A method for reducing lithium content in lithium-rich electrolyte for aluminum electrolysis and recovering lithium

Abstract

The present invention relates to the technical field of aluminum electrolyte recycling and utilization, and discloses a method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium. The method comprises: uniformly mixing the lithium-rich electrolyte for aluminum electrolysis with an additive and then calcining the mixture to obtain a calcined product A; leaching the calcined product A to obtain a lithium-containing leachate B and a filter residue C; washing and drying the filter residue C, and returning the filter residue C to the aluminum electrolysis cell to continue serving as an electrolyte; purifying the lithium-containing leachate B, removing impurities, and evaporating and concentrating it to obtain a lithium-rich solution; adding a lithium precipitator to the lithium-rich solution to obtain a lithium salt product D. The method provided by the present invention converts the insoluble lithium fluoride in the lithium-rich electrolyte for aluminum electrolysis into readily soluble lithium sulfate, thereby reducing the lithium content in the lithium-rich electrolyte for aluminum electrolysis and achieving the comprehensive utilization of the lithium and fluorine resources in the electrolyte. Furthermore, the present invention has the characteristics of high lithium recovery rate, simple operation, and environmental friendliness.

Description

Technical Field

[0001] The present invention relates to the technical field of aluminum electrolyte recycling, and in particular to a method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium. Background Art

[0002] my country is a major producer of electrolytic aluminum, accounting for approximately 60% of global production capacity. Due to the low grade of bauxite and the large reserves of aluminum and lithium co-existing minerals, a large amount of lithium-containing bauxite is used to produce alumina. Lithium-containing alumina is continuously fed into aluminum electrolysis cells for aluminum production. During this process, lithium accumulates in the aluminum electrolyte due to its low potential and difficulty in precipitation, ultimately forming a lithium-rich electrolyte (lithium content can reach over 1.35%). In aluminum electrolysis molten salt systems, the appropriate addition of lithium fluoride is beneficial for lowering the electrolyte's crystallization temperature, improving conductivity, and current efficiency. However, high lithium content can lead to lower electrolysis temperatures, reduced alumina dissolution capacity, and poor electrode stability. To ensure proper and stable operation of aluminum electrolysis cells, the lithium-rich electrolyte must be partially replaced with an electrolyte with a lower lithium content after a period of operation to control the lithium content in the electrolyte.

[0003] In 2022, my country's electrolytic aluminum production reached 40.214 million tons. It is estimated that every 10,000 tons of electrolytic aluminum produced will generate approximately 100 tons of lithium-rich electrolyte for aluminum electrolysis, a significant amount of which is considered a significant waste. Because this electrolyte contains large amounts of toxic and hazardous substances such as fluoride and cyanide, it is classified as hazardous solid waste and must undergo harmless disposal. At the same time, with the rapid development of new energy vehicles, the demand for lithium salts has surged. Lithium-rich electrolytes, with lithium contents comparable to spodumene and lepidolite, have gradually become a new lithium-containing raw material, attracting significant attention from researchers.

[0004] Chinese invention patent publication number CN109179457B, "A method for extracting lithium from waste aluminum electrolytic residues," discloses a method for reacting concentrated sulfuric acid with waste aluminum electrolytic residues to produce a lithium-containing solution. This solution is then neutralized with calcium oxide to remove excess sulfuric acid, followed by evaporation and concentration, and lithium carbonate is produced. Chinese invention patent publication number CN105293536A, "A method for extracting lithium from waste aluminum electrolytic residues," similarly reacts waste aluminum electrolytic residues with concentrated sulfuric acid at 200-400°C to transfer lithium from the waste aluminum electrolytic residues into a solution for recovery. Chinese invention patent publication number CN116334410A, "A method for separating lithium from waste aluminum electrolytic residues containing lithium electrolytes," utilizes concentrated sulfuric acid, a reaction aid, and deionized water to ball-mill a mixture of the electrolyte containing lithium and then calcines it. The mixture is then water-soaked to produce a lithium-containing solution, thereby recovering lithium from the electrolyte. The above patents can all realize the recovery of lithium in lithium-rich electrolyte of aluminum electrolysis. However, since the lithium-rich electrolyte contains a large amount of fluorine, the concentrated sulfuric acid method is likely to produce highly corrosive hydrofluoric acid when treating the lithium-rich electrolyte of aluminum electrolysis, which places extremely high requirements on the corrosion resistance of the equipment. In addition, the above patents fail to fully utilize the fluorine resources in the lithium-rich electrolyte.

[0005] Chinese invention patent publication number CN114438329A, "A Comprehensive Recovery Method for Wasted Lithium-Containing Aluminum Electrolytes," uses a leaching enhancer to react with a lithium-rich electrolyte in an acidic solution, transferring lithium from the electrolyte into the leachate for recovery. However, this method also suffers from the release of hydrogen fluoride under acidic conditions. Chinese invention patent publication number CN114890447A, "A Method for Directly Preparing Aluminum Fluoride from Aluminum Electrolyte without Calcination," high-energy ball mills aluminum salt additives with aluminum electrolytes. The resulting ball-milled material is then leached in water to produce a lithium-containing solution and leached residue. While this patent can achieve lithium recovery, it places stringent requirements on the molar ratio of sodium fluoride to aluminum fluoride in the aluminum electrolyte. The Chinese invention patent with publication number CN116768246A, "A method for efficient separation of lithium and aluminum and enrichment of lithium from aluminum electrolyte waste residue," mixes aluminum electrolyte with aluminum sulfate, then undergoes a low-temperature roasting step and a high-temperature thermal decomposition step. The resulting roasted material is then leached in water to obtain a lithium sulfate solution, thereby recovering lithium from the residue. However, this method has high roasting temperatures, high energy consumption, and produces sulfur-containing flue gas that requires additional treatment.

[0006] In summary, the existing technology for treating lithium-rich electrolytes in aluminum electrolysis is prone to produce highly corrosive hydrogen fluoride, which places high demands on the corrosion resistance of the equipment and fails to fully utilize the valuable resources in the electrolyte (such as fluorine resources), while fluoride is the main component of the aluminum electrolysis molten salt system. Therefore, the development of new technologies to reduce the lithium content in lithium-rich electrolytes in aluminum electrolysis and comprehensively recover valuable resources (lithium, fluorine, aluminum, etc.) in the electrolyte is of great significance for realizing the comprehensive utilization of aluminum electrolysis by-products, promoting clean production in the aluminum electrolysis industry, and reducing my country's dependence on foreign lithium resources. Summary of the Invention

[0007] In view of the deficiencies of the above-mentioned prior art, the present invention aims to provide a method for reducing the lithium content in the lithium-rich electrolyte of aluminum electrolysis and recovering lithium. Alum-based double salt is uniformly mixed with the lithium-rich electrolyte of aluminum electrolysis and then roasted, so that the insoluble lithium fluoride and lithium cryolite in the lithium-rich electrolyte of aluminum electrolysis are converted into lithium sulfate that is easily soluble in water, and fluorine is converted into aluminum fluoride that is insoluble in water, thereby achieving effective separation of lithium and elements such as aluminum and fluorine in the lithium-rich electrolyte of aluminum electrolysis. In addition, the filter residue obtained after water leaching of the roasted product mainly contains aluminum fluoride, aluminum oxide, etc., which can be returned to the aluminum electrolysis cell for continued use after being fully washed and dried, thereby realizing the comprehensive utilization of aluminum and fluorine resources. The lithium-containing leachate is purified, impurities removed, evaporated and concentrated, and then used for lithium salt products, thereby achieving efficient recovery of lithium.

[0008] To achieve the above object, the present invention is implemented according to the following technical solutions:

[0009] A method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium comprises the following steps:

[0010] S1, mixing the aluminum electrolysis lithium-rich electrolyte and the additive uniformly and then calcining to obtain a calcined product A;

[0011] S2, leaching the calcined product A to obtain a lithium-containing leachate B and a filter residue C; the filter residue C is washed and dried and then returned to the aluminum electrolysis cell to continue serving as an electrolyte;

[0012] S3, the lithium-containing leachate B is purified and impurities removed, evaporated and concentrated to obtain a lithium-rich solution, and a lithium precipitating agent is added to the lithium-rich solution to obtain a lithium salt product D.

[0013] The lithium-rich electrolyte for aluminum electrolysis comes from the lithium-containing electrolytic slag produced during the aluminum electrolysis process, especially the electrolytic slag produced by lithium-containing alumina electrolysis in northern my country. In order to accelerate the reaction rate and make the reaction more complete, the lithium-rich electrolyte needs to be crushed and screened.

[0014] Preferably, in step S1, the particle size of the lithium-rich electrolyte for aluminum electrolysis is 75-500 mesh.

[0015] Further preferably, in step S1, the particle size of the lithium-rich electrolyte for aluminum electrolysis is 100-500 mesh.

[0016] In order to convert the lithium fluoride that is insoluble in water in the lithium-rich electrolyte of aluminum electrolysis into a lithium salt that is easily soluble in water, and thus realize the extraction of lithium from the lithium-rich electrolyte of aluminum electrolysis. The present invention uses alum-based double salts as additives to carry out a high-temperature solid-phase reaction with the lithium-rich electrolyte of aluminum electrolysis, so that the lithium fluoride is converted into aluminum sulfate that is soluble in water, and the fluorine in the lithium-rich electrolyte of aluminum electrolysis combines with the aluminum in the alum-based double salts to generate aluminum fluoride that is more thermodynamically stable and insoluble in water, thereby realizing the separation of lithium from elements such as aluminum and fluorine. At the same time, during the high-temperature solid-phase reaction, the sodium, potassium, ammonium and other ions in the alum-based double salts easily react with the lithium ions in the lithium-rich electrolyte by isomorphism, such as sodium and lithium isomorphism to generate sodium fluoride. The reaction of sodium fluoride with the aluminum in the alum-based double salts has a larger equilibrium constant, that is, the reaction can occur more easily, thereby realizing the efficient extraction of lithium. In addition, alum-based double salts are mainly sulfate double salts of aluminum and alkali metals or ammonium. After being used as additives and roasted with lithium-rich electrolytes for aluminum electrolysis and then immersed in water, alkali metal or ammonium ions enter the aqueous solution, and only aluminum exists in the slag in the form of fluoride, that is, the leached slag returned to the electrolytic cell will not introduce other impurity ions.

[0017] Preferably, in step S1, the additive is an alum-based double salt.

[0018] Further preferably, in step S1, the alum-based double salt is a sulfate double salt of aluminum and an alkali metal or ammonium.

[0019] Preferably, in step S1, the alum-based double salt is at least one of potassium alum, ammonium alum and sodium alum.

[0020] To fully convert the insoluble lithium salts in the lithium-rich electrolyte used in aluminum electrolysis into water-soluble lithium salts, the amount of additives added must be controlled. Too little additives will result in incomplete conversion of insoluble components like lithium fluoride. Too much additives will result in a high impurity content in the leaching solution obtained after calcination, adding a burden to the subsequent lithium carbonate preparation process.

[0021] Preferably, in step S1, the ratio of the lithium-rich electrolyte to the additive in aluminum electrolysis is 1 to 3:1 in molar ratio based on the lithium content in the lithium-rich electrolyte and the aluminum content in the additive.

[0022] Since the original aluminum fluoride, cryolite and aluminum fluoride generated by subsequent reactions in the lithium-rich electrolyte of aluminum electrolysis are easy to react with water vapor at high temperature to generate aluminum oxide and hydrogen fluoride gas, and alum-based complex salts generally contain crystal water. In order to ensure that the fluorine in the lithium-rich electrolyte is recovered in the form of aluminum fluoride, rather than generating hydrogen fluoride gas and existing in the tail gas, it is necessary to adopt a certain temperature increase program during the roasting process to fully remove the crystal water in the alum-based complex salt. Therefore, the present invention adopts a two-stage temperature increase program, that is, first keeping warm at low temperature for a period of time to fully remove moisture, and then reacting at high temperature to realize the conversion of insoluble lithium salts into water-soluble lithium salts; at the same time, the high-temperature reaction stage should avoid using too high a reaction temperature to prevent the generation of sulfur-containing flue gas.

[0023] Preferably, in step S1, the roasting process is to keep the temperature at 200-300°C for 1-2 hours, and then raise the temperature to 400-750°C and keep the temperature for 0.5-12 hours.

[0024] Specifically, the process of step S1 is as follows: the crushed and screened aluminum electrolysis lithium-rich electrolyte and the additive are evenly mixed in a certain proportion, and the calcined product A is obtained after the calcined reaction according to a set temperature program in an air atmosphere; the aluminum electrolysis lithium-rich electrolyte is sieved through a 75-500 mesh sieve; the additive is an alum-based double salt, preferably at least one of potassium alum, ammonium alum, sodium alum, etc.; the ratio of the aluminum electrolysis lithium-rich electrolyte to the additive is 1:1 to 3:1 according to the molar ratio of the lithium content in the lithium-rich electrolyte to the aluminum content in the additive; the calcination reaction temperature program is first kept at 200-300°C for 1-2 hours for low-temperature dehydration, and then continued to heat to 400-750°C and kept for 0.5-12 hours for solid-phase reaction.

[0025] The main components of the calcined product A obtained after the high-temperature solid-phase calcination reaction of the lithium-rich electrolyte and additives in aluminum electrolysis are lithium sulfate, aluminum fluoride, aluminum oxide, etc. The different solubilities of these products in water or dilute acid can be used to effectively separate lithium from other components; the filter residue obtained after leaching basically does not contain lithium-containing aluminum fluoride, aluminum oxide, etc., which are all the main components of the aluminum electrolysis molten salt system. After being fully washed with deionized water and dried, it can be returned to the aluminum electrolysis cell to continue to be used as an electrolyte.

[0026] Preferably, in step S2, deionized water or acid is used in the leaching process; the acid is one of dilute sulfuric acid, dilute hydrochloric acid, and dilute nitric acid.

[0027] Specifically, the liquid-to-solid ratio of the leaching process is 2 to 5:1.

[0028] Preferably, in step S2, the concentration of the acid is 0.1 to 1 mol / L.

[0029] In order to fully transfer the lithium in the roasted product A into the solution and facilitate subsequent recovery, the reaction conditions during the leaching process need to be controlled.

[0030] Preferably, in step S2, the leaching temperature is 20-95° C., and the leaching time is 0.5-6 h.

[0031] Specifically, the process of step S2 is as follows: the roasted product A is leached with deionized water or dilute acid to obtain a lithium-containing leachate B and a filter residue C; the filter residue C is fully washed with deionized water, dried, and returned to the aluminum electrolysis cell to continue to serve as an electrolyte; the dilute acid is one of dilute sulfuric acid, dilute hydrochloric acid, and dilute nitric acid, and the acid concentration is 0.1 to 1 mol / L; the leaching temperature during the leaching process is 20 to 95° C., and the leaching time is 0.5 to 6 hours.

[0032] Since the lithium-containing leachate B contains a small amount of impurity ions such as aluminum, these impurity ions are easily co-precipitated with lithium when the leachate recovers lithium and enters the lithium salt product. Therefore, the lithium-containing leachate B needs to be purified and impurities removed. Since aluminum hydroxide is a colloid that is difficult to filter and easily adsorbs lithium ions, resulting in lithium loss, it is impossible to remove aluminum by a step-by-step precipitation method. The present invention utilizes the characteristic of aluminum hydroxide being an amphoteric hydroxide, that is, under alkaline conditions, aluminum will exist in the form of aluminate ions, and aluminate will not react with carbonate, phosphate, etc. in the subsequent lithium precipitation process, thereby achieving the separation of lithium from impurity ions such as aluminum in the leachate.

[0033] Preferably, in step S3, the purification and impurity removal process is to add an alkaline solution to the lithium-containing leachate B to adjust its pH value to be greater than 10.

[0034] Preferably, the alkaline solution is at least one of a sodium hydroxide solution and a potassium hydroxide solution.

[0035] Because lithium salts such as lithium carbonate and lithium phosphate are slightly soluble in water, a high lithium concentration must be maintained during the lithium precipitation process to improve lithium recovery. Therefore, the lithium-containing leachate must be properly evaporated and concentrated to obtain a lithium-rich solution with a high lithium concentration.

[0036] Preferably, in step S3, the lithium concentration of the lithium-rich solution is greater than 10 g / L.

[0037] In order to recover the lithium in the solution as lithium salt, a lithium precipitating agent is required to precipitate lithium from the lithium-rich solution. The lithium precipitating agent used must be a lithium salt that forms a poorly soluble lithium salt with lithium ions.

[0038] Preferably, in step S3, the lithium precipitation agent is at least one of soluble fluoride, carbonate, and phosphate.

[0039] Specifically, the process of step S3 is as follows: the lithium-containing leachate B is purified and impurities removed, and then evaporated and concentrated to obtain a lithium-rich solution, and then a lithium precipitating agent is added to the lithium-rich solution to obtain a lithium salt product D. The purification and impurity removal method is to add an alkaline solution such as sodium hydroxide or potassium hydroxide to the lithium-containing leachate B to adjust its pH to greater than 10; the lithium-rich solution obtained by evaporation and concentration has a lithium concentration greater than 10g / L; the lithium precipitating agent can be at least one of a soluble fluoride, a carbonate, and a phosphate.

[0040] The main chemical reactions involved in the method for reducing the lithium content in the lithium-rich electrolyte of aluminum electrolysis and recovering lithium provided by the present invention are as follows (taking potassium alum as an example):

[0041] 6LiF+2KAl(SO4)2·12H2O=3Li2SO4+K2SO4+2AlF3+24H2O;

[0042] 2Li3AlF6+2KAl(SO4)2·12H2O=3Li2SO4+K2SO4+4AlF3+24H2O.

[0043] Beneficial effects:

[0044] 1) Alum-based double salts are used to react with lithium-rich electrolytes for aluminum electrolysis, so that lithium fluoride, lithium cryolite, etc., which are insoluble in water, are converted into lithium sulfate, which is easily soluble in water. Fluorine is converted into aluminum fluoride, one of the important components in the aluminum electrolysis molten salt system. This reduces the lithium content in the lithium-rich electrolyte for aluminum electrolysis and realizes the comprehensive utilization of lithium and fluorine resources in the lithium-rich electrolyte for aluminum electrolysis.

[0045] 2) Alum-based double salts are mainly sulfates of alkali metals and aluminum. The presence of alkali metals can react isomorphously with lithium, improving the conversion efficiency of lithium. After the obtained product is leached with water, the alkali metal ions are transferred to the solution. No other impurity ions are introduced during the lithium recovery process. The obtained leaching residue mainly consists of aluminum fluoride and aluminum oxide, etc., which can be returned to the aluminum electrolysis cell for further use.

[0046] 3) A two-stage heating process of low-temperature dehydration followed by high-temperature calcination is used to ensure that the fluorine in the lithium-rich electrolyte does not generate toxic hydrogen fluoride gas during the calcination reaction, but instead reacts with the aluminum in the alum-based double salt to generate aluminum fluoride, thereby maximizing the recovery of fluorine and aluminum resources.

[0047] 4) The method provided by the present invention has the advantages of high recovery rate, simple operation, environmental friendliness and low maintenance cost, and has important practical significance for the treatment of lithium-rich electrolytes in aluminum electrolysis and the recycling of valuable resources. BRIEF DESCRIPTION OF THE DRAWINGS

[0048] Figure 1 is a process flow chart of the present invention;

[0049] Figure 2 The figure is a comparison chart of the lithium recovery rates of Example 1, Example 2, Example 3, Example 4, Example 5 and Comparative Example 1 and Comparative Example 2 of the present invention. DETAILED DESCRIPTION

[0050] The present invention will be further described in detail below with reference to specific embodiments. It should be understood that the specific embodiments described herein are only used to explain the present invention and are not intended to limit the present invention.

[0051] There is no particular limitation on the purity of all raw materials in the present invention. The present invention preferably uses industrial purity or conventional purity used in the art.

[0052] The devices used in the present invention are not particularly limited and can be devices commonly used in the art.

[0053] Table 1 Main components of lithium-rich electrolyte for aluminum electrolysis

[0054] Element Li Al F Na Ca K content,% 2.17 18.96 50.03 13.26 2.31 1.98

[0055] Example 1

[0056] This embodiment provides a method for reducing the lithium content in lithium-rich electrolytes for aluminum electrolysis and recovering lithium. The principle process flow chart is as follows: Figure 1 As shown, potassium alum is used as an additive and deionized water is used for leaching, which includes the following steps:

[0057] (1) 100 g of aluminum electrolysis lithium-rich electrolyte (main components shown in Table 1) obtained by screening with a 100-mesh sieve was mixed evenly with 147.03 g of an additive (potassium alum) in a molar ratio of lithium content in the lithium-rich electrolyte to aluminum content in the additive of 1:1, and then heated at 300°C for 1 hour for dehydration, and then reacted at 750°C for 0.5 hour to obtain 180.14 g of calcined product A;

[0058] (2) The calcined product A was leached with 500 mL of deionized water at 90° C. for 0.5 h to obtain 500 mL of lithium-containing leachate B with a lithium concentration of 4.28 g / L and 106.20 g of filter residue C. The filter residue C was fully washed with 5 L of deionized water, dried, and returned to the aluminum electrolysis cell to continue to serve as a molten salt electrolyte in the aluminum electrolysis process;

[0059] (3) The lithium-containing leachate B was purified and impurities removed by adding sodium hydroxide solution to adjust the pH to 11.32. The pH-adjusted solution was then evaporated and concentrated to obtain 203 mL of a lithium-rich solution having a lithium concentration of 10.56 g / L. Trisodium phosphate was then added to the lithium-rich solution to obtain 11.71 g of lithium phosphate product D.

[0060] The lithium content in the filter residue C obtained after washing and drying was analyzed. Compared with the lithium content of 2.17% in the initial lithium-rich electrolyte, the lithium content in the filter residue C dropped to 0.03%, effectively reducing the lithium content in the electrolyte. It was calculated that the lithium recovery rate in the lithium-rich electrolyte was as high as 98.53%.

[0061] Example 2

[0062] This embodiment provides a method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium, specifically using soda alum as an additive and leaching with dilute hydrochloric acid, comprising the following steps:

[0063] (1) 500 g of aluminum electrolysis lithium-rich electrolyte (its main components are shown in Table 1) obtained by screening through a 500-mesh sieve was mixed evenly with 236.63 g of an additive (sodium alum) at a molar ratio of lithium content in the lithium-rich electrolyte to aluminum content in the additive of 3:1, and then heated at 250°C for 2 h for dehydration, and then reacted at 600°C for 6 h to obtain 875.09 g of calcined product A;

[0064] (2) The calcined product A was leached with 2000 mL of 1 mol / L hydrochloric acid at 20°C for 3 h to obtain a lithium-containing leachate B with a lithium concentration of 5.23 g / L and 529.87 g of a filter residue C. The filter residue C was fully washed with 10 L of deionized water, dried, and returned to the aluminum electrolysis cell to continue to serve as a molten salt electrolyte in the aluminum electrolysis process;

[0065] (3) The lithium-containing leachate B was purified and impurities removed by adding sodium hydroxide solution to adjust the pH to 12.49. The pH-adjusted solution was then evaporated and concentrated to obtain 490 mL of a lithium-rich solution having a lithium concentration of 21.35 g / L. Sodium carbonate was then added to the lithium-rich solution to obtain 51.09 g of lithium phosphate product D.

[0066] The lithium content in the filter residue C obtained after washing and drying was analyzed. Compared with the lithium content of 2.17% in the initial lithium-rich electrolyte, the lithium content in the filter residue C dropped to 0.08%, effectively reducing the lithium content in the electrolyte. It was calculated that the lithium recovery rate in the lithium-rich electrolyte was as high as 96.08%.

[0067] Example 3

[0068] This embodiment provides a method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium, specifically using ammonium alum as an additive and leaching with dilute sulfuric acid, including the following steps:

[0069] (1) 200 g of aluminum electrolysis lithium-rich electrolyte (its main components are shown in Table 1) obtained by screening with a 75-mesh sieve was mixed evenly with 181.87 g of an additive (ammonium alum) at a molar ratio of lithium content in the lithium-rich electrolyte to aluminum content in the additive of 2:1, and then heated at 200°C for 1.5 h for dehydration, and then reacted at 400°C for 12 h to obtain 292.59 g of calcined product A;

[0070] (2) The calcined product A was leached with 1000 mL of a 0.1 mol / L dilute sulfuric acid solution at 50°C for 6 h to obtain a lithium-containing leachate B with a lithium concentration of 4.24 g / L and 212.38 g of a filter residue C. The filter residue C was fully washed with 8 L of deionized water, dried, and returned to the aluminum electrolysis cell to continue serving as a molten salt electrolyte in the aluminum electrolysis process;

[0071] (3) The lithium-containing leachate B was purified and impurities removed by adding sodium hydroxide solution to adjust the pH to 10.49. The pH-adjusted solution was then evaporated and concentrated to obtain 250 mL of a lithium-rich solution having a lithium concentration of 16.93 g / L. Sodium fluoride was then added to the lithium-rich solution to obtain 15.09 g of lithium phosphate product D.

[0072] The lithium content in the filter residue C obtained after washing and drying was analyzed. Compared with the lithium content of 2.17% in the initial lithium-rich electrolyte, the lithium content in the filter residue C dropped to 0.05%, effectively reducing the lithium content in the electrolyte. It was calculated that the lithium recovery rate in the lithium-rich electrolyte was as high as 97.55%.

[0073] Example 4

[0074] This embodiment provides a method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium, specifically using potassium alum as an additive and leaching with dilute nitric acid, including the following steps:

[0075] (1) 300 g of aluminum electrolysis lithium-rich electrolyte (its main components are shown in Table 1) obtained by screening through a 200-mesh sieve was mixed evenly with 147.09 g of an additive (potassium alum) at a molar ratio of lithium content in the lithium-rich electrolyte to aluminum content in the additive of 3:1, and then heated at 200°C for 1.5 h for dehydration, and then reacted at 400°C for 12 h to obtain 380.14 g of calcined product A;

[0076] (2) The calcined product A was leached with 750 mL of 0.5 mol / L dilute nitric acid at 30° C. for 2 h to obtain a lithium-containing leachate B with a lithium concentration of 8.32 g / L and 301.86 g of a filter residue C. The filter residue C was fully washed with 6 L of deionized water, dried, and returned to the aluminum electrolysis cell to continue to serve as a molten salt electrolyte in the aluminum electrolysis process;

[0077] (3) The lithium-containing leachate B was purified by adding potassium hydroxide solution to adjust the pH to 11.19. The pH-adjusted solution was then evaporated and concentrated to obtain 280 mL of a lithium-rich solution having a lithium concentration of 22.29 g / L. Sodium carbonate was then added to the lithium-rich solution to obtain 30.62 g of lithium phosphate product D.

[0078] The lithium content in the filter residue C obtained after washing and drying was analyzed. Compared with the lithium content of 2.17% in the initial lithium-rich electrolyte, the lithium content in the filter residue C dropped to 0.09%, effectively reducing the lithium content in the electrolyte. It was calculated that the lithium recovery rate in the lithium-rich electrolyte was as high as 95.60%.

[0079] Example 5

[0080] This embodiment provides a method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium, specifically using soda alum as an additive and leaching with deionized water, including the following steps:

[0081] (1) 1000 g of aluminum electrolysis lithium-rich electrolyte (its main components are shown in Table 1) obtained by screening with a 200-mesh sieve was mixed evenly with 473.27 g of an additive (sodium alum) at a molar ratio of lithium content in the lithium-rich electrolyte to aluminum content in the additive of 3:1, and then heated at 300°C for 1 hour for dehydration, and then reacted at 550°C for 8 hours to obtain 1250.07 g of calcined product A;

[0082] (2) The calcined product A was leached with 2500 mL of deionized water at 60° C. for 1 h to obtain a lithium-containing leachate B with a lithium concentration of 8.28 g / L and 1006.20 g of a filter residue C. The filter residue C was fully washed with 15 L of deionized water, dried, and returned to the aluminum electrolysis cell to continue serving as a molten salt electrolyte in the aluminum electrolysis process;

[0083] (3) The lithium-containing leachate B was purified and impurities removed by adding sodium hydroxide solution to adjust the pH to 12.09. The pH-adjusted solution was then evaporated and concentrated to obtain 815 mL of a lithium-rich solution having a lithium concentration of 25.41 g / L. Sodium carbonate was then added to the lithium-rich solution to obtain 101.66 g of lithium phosphate product D.

[0084] The lithium content in the filter residue C obtained after washing and drying was analyzed. Compared with the lithium content of 2.17% in the initial lithium-rich electrolyte, the lithium content in the filter residue C dropped to 0.10%, effectively reducing the lithium content in the electrolyte. It was calculated that the lithium recovery rate in the lithium-rich electrolyte was as high as 95.36%.

[0085] Comparative Example 1

[0086] This comparative example differs from Example 1 in that no alum-based additive is added to the aluminum electrolytic lithium-rich electrolyte. Instead, 100 g of the aluminum electrolytic lithium-rich electrolyte is directly calcined. The calcination and leaching conditions are the same as those in Example 1. The product obtained in this comparative example has a lithium concentration in the leachate of only 0.09 g / L, a lithium content in the filter residue C of 2.13%, and a lithium recovery rate of only 2.12% in the lithium-rich electrolyte.

[0087] Comparative Example 2

[0088] This comparative example differs from Example 1 in that a mixture of 100 g of the lithium-rich electrolyte and potassium alum was calcined at 340°C for 12 hours. All other conditions were the same as in Example 1. The product obtained in this comparative example had a lithium concentration of only 0.25 g / L in the leachate, a lithium content of 2.05% in the filter residue C, and a lithium recovery rate of only 5.76% in the lithium-rich electrolyte.

[0089] The lithium recovery rates of Example 1, Example 2, Example 3, Example 4, Example 5 and Comparative Example 1 and Comparative Example 2 are plotted, and the results are as follows. Figure 2 As shown in the figure, it can be seen that the excellent effect of the method for reducing the lithium content in the lithium-rich electrolyte of aluminum electrolysis and recovering lithium provided by the present invention depends on both the addition of additives and the reasonable control of the roasting temperature, which is the result of a comprehensive effect.

[0090] Comparative Example 3

[0091] The difference between this comparative example and Example 1 is that the mixture of 100g of lithium-rich electrolyte and potassium alum does not adopt the heating procedure of low-temperature dehydration and high-temperature calcination. Instead, the mixture is directly heated to 750°C and calcined for 0.5h. Other conditions are the same as those in Example 1. Analysis of the product obtained in this comparative example showed that the lithium concentration in the leachate was 4.26g / L, the lithium content in the filter residue C was 0.04%, and the lithium recovery rate in the lithium-rich electrolyte was only 98.34%. However, the obtained calcined product was only 159.08g, and the mass of the obtained filter residue C was only 85.14g, indicating that a portion of the fluorine in the lithium-rich electrolyte reacted with water vapor at high temperature to generate hydrogen fluoride and was lost in the tail gas, and could not be recovered in the form of aluminum fluoride for aluminum electrolysis.

[0092] Comparative Example 4

[0093] This comparative example differs from Example 2 in that aluminum sulfate (34.43 g, with a molar ratio of lithium content in the lithium-rich electrolyte to aluminum content in the additive of 3:1) was used as an additive and then mixed with the lithium-rich electrolyte before calcination. Other conditions were the same as in Example 2. The product obtained in this comparative example had a lithium concentration in the leachate of 3.41 g / L and a lithium content in the filter residue C of 0.47%, which was significantly higher than the lithium content of 0.08% in the residue in Example 2. However, the recovery rate of lithium in the lithium-rich electrolyte was only 78.57%.

[0094] By comparing Example 2 and Comparative Example 4, it can be clearly found that when an alum-based double salt is used as an additive and mixed with a lithium-rich electrolyte for calcination, the alkali metal or ammonium ion in the alum-based double salt can react isomorphously with the lithium in the lithium-rich electrolyte during the calcination process, which is beneficial to the conversion of lithium fluoride, thereby effectively reducing the lithium content in the slag and achieving deep extraction of lithium from the lithium-rich electrolyte.

[0095] The technical solution of the present invention is not limited to the above-mentioned specific embodiments. Any technical variations made according to the technical solution of the present invention fall within the protection scope of the present invention.

Claims

1. A method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium, characterized by: The following steps are involved: S1, mixing the aluminum electrolysis lithium-rich electrolyte and the additive uniformly and then calcining to obtain a calcined product A; S2, leaching the calcined product A to obtain a lithium-containing leachate B and a filter residue C; the filter residue C is washed and dried and then returned to the aluminum electrolysis cell to continue serving as an electrolyte; S3, the lithium-containing leachate B is purified and concentrated by evaporation to obtain a lithium-rich solution, and a lithium precipitation agent is added to the lithium-rich solution to obtain a lithium salt product D; In step S1, the additive is an alum-based double salt; In step S1, the roasting process is to keep the temperature at 200-300° C. for 1-2 hours, and then raise the temperature to 400-750° C. and keep the temperature for 0.5-12 hours.

2. The method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium according to claim 1, characterized in that: In step S1, the alum-based double salt is at least one of potassium alum, ammonium alum, and sodium alum.

3. The method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium according to claim 1, characterized in that: In the step S1, the ratio of the lithium-rich electrolyte to the additive in aluminum electrolysis is 1 to 3:1 in molar ratio based on the lithium content in the lithium-rich electrolyte and the aluminum content in the additive.

4. The method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium according to claim 1, characterized in that: In step S2, deionized water or acid is used in the leaching process; the acid is one of dilute sulfuric acid, dilute hydrochloric acid, and dilute nitric acid.

5. The method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium according to claim 4, characterized in that: In step S2, the concentration of the acid is 0.1 to 1 mol / L.

6. The method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium according to claim 1, characterized in that: In step S2, the leaching temperature is 20-95° C., and the leaching time is 0.5-6 hours.

7. The method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium according to claim 1, characterized in that: In step S3, the purification and impurity removal process is to add an alkaline solution to the lithium-containing leachate B to adjust its pH value to be greater than 10.

8. The method for reducing the lithium content in a lithium-rich electrolyte for aluminum electrolysis and recovering lithium according to claim 1, characterized in that: In step S3, the lithium precipitation agent is at least one of soluble fluoride, carbonate, and phosphate.

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