Resource recycling method for electrolytic aluminum overhaul slag

Abstract

The invention discloses a resource recycling method for electrolytic aluminum overhaul slag, and relates to the technical field of resource recycling. The recovery method comprises the following steps: crushing the overhaul slag, stirring with water, washing, and separating to obtain water washing liquid and washing slag; washing residues are leached by adopting an acid solution in a synergistic heating manner, cryolite and other substances in the washing residues are efficiently dissolved, and leached residues are mainly carbon materials. And hydrogen fluoride gas generated in the leaching process is introduced into alkaline water washing liquid, aluminum salt is added for impurity removal, and a sodium fluoride product is obtained through evaporative crystallization. And adding ammonia water into the leachate to adjust the pH value, cooling and crystallizing to obtain an aluminum ammonium sulfate product for recovering aluminum. And the separated mother liquor is subjected to impurity removal, alkalization, evaporation and concentration, ammonia gas is recovered, and a lithium-rich concentrated solution is obtained. Extracting lithium from the concentrated solution, and evaporating raffinate to obtain a sodium sulfate product; and the lithium-loaded organic phase is subjected to reverse extraction by phosphoric acid, and a lithium dihydrogen phosphate product is directly obtained after reverse extraction liquid is subjected to pH-controlled evaporation. According to the invention, collaborative recovery and high-value utilization of fluorine, aluminum, sodium and lithium resources are realized.

Description

Technical Field

[0001] This invention relates to the field of resource recycling technology, and in particular to a method for the resource recovery of electrolytic aluminum overhaul slag. Background Technology

[0002] Waste residue generated during the maintenance and replacement of cathode liners in electrolytic cells during aluminum production is called overhaul residue. Because this type of waste residue contains toxic components such as fluorides, it is classified as hazardous waste and poses a serious threat to the environment if not properly disposed of.

[0003] Currently, the harmless treatment technology for overhaul slag is relatively mature, and the treated residue is often used as a raw material for cement or brick making. At the same time, overhaul slag is also rich in various components such as fluorine, sodium, lithium, aluminum, and carbon materials, possessing considerable resource recycling potential. Effectively extracting valuable elements from it can not only turn waste into treasure and alleviate some of the pressure of resource scarcity, but also fundamentally reduce environmental pollution and promote resource recycling and green sustainable development.

[0004] Regarding resource utilization, existing research has proposed various technical approaches, but significant limitations remain. For example, patent CN117327923A proposes a method for jointly extracting lithium from waste aluminum electrolyte and overhaul slag, but this process only focuses on lithium recovery and the preparation of lithium carbonate, without considering other valuable resources in the slag. Another patent CN117327923A achieves the separate recovery of aluminum slag, carbon blocks, sodium fluoride, etc. through sorting, wet grinding, and electrolysis; however, the value of its recovered products is limited, and it does not effectively extract high-value elements such as lithium and aluminum.

[0005] Overall, current resource recovery processes for overhaul slag are mostly limited to recovering individual elements, failing to achieve the synergistic and efficient recovery of multiple high-value components in the slag. Therefore, developing a comprehensive process that can fully and economically recover valuable elements from overhaul slag, especially high-value resources such as lithium and aluminum, has become a critical issue that urgently needs to be addressed. Summary of the Invention

[0006] The technical problem to be solved by this invention is that existing resource recovery processes for overhaul slag are mostly limited to the recovery of individual elements, and fail to achieve the synergistic and efficient recovery of multiple high-value components in the slag.

[0007] To address the above problems, the present invention proposes the following technical solution: A method for resource recovery of electrolytic aluminum overhaul slag includes the following steps: S1. Crushing and washing pretreatment: After crushing the slag from the major overhaul of electrolytic aluminum, it is mixed with water, stirred and washed, and then separated into solid and liquid to obtain washing liquid and washing slag. S2. Preparation of sodium fluoride: Acidic gas containing hydrogen fluoride is introduced into the washing solution to adjust the pH of the system to 7-9. Aluminum salt is added to remove impurities to obtain a sodium fluoride solution. After evaporation and crystallization, the sodium fluoride product is obtained. S3, Acid Leaching: The washing residue and acidic liquid are reacted at a solid-liquid mass ratio of 1:3-5 under pH 0-1 conditions. After the reaction is completed, solid-liquid separation is performed to obtain acidic leachate and carbon residue. S4. Aluminum resource recovery: Add an ammonium-containing reagent to the acidic leachate, adjust the pH of the system to 2-4, cool and crystallize, and obtain ammonium aluminum sulfate product after solid-liquid separation; S5. Solution purification and concentration: After the mother liquor from which the aluminum ammonium sulfate product is separated is filtered to remove impurities, the pH of the system is adjusted to 11-13 for alkalization treatment, and then concentrated by evaporation to obtain lithium-rich concentrate and recover the volatilized ammonia. S6. Lithium Resource Recovery: The lithium-rich concentrate is extracted to separate a lithium-containing organic phase and a raffinate; the raffinate is evaporated and crystallized to obtain sodium sulfate product; the lithium-containing organic phase is back-extracted with a phosphoric acid solution to separate a back-extraction solution containing lithium dihydrogen phosphate and a regenerated organic phase; the back-extraction solution is evaporated and crystallized to obtain lithium dihydrogen phosphate product.

[0008] A further technical solution is that, in step S1, the particle size of the crushed slag is less than or equal to 100 mesh; the solid-liquid mass ratio of the stirring and washing is 1:3-5, and the stirring is carried out at room temperature for 1-2 hours.

[0009] A further technical solution is that, in step S2, the acidic gas containing hydrogen fluoride includes the gas generated in the reaction in step S3.

[0010] A further technical solution is that the aluminum salt is one of aluminum sulfate, polyaluminum chloride, or sodium aluminate.

[0011] A further technical solution is that, in step S3, the acidic solution is a mixed solution of sulfuric acid and ammonium sulfate, wherein the molar ratio of sulfuric acid to ammonium sulfate is 2:1-4:1.

[0012] The further technical solution is that in step S3, the reaction temperature is 90-110℃ and the reaction time is 2-4 hours.

[0013] A further technical solution is that, in step S4, the ammonium-containing reagent is ammonia water; the cooling crystallization temperature is 0-10℃, and the crystallization time is 1-2 hours.

[0014] A further technical solution is that the ammonia recovered in step S5 is returned to step S4 for recycling; and the organic phase regenerated in step S6 is returned to the extraction treatment step of step S6 for recycling.

[0015] The further technical solution is that step S5 specifically includes: adjusting the pH of the mother liquor after separating the aluminum ammonium sulfate product to 6-8 to precipitate and remove iron and aluminum impurities, and filtering; then adjusting the pH of the filtrate to 11-13, and then evaporating and concentrating it.

[0016] A further technical solution is that, in step S6, the volume ratio of the organic phase to the lithium-rich concentrate in the extraction process is 3:1-5:1; in the back-extraction, the concentration of the phosphoric acid solution is 5-8 wt%, and the volume ratio of the organic phase to the phosphoric acid solution is 1:5-1:10; when preparing lithium dihydrogen phosphate products, the pH of the evaporation crystallization mother liquor is controlled to be 1.5-2.5.

[0017] Compared with the prior art, the beneficial effects of the present invention are as follows: The present invention provides a method for the resource recovery of electrolytic aluminum overhaul slag. Through sequential steps including crushing and washing, acid leaching, aluminum recovery, concentration, and lithium recovery, the method converts multiple elements such as fluorine, aluminum, sodium, and lithium in the overhaul slag into specific products such as sodium fluoride, ammonium aluminum sulfate, sodium sulfate, and lithium dihydrogen phosphate. This method offers significant economic value and allows for the internal reuse of hydrogen fluoride vapor and ammonia generated during the process. It overcomes the limitations of existing technologies that are mostly limited to recovering only one or two elements, significantly improving the comprehensive resource utilization rate of single hazardous waste sources.

[0018] Compared to the traditional high-temperature leaching process using concentrated sulfuric acid (which requires additional sodium hydroxide treatment for the hydrogen fluoride tail gas), this invention adds a water washing pretreatment step. This step dissolves soluble sodium fluoride, sodium hydroxide, etc., in the overhaul slag, forming a strongly alkaline washing solution. The resulting strongly alkaline washing solution has a dual function: firstly, it significantly reduces the consumption of concentrated sulfuric acid in the washing slag during subsequent acid leaching; secondly, it can be directly used as an absorbent for the hydrogen fluoride tail gas, greatly replacing or reducing the use of purchased sodium hydroxide, thus maximizing resource utilization while effectively reducing production costs.

[0019] The hydrogen fluoride gas produced by acid leaching is absorbed by the aforementioned washing solution, which increases the yield of sodium fluoride while completing the tail gas treatment. After absorption, aluminum salt is added to the solution to remove impurities, which can remove silicon ions, and the resulting sodium aluminosilicate colloid can adsorb pigments, so that the obtained sodium fluoride product meets the NF-1 standard requirements in terms of purity and color.

[0020] This invention presents a process for preparing sodium fluoride from a washing solution, followed by cooling and crystallization to recover ammonium aluminum sulfate. Compared to traditional cryolite recovery processes, this significantly improves lithium recovery rates. Traditional cryolite synthesis processes, within the pH range of 5.5-7, easily generate aluminum hydroxide colloids (referencing lithium extraction processes from lithium salt lakes). These colloids form complexes or encapsulate lithium, leading to increased lithium loss. The process route employed in this invention can significantly improve lithium resource recovery rates.

[0021] In traditional lithium carbonate synthesis, a significant amount of lithium remains in the mother liquor, requiring acidification to remove carbon and recover sodium sulfate, followed by lithium recovery from the mother liquor. The lithium carbonate is then used to prepare lithium dihydrogen phosphate. This invention uses an extraction method to directly extract and separate lithium from a concentrated lithium sulfate solution. The remaining sodium sulfate solution is then evaporated and concentrated to prepare sodium sulfate. The organic phase after lithium extraction is used to directly prepare lithium dihydrogen phosphate by evaporating and concentrating the resulting lithium dihydrogen phosphate solution. This invention has a shorter process and a higher lithium recovery rate. Furthermore, lithium dihydrogen phosphate, as a battery material precursor, offers advantages such as low energy consumption, strong adaptability, low overall cost, and stable battery performance, resulting in significant economic value. The remaining solution after lithium extraction can be directly evaporated to prepare sodium sulfate, achieving full utilization of the materials. Attached Figure Description

[0022] To more clearly illustrate the technical solutions of the embodiments of the present invention, the drawings used in the following description of the embodiments will be briefly introduced. Obviously, the drawings described below are some embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.

[0023] Figure 1 This is a schematic flowchart of the method for resource recovery of electrolytic aluminum overhaul slag provided in an embodiment of the present invention; Figure 2 A detailed flowchart of the method for resource recovery of electrolytic aluminum overhaul slag provided in this embodiment of the invention. Detailed Implementation

[0024] The technical solutions of the embodiments of the present invention will be clearly and completely described below with reference to the accompanying drawings. Obviously, the described embodiments are only some, not all, of the embodiments of the present invention. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0025] It should be understood that, when used in this specification and the appended claims, the terms "comprising" and "including" indicate the presence of the described features, integrals, steps, operations, elements and / or components, but do not exclude the presence or addition of one or more other features, integrals, steps, operations, elements, components and / or collections thereof.

[0026] It should also be understood that the terminology used in this specification is for the purpose of describing particular embodiments only and is not intended to limit the invention. As used in this specification and the appended claims, the singular forms “a,” “an,” and “the” are intended to include the plural forms unless the context clearly indicates otherwise.

[0027] It should also be further understood that the term "and / or" as used in this specification and the appended claims refers to any combination of one or more of the associated listed items and all possible combinations, and includes such combinations.

[0028] See Figure 1 and Figure 2 This invention provides a method for the resource recovery of electrolytic aluminum overhaul slag, comprising the following steps: S1. Crushing and washing pretreatment: After crushing the electrolytic aluminum overhaul slag, it is mixed with water, stirred and washed, and then separated into solid and liquid to obtain washing liquid (liquid 1) and washing slag (slag 1).

[0029] In practice, the slag from the overhaul of electrolytic aluminum is mechanically crushed and ball-milled until it passes through a 100-mesh sieve (approximately 0.15 mm) to obtain powdered material. Smaller particle size significantly increases the specific surface area, promoting the dissolution of soluble substances during the washing process. Excessively coarse particle size reduces dissolution efficiency, while excessively fine particle size increases the difficulty and energy consumption of subsequent solid-liquid separation.

[0030] Mix the powder with water at a solid-liquid mass ratio of 1:3-5 (e.g., 1:3, 1:4, 1:5), stir at room temperature for 1-2 hours to fully dissolve the soluble sodium fluoride and sodium hydroxide, and then perform solid-liquid separation.

[0031] The washing solution is a strongly alkaline solution rich in sodium fluoride. The washing residue mainly consists of carbon materials, cryolite (Na3AlF6), lithium-containing compounds, and other sparingly soluble or insoluble substances.

[0032] S2. Preparation of sodium fluoride: Acidic gas containing hydrogen fluoride is introduced into the washing solution (liquid 1) to adjust the pH of the system to 7-9. Aluminum salt is added to remove impurities to obtain sodium fluoride solution (liquid 2). After evaporation and crystallization, sodium fluoride product is obtained.

[0033] In practice, the acidic gas mainly originates from the hydrogen fluoride (HF) gas generated during the acidic leaching process in step S3. HF reacts with OH- in the alkaline washing solution. - The reaction produces F - Adding water not only treats the waste gas but also increases the concentration of sodium fluoride in the solution.

[0034] Adding aluminum salts removes silicon ions and pigments from the solution. A weakly alkaline environment (pH 7-9, e.g., pH 7, 8, 9) allows the aluminum salts to react fully with the silicon ions and prevents excessive hydrolysis, resulting in a purified sodium fluoride-rich solution. The purified sodium fluoride-rich solution is then evaporated to supersaturation under normal or reduced pressure, cooled, and crystallized to obtain industrial-grade sodium fluoride (NaF) product.

[0035] In specific implementation, the aluminum salt is aluminum sulfate, polyaluminum chloride, sodium aluminate, or other reagents that can provide aluminum ions. Its main function is to remove silicon and adsorb impurities such as pigments by generating sodium aluminosilicate. This invention does not strictly limit the specific type of aluminum salt; any aluminum-containing compound that can achieve similar impurity removal functions is applicable. The amount of aluminum salt is usually added in approximately equimolar amounts to the silicon impurities or in slight excess, for example, at a sodium silicate:aluminum salt molar ratio of 1:1-1.2 (e.g., 1:1, 1:2). The specific amount can be adjusted according to the silicon content in the washing solution to ensure the formation of aluminosilicate precipitates.

[0036] Taking sodium aluminate as an example, this step preferably uses the absorption of HF gas volatilized from acid leaching and heating to adjust the pH to neutral. Simultaneously, sodium aluminate is used to synergistically remove silicon. Sodium silicate reacts with hydrogen fluoride to form silicic acid precipitate, achieving the absorption of hydrogen fluoride tail gas and removal of impurities from the washing solution (sodium fluoride salts have silicon compound concentration requirements). The reactions involved include: 2HF + Na₂SiO₃ = 2NaF + H₂SiO₃↓ 2Na2SiO3+2NaAlO2+2H2O═Na2Al2Si2O8↓+4NaOH.

[0037] S3, Acid Leaching: The washing residue (residue 1) is reacted with acidic liquid at a solid-liquid mass ratio of 1:3-5 under pH 0-1 conditions. After the reaction is completed, solid-liquid separation is performed to obtain acidic leachate (liquid 3) and carbon residue (residue 2).

[0038] In practice, the acidic solution is a mixture of sulfuric acid and ammonium sulfate, wherein the molar ratio of sulfuric acid to ammonium sulfate is 2:1 to 4:1 (e.g., 2:1, 3:1, 4:1). In the mixed solution, sulfuric acid provides H₂. + NH4 in ammonium sulfate + Can be used with Al 3+ By forming complexes or double salt precursors, the decomposition of cryolite is promoted. Compared with conventional acid washing / leaching, the leaching rate of cryolite in this invention can reach 95%, which greatly improves the recovery rate of fluoride and aluminum in overhaul slag.

[0039] To fully demonstrate the advantages of using sulfuric acid and ammonium sulfate as the acidic solutions in this invention for cryolite, the following comparative experiments were conducted: Sodium fluoroaluminate was leached at 95°C for 2 hours using single sulfuric acid or hydrochloric acid solutions of different concentrations, then filtered, the solid was dried and weighed, and the solubility rate was calculated. The experimental results are shown in Table 1.

[0040] Table 1. Leaching results of sodium fluoroaluminate with different acidic solvents

[0041] The results showed that even when the sulfuric acid concentration was increased to 20%, or the liquid-to-solid ratio was significantly increased (to 1:15), the solubility of sodium fluoroaluminate was still only about 37%. This indicates that simply increasing the acidity or using excessive solvent is insufficient to efficiently destroy the cryolite structure.

[0042] The present invention introduces ammonium sulfate into sulfuric acid, which provides ammonium ions (NH4+) in a strongly acidic environment. + It can react with aluminum ions (Al) 3+ The formation of soluble aluminum ammonium complexes (such as NH4Al(SO4)2) greatly promotes the shift of the decomposition equilibrium of cryolite (Na3AlF6) towards dissolution, thereby increasing the leaching rate of cryolite to about 95%.

[0043] The main reactions involved in this step are: 2Na3AlF6 + 6H2SO4 = 3Na2SO4 + Al2(SO4)3 + 12HF↑, and (NH4)2SO4 + Al2(SO4)3 = 2NH4Al(SO4)2. During the acid leaching process, cryolite reacts with sulfuric acid to produce aluminum sulfate, and aluminum sulfate reacts with ammonium sulfate to produce ammonium aluminum sulfate. The solubility of ammonium aluminum sulfate is greater than that of aluminum sulfate, which promotes the reaction between cryolite and sulfuric acid, thereby increasing the leaching rate of cryolite in the acid leaching of electrolytic aluminum overhaul slag. Most of the HF generated in the reaction volatilizes into the gas phase and is collected and recycled in step S2, while a small portion remains in the liquid phase.

[0044] The present invention controls the system to a strongly acidic condition with a pH of 0-1 (e.g., pH 0, 0.5, 0.8, 1.0), which is beneficial for the decomposition of cryolite (Na3AlF6) and the leaching of aluminum and fluorine in the washing residue. Excessively high pH will lead to a decrease in the leaching rate.

[0045] In practice, appropriately increasing the temperature of the leaching system helps to improve the reaction rate, leaching equilibrium, and the volatilization and release of hydrogen fluoride gas. The suitable reaction temperature is 90-110℃ (e.g., 90℃, 100℃, 110℃), and the reaction time is 2-4 hours. If the temperature is too low, the reaction will be slow, and if it is too high, the energy consumption will increase and may aggravate equipment corrosion.

[0046] The acidic leachate obtained in this step mainly contains Al. 3+ Li + Na + SO4 2- F - Etc. Carbon slag is an unreacted carbon material that can be used as an auxiliary material in cement kilns.

[0047] S4. Aluminum resource recovery: Add an ammonium-containing reagent to the acidic leachate (liquid 3), adjust the pH of the system to 2-4, cool and crystallize, and obtain ammonium aluminum sulfate product after solid-liquid separation.

[0048] In practice, an ammonium-containing reagent (such as ammonia) is added to the acidic leachate obtained from S3 to adjust the pH to 2-4 (e.g., pH 2, 3, 3.5, 4). Then, the solution is cooled and crystallized at 0-10℃ for 1-2 hours, and the solid-liquid separation yields the aluminum ammonium sulfate (NH4Al(SO4)2·12H2O) product.

[0049] pH 2-4 is the stable crystallization region for ammonium aluminum sulfate, where aluminum crystals almost entirely in the form of Al. 3+ Alternatively, it can exist in the form of aluminum ammonium double salt to avoid the formation of Al(OH)3 colloids (which typically begin to form at pH > 4.5). Colloids adsorb lithium ions, resulting in lithium loss.

[0050] Generally speaking, the amount of ammonium-containing reagent added is expressed as n(NH4) + ) : n(Al 3+ A ratio of approximately 1-1.1:1 is recommended to ensure that aluminum fully forms a complex salt.

[0051] In practice, ammonia water can be obtained from the ammonia gas recovered in step S5.

[0052] S5. Solution purification and concentration: After the mother liquor (liquid 4) after separating the aluminum ammonium sulfate product is filtered to remove impurities, the pH of the system is adjusted to 11-13 for alkalization treatment. After evaporation and concentration, lithium-rich concentrate (liquid 6) is obtained and the volatilized ammonia is recovered.

[0053] In specific implementation, the mother liquor after separating ammonium aluminum sulfate in step S4 is first adjusted to pH 6-8 (e.g., pH 6, 7, 8) to remove residual Fe. 3+ Al 3+ After a hydroxide precipitate forms (slag 3, i.e., iron-containing slag), it is removed by filtration, yielding the filtrate (liquid 5). Li + Na + The filtrate (liquid 5) remains in solution (liquid 5); then adjust the pH of the filtrate (liquid 5) to 11-13 with NaOH (e.g., pH 11, 12, 13). Strongly alkaline conditions ensure the presence of residual NH4+. + It is converted into NH3, escapes, and is recovered (reused in step S4).

[0054] Evaporation and concentration are carried out under normal or reduced pressure to obtain lithium-rich concentrate. The final lithium concentration is controlled at approximately 10–15 g / L to provide a suitable concentration for subsequent extraction.

[0055] S6. Lithium Resource Recovery: The lithium-rich concentrate (liquid 6) is extracted to separate a lithium-containing organic phase (organic phase 1) and a raffinate (liquid 7); the raffinate is evaporated and crystallized to obtain sodium sulfate product; the lithium-containing organic phase is back-extracted with a phosphoric acid solution to separate a back-extraction solution (liquid 8) containing lithium dihydrogen phosphate and a regenerated organic phase (organic phase 2); the back-extraction solution is evaporated and crystallized to obtain lithium dihydrogen phosphate product.

[0056] In specific implementation, the lithium-rich concentrate obtained from S5 is mixed with the extraction organic phase (such as tributyl phosphate-kerosene system) at a volume ratio of 1:3-5 (e.g., 1:3, 1:4, 1:5) and subjected to multi-stage extraction. Lithium enters the organic phase, resulting in a lithium-containing organic phase and raffinate (mainly containing Na2SO4).

[0057] The raffinate was evaporated and crystallized to obtain sodium sulfate product.

[0058] Back-extraction is performed using a 5-8% (w / w) phosphoric acid solution and a lithium-containing organic phase at a volume ratio of 5-10:1 (aqueous phase:organic phase). Lithium is transferred to the aqueous phase, yielding a back-extraction solution containing lithium dihydrogen phosphate. The organic phase is then regenerated and reused. A phosphoric acid solution concentration of 5-8% is preferable; too low a concentration results in poor back-extraction efficiency, while too high a concentration may introduce impurities. Those skilled in the art can adjust the phosphoric acid concentration appropriately based on the lithium loading in the organic phase, the back-extraction efficiency, and the purity requirements of the target product.

[0059] The back-extraction solution was evaporated and concentrated at pH 1.5–2.5, which is the stable crystallization region of LiH2PO4. If the pH is too high, Li2HPO4 or other lithium salts may be formed, while if the pH is too low, the crystallization rate will decrease. After cooling and crystallization, battery-grade lithium dihydrogen phosphate (LiH2PO4) is obtained.

[0060] Understandably, in specific implementations, the extractant can be tributyl phosphate (TBP), di(2-ethylhexyl) phosphate (P204), naphthenic acid, or a mixture thereof with diluents such as kerosene or sulfonated kerosene. The specific type and ratio can be selected and optimized based on the extraction efficiency and selectivity of lithium ions, but this invention is not limited thereto. In actual production, those skilled in the art can select other organic reagents or combinations thereof suitable for lithium extraction based on the solution composition and process conditions.

[0061] This invention achieves the stepwise separation and high-value recovery of fluorine, aluminum, sodium, and lithium through precise control of parameters such as pH and internal material circulation. Simultaneously, it reuses waste gas (HF), waste heat, ammonia, and organic phases within the system, forming a highly efficient, low-consumption, and low-emission resource-based process route. The parameters for each step are optimized based on reaction thermodynamics and kinetics, ensuring high recovery rates and product quality while also considering operational feasibility and economic efficiency.

[0062] It should be noted that the pH ranges involved in each step of the method of the present invention are preferred examples based on specific material systems and process conditions. In practical applications, the pH ranges can be reasonably adjusted near these ranges according to the specific ionic composition and requirements of the solution. The present invention is not limited to the specific values ​​listed. Those skilled in the art will understand that any specific pH value selected within the defined range can achieve the purpose of the present invention.

[0063] To verify the actual effect and parameter feasibility of the aforementioned process route of the present invention, the following detailed description of the resource recovery method for electrolytic aluminum overhaul slag of the present invention is provided in conjunction with specific embodiments. The raw material used in the embodiments is electrolytic aluminum overhaul slag produced by a certain enterprise, and its typical element mass percentage content is: F 16.2%, Al 8.9%, Na 15.5%, Fe 0.84%, Li 1.18%, Si 2.01%, K 1.44%, Ca 2.62%, with the remainder being carbon powder and dust.

[0064] Example 1 The mass percentage of each element in the electrolytic aluminum overhaul slag raw material processed in this embodiment is as follows: F 16.2%, Al 8.9%, Na 15.5%, Fe 0.84%, Li 1.18%, Si 2.01%, K 1.44%, Ca 2.62%, with the remainder being carbon powder and dust.

[0065] S1. Crushing and Water Washing Pretreatment: The electrolytic aluminum overhaul slag is crushed, ball-milled, and passed through a 100-mesh sieve to obtain powder. 500.0g of the powder is weighed and mixed with 1.5L of deionized water (solid-liquid mass ratio of approximately 1:3). After stirring at room temperature for 2 hours, the solid and liquid are separated to obtain water washing liquid and washing slag.

[0066] S2, Sodium fluoride preparation: The hydrofluoric acid vapor generated in subsequent step S3 is passed into the above washing solution to adjust the pH of the system to 7.70. 0.58 g of sodium aluminate is added, and the mixture is stirred at room temperature for 2 hours. After filtration to remove trace impurities such as silicon and pigments, a purified sodium fluoride solution is obtained. This solution is evaporated and crystallized to obtain the sodium fluoride product with a purity of 95.65%, meeting the secondary standard of YS / T517-2009. The fluorine recovery rate is 85.8%.

[0067] S3, Acid Leaching: The washing residue obtained in S1 is mixed with acidic solution (a mixed solution prepared from 7.2% sulfuric acid, 3.4% ammonium sulfate, and water) at a solid-liquid mass ratio of approximately 1:4 (500g of residue corresponds to 2L of leachate), and reacted at 95℃ and pH 0-1 for 2 hours. After the reaction, the solid and liquid are separated to obtain carbon slag (used as an auxiliary material for cement kilns and activated carbon material) and acidic leachate.

[0068] S4. Aluminum Resource Recovery: Add 140 mL of 25% ammonia solution to the above acidic leachate to adjust the pH of the system to 2-4. Then, while slowly stirring, cool the solution to 4°C and react for 1 hour. After solid-liquid separation, aluminum ammonium sulfate dodecahydrate (industrial grade) is obtained, with an aluminum recovery rate of 90.7%.

[0069] S5. Solution Purification and Concentration: The mother liquor after separating ammonium aluminum sulfate is adjusted to pH 6.8 with sodium hydroxide, filtered to remove trace amounts of impurities such as iron and aluminum; then sodium hydroxide is added again to adjust the pH to 12.5 for alkalization treatment, followed by filtration. The filtrate is evaporated and concentrated to obtain a lithium-rich concentrate with a lithium concentration of 12.5 g / L, and the volatilized ammonia gas is recovered and reused in step S4.

[0070] S6. Lithium Resource Recovery: The lithium-rich concentrate is extracted at a volume ratio of 1:4 for 10 minutes. After extraction, the lithium-containing organic phase and raffinate are separated. The raffinate is evaporated and crystallized to obtain sodium sulfate product with a purity of 99.7% (meeting GB / T 6009-2014 Class I superior grade). The lithium-containing organic phase is back-extracted with phosphoric acid solution at a volume ratio of 1:6 for 10 minutes. After extraction, the extractant and phosphoric acid solution are separated. The back-extracted phosphoric acid solution is evaporated and concentrated, and the pH of the concentrated mother liquor is controlled at 1.8 to obtain battery-grade lithium dihydrogen phosphate product with a purity of 99.6% (YS / T 967-2014). The total lithium recovery rate is 91.6%. The organic phase extractant after back-extraction is recycled for the extraction step.

[0071] In the final product obtained in this embodiment, the purity of the battery-grade lithium dihydrogen phosphate product is 99.6% and the lithium recovery rate is 91.6%; the fluorine content of the sodium fluoride product is 43.3% and the fluorine recovery rate is 85.8%; the aluminum ammonium sulfate hydrate product has an aluminum content of 5.9% and an aluminum recovery rate of 90.7%.

[0072] Example 2 The raw materials used in this embodiment are the same as those in Embodiment 1.

[0073] S1. Crushing and Water Washing Pretreatment: The electrolytic aluminum overhaul slag is crushed, ball-milled, and passed through a 100-mesh sieve to obtain powder. 500.0g of the powder is weighed and mixed with 1.5L of deionized water (solid-liquid mass ratio of approximately 1:3). After stirring at room temperature for 2 hours, the solid and liquid are separated to obtain water washing liquid and washing slag.

[0074] S2. Preparation of sodium fluoride: Hydrofluoric acid vapor was introduced to adjust the pH of the washing solution to 6.92. 0.63g of sodium aluminate was added to remove impurities, and the solution was filtered, evaporated, and crystallized to obtain sodium fluoride product with a purity of 96.72% (YS / T 517-2009 secondary standard) and a fluorine recovery rate of 87.8%.

[0075] S3, Acid Leaching: The washing residue is mixed with an acidic solution (prepared with 7.0% sulfuric acid, 2.5% ammonium sulfate, and water) at a solid-liquid ratio of approximately 1:4 (500g of residue corresponds to 2L of leachate). The mixture is reacted at 100℃ and pH 0-1 for 2 hours, followed by solid-liquid separation. After solid-liquid separation, carbon residue and acidic leachate are obtained.

[0076] S4. Aluminum Resource Recovery: Add 140mL of 25% ammonia water to the leaching solution, adjust the pH to 2-4, and slowly stir while cooling to 4℃. After reacting for 1 hour, crystals precipitate. Separate the solid and liquid to obtain ammonium aluminum sulfate product (industrial grade). The aluminum recovery rate is 91.3%.

[0077] S5. Solution purification and concentration: Add sodium hydroxide to the mother liquor to adjust the pH to 7.2, filter to remove impurities, add sodium hydroxide again to adjust the pH to 12.4, filter again, evaporate and concentrate the filtrate until the lithium concentration in the mother liquor is 12.24 g / L, and recover ammonia gas for use in step S4.

[0078] S6. Lithium Resource Recovery: The concentrated mother liquor is extracted at a volume ratio of 1:5 for 10 minutes. After extraction, the lithium-containing organic phase and raffinate are separated. The raffinate is evaporated and crystallized to obtain sodium sulfate with a purity of 99.6%. The lithium-containing organic phase is back-extracted with phosphoric acid solution at a volume ratio of 1:8 for 10 minutes. After extraction, the extractant and phosphoric acid solution are separated. The back-extracted phosphoric acid solution is evaporated and concentrated, and the pH of the concentrated mother liquor is controlled at 2.3 to obtain battery-grade lithium dihydrogen phosphate with a purity of 99.5%. The total lithium recovery rate is 92.1%. The organic phase is reused.

[0079] In this embodiment, the final product has the following properties: battery-grade lithium dihydrogen phosphate with a purity of 99.5% and a lithium recovery rate of 92.1%; sodium fluoride with a fluorine content of 43.5% and a fluorine recovery rate of 87.8%; and aluminum ammonium sulfate hydrate with an aluminum content of 5.86% and an aluminum recovery rate of 91.3%.

[0080] As can be seen from the above embodiments, the electrolytic aluminum overhaul slag treated by the method of the present invention can be used to prepare cement kiln auxiliary materials, sodium fluoride, ammonium aluminum sulfate, sodium sulfate, and lithium dihydrogen phosphate, etc., realizing the comprehensive resource utilization of solid waste and having high economic value. By using sulfuric acid and ammonium sulfate for co-leaching, cryolite, which is difficult to dissolve in sulfuric acid, can be converted into ammonium aluminum sulfate with high solubility, realizing the conversion of cryolite in the electrolytic aluminum overhaul slag and improving the aluminum resource recovery rate. Simultaneously, the use of water washing liquid evaporation and concentration to prepare sodium fluoride, followed by freezing to prepare ammonium aluminum sulfate for aluminum resource recovery, is superior to the traditional impurity removal process where fluorides and aluminum form different forms of cryolite and aluminum hydroxide colloids, reducing the impact of lithium loss caused by pH increase during lithium precipitation, the formation of aluminum hydroxide colloids, and the formation of lithium complexes, thus improving the subsequent lithium resource recovery efficiency. Furthermore, currently, the main product recovered from lithium resources in electrolytic aluminum overhaul slag in industry is lithium carbonate, which is then further processed into related battery products. Lithium dihydrogen phosphate (LDH) is superior to lithium carbonate in battery production, primarily due to its lower energy consumption, greater adaptability, lower material usage and overall lower cost, and safer, more stable battery performance with a wider range of applications. Traditional processes involve lithium carbonate precipitation with sodium carbonate, followed by acidification and carbon removal to recover sodium sulfate from the mother liquor, resulting in low lithium recovery. Subsequent reprocessing with lithium carbonate to produce LDH is cumbersome and costly. This invention directly concentrates and separates the lithium sulfate solution to produce both LDH and sodium sulfate, a superior process compared to traditional lithium carbonate production. Furthermore, this method enables the recovery of high-value resources such as fluorine, lithium, and aluminum from electrolytic aluminum overhaul slag, demonstrating significant practicality and economic value.

[0081] In the above embodiments, the descriptions of each embodiment have different focuses. For parts that are not described in detail in a certain embodiment, please refer to the relevant descriptions in other embodiments.

[0082] Obviously, those skilled in the art can make various modifications and variations to this invention without departing from its spirit and scope. Since these modifications and variations fall within the scope of the claims and their equivalents, this invention also intends to include these modifications and variations.

[0083] The above description is merely a specific embodiment of the present invention, but the scope of protection of the present invention is not limited thereto. Any person skilled in the art can easily conceive of various equivalent modifications or substitutions within the technical scope disclosed in the present invention, and these modifications or substitutions should all be covered within the scope of protection of the present invention. Therefore, the scope of protection of the present invention should be determined by the scope of the claims.

Claims

1. A method for resource recovery of electrolytic aluminum overhaul slag, characterized in that, Includes the following steps: S1. Crushing and washing pretreatment: After crushing the slag from the major overhaul of electrolytic aluminum, it is mixed with water, stirred and washed, and then separated into solid and liquid to obtain washing liquid and washing slag. S2. Preparation of sodium fluoride: Acidic gas containing hydrogen fluoride is introduced into the washing solution to adjust the pH of the system to 7-9. Aluminum salt is added to remove impurities to obtain a sodium fluoride solution. After evaporation and crystallization, the sodium fluoride product is obtained. S3, Acid Leaching: The washing residue and acidic liquid are reacted at a solid-liquid mass ratio of 1:3-5 under pH 0-1 conditions. After the reaction is completed, solid-liquid separation is performed to obtain acidic leachate and carbon residue. S4. Aluminum resource recovery: Add an ammonium-containing reagent to the acidic leachate, adjust the pH of the system to 2-4, cool and crystallize, and obtain ammonium aluminum sulfate product after solid-liquid separation; S5. Solution purification and concentration: After the mother liquor from which the aluminum ammonium sulfate product is separated is filtered to remove impurities, the pH of the system is adjusted to 11-13 for alkalization treatment, and then concentrated by evaporation to obtain lithium-rich concentrate and recover the volatilized ammonia. S6. Lithium Resource Recovery: The lithium-rich concentrate is extracted to separate a lithium-containing organic phase and a raffinate; the raffinate is evaporated and crystallized to obtain sodium sulfate product; the lithium-containing organic phase is back-extracted with a phosphoric acid solution to separate a back-extraction solution containing lithium dihydrogen phosphate and a regenerated organic phase; the back-extraction solution is evaporated and crystallized to obtain lithium dihydrogen phosphate product.

2. The method according to claim 1, characterized in that, In step S1, the particle size of the crushed slag is less than or equal to 100 mesh; the solid-liquid mass ratio of the stirring and washing is 1:3-5, and the mixture is stirred at room temperature for 1-2 hours.

3. The method according to claim 1, characterized in that, In step S2, the acidic gas containing hydrogen fluoride includes the gas produced in the reaction in step S3.

4. The method according to claim 1, characterized in that, The aluminum salt is one of aluminum sulfate, polyaluminum chloride, or sodium aluminate.

5. The method according to claim 1, characterized in that, In step S3, the acidic solution is a mixed solution of sulfuric acid and ammonium sulfate, wherein the molar ratio of sulfuric acid to ammonium sulfate is 2:1-4:

1.

6. The method according to claim 1 or 5, characterized in that, In step S3, the reaction temperature is 90-110℃ and the reaction time is 2-4 hours.

7. The method according to claim 1, characterized in that, In step S4, the ammonium-containing reagent is ammonia water; the cooling crystallization temperature is 0-10℃, and the crystallization time is 1-2 hours.

8. The method according to claim 1, characterized in that, The ammonia recovered in step S5 is returned to step S4 for recycling; the regenerated organic phase in step S6 is returned to the extraction process in step S6 for recycling.

9. The method according to claim 1, characterized in that, Step S5 specifically includes: adjusting the pH of the mother liquor after separating the ammonium aluminum sulfate product to 6-8 to precipitate and remove iron and aluminum impurities, and then filtering; adjusting the pH of the filtrate to 11-13, and then evaporating and concentrating it.

10. The method according to claim 1, characterized in that, In step S6, the volume ratio of the organic phase to the lithium-rich concentrate in the extraction process is 3:1-5:1; in the back-extraction, the concentration of the phosphoric acid solution is 5-8 wt%, and the volume ratio of the organic phase to the phosphoric acid solution is 1:5-1:10; when preparing lithium dihydrogen phosphate product, the pH of the evaporation crystallization mother liquor is controlled to be 1.5-2.5.

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