carried out at 1200℃ for 30 minutes to obtain an alumina-silica mixture and silicon tetrafluoride gas. The purity of the prepared alumina-silica mixture is 98.6%. The purity of the silicon tetrafluoride gas is 99.82%.

[0039] S4: Add sodium carbonate to the first filtrate and heat and stir at 100℃. Filter to obtain lithium carbonate product and delithiation filtrate. The lithium carbonate product has a purity of 98.5%. Evaporate the delithiation filtrate at 120℃ to obtain sodium nitrate product.

[0040] Example 4

[0041] This embodiment uses method one for processing.

[0042] S1: The crushed bottom precipitate of the electrolytic aluminum furnace is mixed evenly with aluminum sulfate and subjected to a first-stage roasting reaction to obtain a first-stage roasting product. The mass ratio of bottom precipitate to aluminum sulfate is 1:1.5, the roasting temperature is 600℃, and the roasting time is 180min.

[0043] S2: A section of the roasted product is leached, and filtered to obtain a section of residue and a section of filtrate. The leaching agent is water. The liquid-to-solid ratio of the leaching agent to the roasted product is 5:1, the leaching temperature is 80℃, and the leaching time is 40 min.

[0044] S3: After drying the filter residue at 110℃, it is mixed evenly with the aluminum-silicon oxide obtained after impurity removal from coal gangue (the molar ratio of fluorine in the furnace bottom precipitate to silicon in the aluminum-silicon oxide is 4:1), and a two-stage calcination reaction is carried out at 1000℃ for 180 min to obtain pure alumina product and silicon tetrafluoride gas. The XRD pattern of the obtained pure alumina product is shown below. Figure 3 As shown, the purity of the pure alumina product is 98.7%. The purity of the silicon tetrafluoride gas is 99.83%.

[0045] S4: Add sodium carbonate solution to the first filtrate and heat and stir at 70℃. Filter to obtain lithium carbonate product and delithiation filtrate. The purity of the lithium carbonate product is 98.6%. Evaporate the delithiation filtrate at 110℃ to obtain sodium sulfate product.

[0046] Example 5

[0047] This embodiment uses method two for processing.

[0048] S1: The crushed electrolytic aluminum anode covering material, aluminum sulfate, and silicon dioxide are mixed evenly (the mass ratio of anode covering material to aluminum sulfate is 1:1, and the molar ratio of fluorine in the anode covering material to silicon in the silicon dioxide is 4:1). A first-stage roasting reaction is performed, followed by a second-stage roasting reaction, ultimately yielding the second-stage roasting product and silicon tetrafluoride gas. The first-stage roasting temperature is 500℃, and the roasting time is 240 min; the second-stage roasting temperature is 900℃, and the roasting time is 240 min. The purity of the silicon tetrafluoride gas is 99.81%.

[0049] S2: The second-stage roasting product was leached, and filtered to obtain second-stage filter residue and second-stage filtrate. Water was used as the leaching agent. The liquid-to-solid ratio of the leaching agent to the second-stage roasting product was 7:1, the leaching temperature was 70℃, and the leaching time was 40 min. The second-stage filter residue was dried at 140℃ to obtain pure alumina. The purity of the prepared pure alumina product was 99.4%.

[0050] S3: Add sodium carbonate solution to the second-stage filtrate and heat and stir at 70°C. Filter to obtain lithium carbonate product and delithiation filtrate. The lithium carbonate product has a purity of 98.0%. Evaporate the delithiation filtrate at 90°C to obtain sodium sulfate product.

[0051] Example 6

[0052] This embodiment uses method two for processing.

[0053] S1: The crushed waste electrolyte from electrolytic aluminum, aluminum sulfate, and aluminum-silicon oxide obtained after removing impurities from fly ash are mixed evenly (the mass ratio of waste electrolyte to aluminum sulfate is 1:2, and the molar ratio of fluorine in the waste electrolyte to silicon in the aluminum-silicon oxide is 4:5). A first-stage roasting reaction is then carried out, followed by a second-stage roasting reaction, ultimately yielding the second-stage roasting product and silicon tetrafluoride gas. The first-stage roasting temperature is 600℃, and the roasting time is 180 min; the second-stage roasting temperature is 1000℃, and the roasting time is 180 min. The purity of the silicon tetrafluoride gas is 99.85%.

[0054] S2: The second-stage roasting product was leached, and filtered to obtain a second-stage filter residue and a second-stage filtrate. The leaching agent was a sodium chloride solution with a concentration of 3 mol / L. The liquid-to-solid ratio of the leaching agent to the second-stage roasting product was 8:1, the leaching temperature was 70℃, and the leaching time was 40 min. The second-stage filter residue was dried at 130℃ to obtain an alumina-silica mixture. The purity of the prepared alumina-silica mixture product was 98.9%.

[0055] S3: Add sodium carbonate solution to the second-stage filtrate and heat and stir at 90℃. Filter to obtain lithium carbonate product and delithiation filtrate. The purity of the lithium carbonate product is 98.2%. Evaporate the delithiation filtrate at 100℃ to obtain sodium sulfate product.

[0056] Example 7

[0057] This embodiment uses method two for processing.

[0058] S1: The crushed bottom precipitate of the electrolytic aluminum furnace, aluminum sulfate, and aluminum-silicon oxide obtained after impurity removal from coal gangue are mixed evenly (the mass ratio of bottom precipitate to aluminum sulfate is 1:3, and the molar ratio of fluorine in the bottom precipitate to silicon in the aluminum-silicon oxide is 4:10). A first-stage roasting reaction is then carried out, followed by a second-stage roasting reaction, ultimately yielding the second-stage roasting product and silicon tetrafluoride gas. The first-stage roasting temperature is 550℃, and the roasting time is 210 min; the second-stage roasting temperature is 1200℃, and the roasting time is 120 min. The purity of the silicon tetrafluoride gas is 99.80%.

[0059] S2: The two-stage roasting products were sequentially subjected to acid leaching and water leaching, and filtered to obtain two-stage filter residue and two-stage filtrate. The leaching agent used for acid leaching was dilute sulfuric acid with a concentration of 3 mol / L. The liquid-to-solid ratio of the leaching agent to the two-stage roasting products was 5:1, the leaching temperature was 80℃, and the leaching time was 30 min. The filter residue after acid leaching was then subjected to water leaching, washing until the pH of the washing solution reached 7. After drying the two-stage filter residue at 120℃, an alumina-silica mixture was obtained. The purity of the prepared alumina-silica mixture product was 99.3%.

[0060] S3: Add sodium carbonate solution to the second-stage filtrate and heat and stir at 80℃. Filter to obtain lithium carbonate product and delithiation filtrate. The lithium carbonate product has a purity of 98.1%. Evaporate the delithiation filtrate at 120℃ to obtain sodium sulfate product.

Claims

1. A method for the resource recovery of complex electrolyte hazardous waste rich in alumina, characterized in that, The process includes the following: crushed, alumina-rich complex electrolyte hazardous waste is uniformly mixed with aluminum salts and silicon oxides, and then subjected to a staged roasting-leaching treatment to obtain alumina products and silicon tetrafluoride gas; the staged roasting-leaching treatment has two methods. Method 1: One-stage roasting-leaching-two-stage roasting, including the following: the complex electrolyte hazardous waste is mixed with the aluminum salt, and after one-stage roasting, a one-stage roasting product is obtained. The one-stage roasting product is leached to obtain a one-stage filter residue and a one-stage filtrate. The one-stage filter residue is then roasted in two stages to obtain alumina product and silicon tetrafluoride gas. Method 2: First-stage roasting-second-stage roasting-leaching, including the following: the complex electrolyte hazardous waste is mixed with the aluminum salt and the silicon-containing oxide, and first-stage roasting is performed, followed by second-stage roasting, to obtain the second-stage roasting product and silicon tetrafluoride gas. The second-stage roasting product is leached to obtain the second-stage filter residue, i.e., alumina product and the second-stage filtrate.

2. The method for resource recovery of complex electrolyte hazardous waste rich in alumina according to claim 1, characterized in that, The first-stage roasting temperature is 300℃-600℃, and the time is 60min-480min; the second-stage roasting temperature is 600℃-1200℃, and the time is 30min-480min.

3. The method for resource recovery of complex electrolyte hazardous waste rich in alumina according to claim 1, characterized in that, The complex electrolyte hazardous waste mainly contains alumina, as well as fluorine, lithium, and sodium elements. After a first-stage roasting, the fluorine is converted into aluminum fluoride, and the lithium and sodium elements are converted into soluble salts. The alumina does not participate in the reaction and together they form the first-stage roasting product. After a second-stage roasting, the fluorine is converted into silicon tetrafluoride gas, thus achieving the separation of fluorine and aluminum.

4. The method for resource recovery of complex electrolyte hazardous waste rich in alumina according to claim 1, characterized in that, The complex electrolyte hazardous waste is one or more of the following: furnace bottom sediment, anode covering material, and waste electrolyte during the electrolytic production of aluminum or aluminum alloys; the aluminum salt includes aluminum chloride, aluminum sulfate, and aluminum nitrate; the silicon oxide is one or more of the following: mullite, fly ash, coal gangue solid waste impurities obtained after purification, aluminum silicon oxide, and silicon dioxide.

5. The method for resource recovery of complex electrolyte hazardous waste rich in alumina according to claim 4, characterized in that, The aluminum-silicon oxide contains ≤0.2% iron oxide, ≤0.1% calcium oxide, and ≤0.02% magnesium oxide, with the remainder being aluminum oxide and silicon dioxide.

6. The method for resource recovery of complex electrolyte hazardous waste rich in alumina according to claim 1, characterized in that, The alumina product is pure alumina or an alumina-silicon dioxide mixture with a purity ≥98.6%, and is used for aluminum electrolysis or aluminum-silicon alloy electrolysis; the silicon tetrafluoride gas has a purity ≥99.80%.

7. The method for resource recovery of complex electrolyte hazardous waste rich in alumina according to claim 1, characterized in that, The mass ratio of the complex electrolyte hazardous waste to the aluminum salt is 1:0.25-1:4; the molar ratio of fluorine in the complex electrolyte hazardous waste to silicon in the silicon-containing oxide is 4:(1-10).

8. The method for resource recovery of complex electrolyte hazardous waste rich in alumina according to claim 7, characterized in that, When the molar ratio of fluorine to silicon in the complex electrolyte hazardous waste is 4:1, the alumina product is pure alumina with a purity ≥98.6%, and is used for aluminum electrolysis.

9. A method for resource recovery of complex electrolyte hazardous waste rich in alumina according to claim 1, characterized in that, The leaching agent for the leaching treatment is one or more of water, salt solution, and dilute acid solution; the salt solution is one or more of sodium salt solution and ammonium salt solution; the dilute acid solution is one or more of dilute hydrochloric acid, dilute sulfuric acid, and dilute nitric acid; the liquid-solid ratio of the leaching agent to the first-stage roasted product or the second-stage roasted product is (4-10):1, the leaching temperature is 20℃-95℃, and the leaching time is 10min-480min.

10. A method for resource recovery of complex electrolyte hazardous waste rich in alumina according to claim 1, characterized in that, The first or second stage filtrate is processed as follows: sodium carbonate is added to the first or second stage filtrate and heated and stirred at a temperature of 50℃-100℃. The filtrate is then filtered to obtain lithium carbonate product and delithiation filtrate, with the lithium carbonate product having a purity of ≥98.0%. The delithiation filtrate is then evaporated and crystallized at 50℃-120℃ to obtain sodium salt product.

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