Method for efficiently leaching waste aluminum electrolyte and recovering fluorine, lithium, aluminum and sodium

Abstract

The invention discloses a method for efficiently leaching waste aluminum electrolyte and recovering fluorine, lithium, aluminum and sodium, belongs to the field of solid waste resource recovery, and solves the following problems of low aluminum and lithium leaching rate in an alkaline process and HF pollution in an acid process. The method comprises the following steps: crushing, grinding and screening waste aluminum electrolyte blocks to obtain waste aluminum electrolyte powder A; uniformly mixing the waste aluminum electrolyte powder A with a calcium-containing reaction auxiliary material to obtain a material B; adding the material B into an alkali solution, and carrying out a leaching reaction; separating the materials to obtain filter residues and filtrate; carrying out acid pickling purification and impurity removal on filter residues, and filtering; extracting the filtrate to obtain a loaded organic phase and raffinate; carrying out reverse extraction on the loaded organic phase; adding reaction auxiliary materials into the raffinate, uniformly stirring, reacting, standing, filtering and separating; continuously supplementing reaction auxiliary materials into the filtrate, uniformly stirring, standing, filtering and separating. According to the method, multiple components in the waste aluminum electrolyte can be efficiently recycled, and the comprehensive utilization rate is high.

Description

Technical Field

[0001] The present invention relates to the technical field of solid waste resource recovery, and in particular to a method for efficiently leaching and recovering fluorine, lithium, aluminum and sodium from lithium-containing waste aluminum electrolyte. Background Art

[0002] During the Bayer process, lithium eventually enters the alumina product. Lithium-containing alumina is used as a raw material in the production process of electrolytic aluminum. Since lithium is more electrochemically active than aluminum, lithium accumulates in the electrolyte, causing great changes in the electrolyte composition and physical and chemical properties. Excessive lithium salt content in the aluminum electrolyte will cause the initial crystallization temperature of the electrolyte system to be too low, resulting in a decrease in the temperature of the electrolytic cell, making the production of the aluminum electrolytic cell unstable and the technical conditions difficult to maintain, which increases the difficulty of electrolytic process operation and is very unfavorable to production stability and energy consumption. At present, the lithium fluoride content of most companies has exceeded 3%, and the highest has reached more than 7%. When the lithium content exceeds the standard, a part of the new electrolyte needs to be replaced, which will produce a certain amount of waste aluminum electrolyte containing more than 2% lithium.

[0003] As an electrode material for high-energy-density batteries, lithium has irreplaceable and important applications in the fields of new energy and electronics. my country is rich in lithium resources, but they mainly exist in the form of salt lake brine, spodumene and lithium mica. The development and utilization of lithium resources face difficulties including uneven distribution, low grade, low recovery rate, and great pressure on environmental protection. In addition, fluorite is also a scarce non-renewable resource. Among them, acid-grade fluorite concentrate and high-grade fluorite are in great demand and in short supply. However, the current mining of fluorite resources in my country has the problem of "low reserves and high output".

[0004] The main components of waste aluminum electrolyte are cryolite and lithium-containing cryolite, and the main components with recycling value include lithium, fluorine, and aluminum. The chemical composition and phase composition of waste aluminum electrolyte are relatively simple. Compared with mineral resources, the process of recovering lithium and fluorine from waste aluminum electrolyte is relatively simple and has important value. At present, the treatment process of waste aluminum electrolyte mainly includes alkaline / acid leaching, leaching after roasting, etc. CN112919507B provides a method for extracting lithium salts from electrolytes, using multiple alkaline leaching, acid leaching, and lithium carbonate precipitation to obtain high-grade cryolite and lithium carbonate; CN117430144A provides a green acid leaching method for resource utilization of waste aluminum electrolytes, by adding sodium salt to medium-concentration acid solution for leaching, lithium extraction and recovery is achieved; CN114890447B provides a method for directly preparing aluminum fluoride without roasting using aluminum electrolyte as raw material, by adding aluminum salt for high-energy ball milling, washing and filtering to obtain aluminum fluoride. In general, in the existing technical routes, the treatment is mainly carried out through alkaline leaching, acid leaching (or acid leaching under the condition of adding aluminum salt or sodium salt), acid leaching after roasting and other processes, but most of them are mainly for lithium extraction. The recovery of fluorine is prone to overflow of harmful and highly corrosive HF gas during the acid leaching process, which not only pollutes the environment, but also causes a waste of fluorine resources. In the alkaline extraction process, due to the dissolution of cryolite and the slightly soluble nature of LiF, the aluminum-lithium leaching rate is generally not high, and cryolite phase and aluminum fluoride exist simultaneously in the leached residue, which is difficult to recycle. Summary of the invention

[0005] In view of the above situation, the present invention aims to provide a method for efficiently leaching and recovering fluorine, lithium, aluminum and sodium from waste aluminum electrolyte, so as to solve one of the following technical problems: low aluminum and lithium leaching rate in alkaline process, HF pollution and fluorine loss in acid process.

[0006] The purpose of the present invention is mainly achieved through the following technical solutions:

[0007] The present invention provides a method for efficiently leaching and recovering fluorine, lithium, aluminum and sodium from waste aluminum electrolyte, the method comprising the following steps:

[0008] Step 1, crushing, grinding and screening the waste aluminum electrolyte blocks to obtain waste aluminum electrolyte powder A;

[0009] Step 2, mixing the waste aluminum electrolyte powder A and the calcium-containing reaction auxiliary material evenly to obtain material B;

[0010] Step 3, adding material B to the alkaline solution, stirring evenly, and adding it to the alkali-resistant reaction device for leaching reaction;

[0011] Step 4: After the reaction is completed, the material is filtered and separated to obtain filter residue and filtrate respectively;

[0012] Step 5a, using an inorganic acid solution to acid-wash the filter residue obtained in step 4, purify and remove impurities, filter, and dry the filter cake to obtain CaF2;

[0013] Step 5b, extracting the filtrate obtained in step 4 with an organic solvent extractant to obtain a loaded organic phase and a raffinate;

[0014] Step 6a, stripping the loaded organic phase obtained in step 5b to obtain a lithium-containing solution;

[0015] Step 6b, adding reaction auxiliary materials to the raffinate obtained in step 5b, stirring and reacting evenly and standing for a period of time, filtering and separating to obtain calcium aluminate precipitate, and obtaining a high-purity calcium aluminate product after acid washing;

[0016] Step 7: Continue to add reaction auxiliary materials to the filtrate obtained in step 6b, stir evenly, let stand, and filter and separate.

[0017] Furthermore, in step 6b, the reaction auxiliary material is at least one of CaO and Ca(OH)2.

[0018] Furthermore, in step 6b, the reaction temperature is 20-80°C.

[0019] Furthermore, in step 6b, the stirring speed is 200 to 500 rpm, and the reaction time is 0.5 to 2 hours.

[0020] Furthermore, in step 6b, the standing time is 0.5 to 1 hour.

[0021] Furthermore, in step 6b, the pH value of the solution during the pickling process is in the range of 2 to 5.

[0022] Furthermore, in step 2, the mass ratio of the calcium-containing reaction auxiliary material to the waste aluminum electrolyte powder A is 0.6 to 0.9:1.

[0023] Furthermore, in step 3, the liquid-to-solid ratio of the alkaline solution to material B is in the range of 4 to 15 mL: 1 g.

[0024] Furthermore, in step 5a, the pH value of the solution during the pickling process is in the range of 3 to 6.

[0025] Furthermore, in step 5b, the organic solvent extractant is a mixed solvent consisting of a phosphate organic reagent, a ketone organic reagent and sulfonated kerosene.

[0026] Furthermore, in step 5b, the molar concentration ratio of the phosphate organic reagent to the ketone organic reagent is 0.5:0.1 to 0.1:0.5.

[0027] Compared with the prior art, the present invention can achieve at least one of the following beneficial effects:

[0028] a) In the method of the present invention, under the action of reaction auxiliary materials, the main cryolite phase in the waste aluminum electrolyte is rapidly decomposed in an alkaline solution, and the strong coordination precipitation effect between the calcium ions in the reaction auxiliary materials and the fluoride ions generated after the cryolite is dissolved is utilized to promote the rapid conversion of the fluoride ions entering the solution through alkaline leaching into CaF2 and precipitate out, thereby avoiding the coordination precipitation of fluoride ions with aluminum and lithium, thereby achieving efficient leaching of aluminum and lithium; compared with the existing alkaline leaching process technology, the present invention realizes the dissolution of the waste aluminum electrolyte in an alkaline solution by adding reaction auxiliary materials, and the fluoride ions generated after the dissolution react with the calcium ions directly to precipitate and remove, thereby avoiding the loss of valuable components such as leached Li / Al due to the complex precipitation of fluoride ions; compared with the existing alkaline roasting or calcium-added alkaline roasting-leaching process, the method of the present invention realizes the reaction process of cryolite dissolution and CaF2 formation in the waste aluminum electrolyte in one step, thereby avoiding the problems of low lithium aluminum leaching rate caused by incomplete reaction or difficult dissolution of LiF formed during the roasting process and the encapsulation of Li / Al components by CaF2 formed during the roasting process.

[0029] b) In the method of the present invention, through leaching in an alkaline system and calcification transformation of fluoride ions, compared with the acidic roasting / leaching system, the overflow loss and dispersion of fluorine in the form of HF gas are avoided, and the fluorine in the waste aluminum electrolyte is basically converted into calcium fluoride, with a high fluorine recovery rate; compared with the existing alkaline leaching and alkaline roasting-leaching process technologies, by regulating the alkaline leaching process, efficient dissolution of the waste aluminum electrolyte can be achieved, and the regulation of the calcium fluoride precipitation process can make the output calcium fluoride slag high in grade and less in impurities, and after acid washing and impurity removal, the grade can reach 98wt.%, which has a higher value.

[0030] c) In the method of the present invention, all major process steps are carried out in an alkaline system, and the alkaline lithium-containing solution obtained is compatible with the existing lithium solution extraction system. Compared with leaching in an acidic system and extracting after the leachate is adjusted to alkaline, the method of the present invention does not need to change the solution pH in a large range, the extraction process is easy to implement, and the lithium extraction recovery rate is high; the raffinate is added with reaction auxiliary materials again, on the one hand, it can realize the efficient removal and recovery of aluminum and the preparation of high-purity calcium aluminate products through the coordination precipitation of calcium ions and aluminate ions, and it is also beneficial to improve the alkalinity of the raffinate; the liquid after aluminum precipitation is treated by adding CaO, which can realize the concentration of alkali solution in a non-evaporation mode to obtain qualified alkali solution, with low energy consumption and easy operation, and the obtained Ca(OH)2 can be used as the calcium-containing reaction auxiliary material of step 2. The method of the present invention has a short process flow and low energy consumption. Through the method of the present invention, multiple components such as fluorine, lithium, aluminum, and sodium in the waste aluminum electrolyte are recovered step by step and efficiently, and the comprehensive utilization rate is high.

[0031] Other features and advantages of the present invention will be described in the following description, and part of them will become obvious from the description, or will be understood by practicing the present invention. The purpose and other advantages of the present invention can be realized and obtained by the contents particularly pointed out in the written description and the accompanying drawings. BRIEF DESCRIPTION OF THE DRAWINGS

[0032] The drawings are only for the purpose of illustrating particular embodiments and are not to be considered limiting of the present invention. Like reference symbols denote like components throughout the drawings.

[0033] Figure 1 It is a process flow chart of the present invention. DETAILED DESCRIPTION

[0034] Preferred embodiments of the present invention will be described in detail below in conjunction with the accompanying drawings, wherein the accompanying drawings constitute a part of the present invention and are used to explain the principles of the present invention together with the embodiments of the present invention.

[0035] The present invention provides a method for efficiently leaching and recovering fluorine, lithium, aluminum and sodium from waste aluminum electrolyte, comprising the following steps:

[0036] Step 1, crushing, grinding and screening the waste aluminum electrolyte blocks to obtain waste aluminum electrolyte powder A;

[0037] Step 2, mixing the waste aluminum electrolyte powder A and the calcium-containing reaction auxiliary material evenly to obtain material B;

[0038] Step 3, adding material B to the alkaline solution, stirring evenly, and adding it to the alkali-resistant reaction device for leaching reaction;

[0039] Step 4: After the reaction is completed, the material is filtered and separated to obtain filter residue (leaching residue) and filtrate (leaching liquid);

[0040] Step 5a, using an inorganic acid solution to acid-wash the filter residue obtained in step 4 to remove impurities, filter, and dry the filter cake to obtain high-grade CaF2;

[0041] Step 5b, extracting the filtrate obtained in step 4 with an organic solvent extractant to obtain a loaded organic phase and a raffinate;

[0042] Step 6a, stripping the loaded organic phase obtained in step 5b to obtain a lithium-containing solution, and the lithium-containing solution is concentrated and used for carbonation precipitation of lithium to obtain industrial-grade lithium carbonate;

[0043] Step 6b, adding a certain multiple of the theoretical amount of reaction auxiliary materials to the raffinate obtained in step 5b, stirring evenly to react and standing for a period of time, filtering and separating to obtain calcium aluminate precipitate, and obtaining a high-purity calcium aluminate product after acid washing;

[0044] Step 7, continue to add reaction auxiliary materials to the filtrate obtained in step 6b, stir evenly and let stand, filter and separate, the filter residue obtained is Ca(OH)2, Ca(OH)2 can be returned to step 2 as a reaction auxiliary material; the filtrate is a qualified alkali solution product.

[0045] In a specific embodiment, in step 1, the waste aluminum electrolyte is solid waste generated during aluminum electrolysis, the main component of which is cryolite, and the lithium content is 0.5-7 wt.%;

[0046] In a specific embodiment, in step 1, it is considered that when the particle size of the waste aluminum electrolyte powder A is too large, the leaching process efficiency is low, which may cause problems such as low lithium aluminum leaching; when the particle size is too small, it is easy to cause material agglomeration in the leaching process; therefore, the waste aluminum electrolyte blocks are crushed, ground, and screened to obtain a waste aluminum electrolyte powder A with a powder particle size of ≤74μm.

[0047] In a specific embodiment, in step 2, the calcium-containing reaction auxiliary material is capable of providing Ca 2+ The calcium-containing substance, for example, the reaction auxiliary material is at least one of CaO, CaCl2, CaF2, Ca(OH)2, Ca(NO3)2 and hydrates thereof. Exemplarily, the reaction auxiliary material is CaO, CaO+CaF2, CaO+Ca(OH)2+CaF2.

[0048] It should be noted that the main purpose of adding calcium-containing reaction auxiliary materials is to provide Ca2+ for binding with fluoride ions during alkali leaching. 2+ , so that the fluoride ions dissolved into the solution are quickly converted into calcium fluoride precipitates, reducing the fluoride ion concentration in the solution, and further promoting the alkaline leaching and dissolution of waste aluminum electrolyte. Calcium-containing substances that can achieve this purpose are within this protection range.

[0049] Compared with the prior art, the method of the present invention promotes the conversion of fluoride ions produced by the dissolution of waste aluminum electrolyte into CaF2 by adding calcium-containing reaction auxiliary materials, which further promotes the dissolution of waste aluminum electrolyte on the one hand, and can effectively avoid the formation of coordination precipitation between fluoride ions and dissolved aluminum, lithium, etc., which affects the aluminum-lithium leaching rate on the other hand. The method of the present invention can effectively realize the efficient recovery of fluorine resources and avoid the dispersion of fluorine and the problem of insufficient purity of the formed precipitate product.

[0050] In a specific embodiment, in step 2, the mass ratio of the calcium-containing reaction auxiliary material to the waste aluminum electrolyte powder A is 0.6 to 0.9:1, preferably, the mass ratio is 0.8:1.

[0051] In a specific embodiment, in step 3, the alkaline solution is prepared using at least one of NaOH and KOH, preferably, NaOH is used, and the concentration of the alkaline solution is 20-100 g / L, for example, 30 g / L, 40 g / L, 50 g / L, 60 g / L, 70 g / L, 80 g / L, 90 g / L.

[0052] It should be noted that when the concentration of the alkaline solution is low, the waste aluminum electrolyte is not completely dissolved; when the concentration of the alkaline solution is high, it is easy to make it difficult for the added calcium salt to enter the solution; both will cause a low lithium aluminum leaching rate; therefore, the concentration of the alkaline solution is controlled to be 20-100g / L.

[0053] In a specific embodiment, in step 3, the alkali-resistant reaction device includes but is not limited to a high-pressure reactor, a pipeline dissolution device, etc.

[0054] In a specific embodiment, in step 3, the liquid-to-solid ratio of the alkaline solution to material B is in the range of (4-15 mL): 1 g; for example, 5 mL: 1 g, 7 mL: 1 g, 10 mL: 1 g, 13 mL: 1 g.

[0055] It should be noted that if the liquid-to-solid ratio is too large, the single processing capacity of the reaction vessel of the same volume will be reduced; if the liquid-to-solid ratio is too small, it will cause problems such as high density of the mixed slurry, slow solid-liquid reaction, and insufficient alkaline solution reactants; therefore, the liquid-to-solid ratio of the alkaline solution to material B is controlled in the range of (4-15mL): 1g.

[0056] In a specific embodiment, in step 3, material B reacts with an alkaline solution at a reaction temperature of 120 to 200°C, a reaction time of 0.5 to 2 hours, and a stirring speed of 200 to 500 rpm. It should be noted that if the reaction temperature is lower than this range, the waste aluminum electrolyte is difficult to dissolve in the alkaline solution and the reaction kinetics are limited. If the temperature is too high, the energy consumption and pressure of the reaction vessel will be high. If the stirring speed is lower than this range, the solution will cause CaF2 precipitation to wrap the unreacted waste aluminum electrolyte, causing the leaching reaction to be hindered. If the stirring speed is too high, a vortex will easily form, which will not have a stirring and mixing effect. If the reaction time is lower than this range, the reaction will be incomplete, and if it is higher than this range, the reaction efficiency will be limited. Therefore, the reaction temperature is controlled to be 120-200°C, the reaction time is 0.5-2 hours, and the stirring speed is 200-500rpm; illustratively, the reaction temperature is 130°C, 150°C, 170°C, 190°C; the reaction time is 0.6 hour, 1.0 hour, 1.5 hours, 1.8 hours; the stirring speed is 250rpm, 300rpm, 350rpm, 400rpm, 450rpm.

[0057] In a specific embodiment, in step 5a, the filter residue obtained in step 4 is pickled with an inorganic acid solution, and the inorganic acid solution is prepared from at least one of hydrochloric acid and nitric acid. Preferably, it is prepared with hydrochloric acid. It should be noted that considering the economic cost, concentrated hydrochloric acid is preferably used for preparing the acid solution. The pickling process is mainly to remove the unreacted auxiliary materials such as CaO / Ca(OH)2 in the filter residue with the acid solution. If the pH of the solution is too low, the dissolution loss of CaF2 may occur. If the pH of the solution is too high, the removal of CaO / Ca(OH)2 will be incomplete, affecting the purity of CaF2. Therefore, the pH range of the solution during pickling is controlled to be 3-6. Exemplarily, the pH of the solution is 3, 4, 4.5, 5, 5.5.

[0058] In a specific embodiment, in step 5b, the organic solvent extractant is mainly a mixed solvent composed of phosphoric acid ester organic reagents, ketone organic reagents and sulfonated kerosene.

[0059] In a specific embodiment, in step 5b, the phosphoric acid ester organic reagents include at least one of tributyl phosphate, trioctyl phosphate, trialkyl phosphonate, trioctyl phosphine oxide, etc.; the ketone organic reagents include at least one of β-diketone, trifluorodimethylhexanedione, etc. The molar concentration ratio of the phosphoric acid ester organic reagent to the ketone organic reagent is 0.5:0.1 - 0.1:0.5.

[0060] In a specific embodiment, in step 5b, the pH value of the filtrate in step 4 is controlled to be 12.5 < pH < 14, the volume ratio of the organic solvent extractant to the filtrate is 1:20 - 1:1, the oscillation time is 5 - 30 min, and the temperature is 10 - 40 °C. Exemplarily, the pH is 12.5, 13, 13.5, 14, the volume ratio of the organic solvent extractant to the filtrate is 1:15, 1:10, 1:5, 1:2, the oscillation time is 10 min, 15 min, 20 min, 25 min, and the temperature is 15 °C, 20 °C, 25 °C, 30 °C, 35 °C.

[0061] It should be noted that the ketone organic reagent has a selective chelating effect on lithium in the solution, and the neutral phosphorus organic reagent, as a strong electron donor, can enhance the selective stability of the extract. At the same time, the extraction process is also affected by the organic phase concentration, the pH value of the solution, the organic phase / water phase volume ratio O / A, the oscillation time and the temperature. To ensure the complete extraction of lithium in the solution and avoid waste of organic reagents, the molar ratio of the phosphoric acid ester to the ketone organic phase is controlled to be 0.5:0.1 - 0.1:0.5.

[0062] In a specific embodiment, in step 6a, the stripping reagent is one of inorganic acids such as nitric acid, hydrochloric acid, sulfuric acid, carbonic acid, etc., and the hydrogen ion concentration in the stripping reagent is controlled to be 1 - 6 mol / L.

[0063] In a specific embodiment, in step 6a, in the stripping, the volume ratio of the loaded organic phase to the stripping reagent is 2:1 to 15:1, preferably 5:1 to 10:1.

[0064] It should be noted that, in order to consider the comprehensive utilization efficiency and cost of the stripping acid, as well as the subsequent regeneration of the organic phase, hydrochloric acid is preferably used as the stripping reagent, and the hydrogen ion concentration is controlled to be 1 to 6 mol / L. Preferably, the hydrogen ion concentration of the hydrochloric acid solution is 4 mol / L. In addition, to ensure normal phase separation and operation of the stripping process, the volume ratio of the stripping organic phase to the aqueous phase is 2:1 to 15:1. Preferably, the volume ratio of the stripping organic phase to the aqueous phase is 10:1.

[0065] In a specific embodiment, in step 6a, after the lithium-containing solution is concentrated, the lithium ion concentration reaches above 20 g / L, and sodium carbonate is added to precipitate lithium to obtain lithium carbonate.

[0066] In a specific embodiment, in step 6b, the reaction auxiliary material is at least one of CaO and Ca(OH)2, and preferably, Ca(OH)2 is used.

[0067] In a specific embodiment, in step 6b, the theoretical amount is the mass of the reaction auxiliary material calculated according to the aluminum concentration in the solution and the molecular formula of calcium aluminate precipitation.

[0068] In a specific embodiment, in step 6b, the added amount is 1 to 1.5 times the theoretical amount, and illustratively, the added amount is 1, 1.3, and 1.5 times; it should be noted that although the use of CaO can achieve the precipitation of the aluminum component in the raffinate, the addition of CaO will increase the alkalinity of the solution, which is not conducive to the reaction process. Therefore, Ca(OH)2 is preferably used as a reaction auxiliary material; the reaction temperature is 20 to 80°C, the stirring speed is 200 to 500 rpm, the reaction time is 0.5 to 2 hours, and the standing time is 0.5 to 1 hour. Schematically, the reaction temperature is 20°C, 40°C, 60°C, and 80°C; the reaction time is 1 hour and 1.5 hours; the stirring speed is 250rpm, 300rpm, 350rpm, 400rpm, and 450rpm.

[0069] In a specific embodiment, in step 6b, HCl and HNO3 solutions are used for pickling, and the pH value of the solution is 2 to 5; preferably, HCl solution is used; it should be noted that the acid solution is preferably prepared with concentrated hydrochloric acid in consideration of economic cost; the pickling process mainly uses an acid solution to remove the Ca(OH)2 auxiliary material that does not participate in the reaction in the filter residue. If the pH value of the solution is too low, it may cause the dissolution loss of calcium aluminate, and if the pH value of the solution is too high, it will cause incomplete removal of Ca(OH)2, affecting the purity of calcium aluminate; therefore, the pH value of the solution is controlled to be in the range of 2 to 5 during the pickling process.

[0070] In a specific embodiment, in step 7, the reaction auxiliary material is CaO.

[0071] It should be noted that the method of the present invention has an obvious beneficial effect on the alkaline leaching of cryolite phase. Since the main components of hazardous wastes generated during aluminum electrolysis, such as aluminum electrolysis overhaul slag and carbon powder, also contain cryolite phase, the method described in the present invention also has a good effect on the above two raw materials.

[0072] Example 1

[0073] This embodiment provides a method for efficiently leaching and recovering fluorine, lithium, aluminum and sodium from waste aluminum electrolyte, comprising the following steps:

[0074] Step 1, crushing, grinding and screening the waste aluminum electrolyte blocks to obtain waste aluminum electrolyte powder with a particle size of ≤74 μm;

[0075] Step 2, mixing the reaction auxiliary materials CaO+Ca(OH)2+CaF2 and the waste aluminum electrolyte powder in a mass ratio of 0.8:1;

[0076] Step 3, prepare 100g / L NaOH solution, stir the mixed material and NaOH solution at a liquid-solid ratio of 10mL:1g, put into an alkali-resistant reactor, and react at 150°C and 300rpm for 1.5 hours;

[0077] Step 4: After the reaction is completed, the material is filtered to obtain filter residue (leaching residue) and filtrate (leaching liquid);

[0078] Step 5a, using low concentration hydrochloric acid to pickle the filter residue to remove impurities, the solution pH is maintained at 4, filtering, and drying the filter cake to obtain high-grade calcium fluoride;

[0079] During the leaching process of step 4, the aluminum and lithium leaching rates were 95.34% and 94.27% respectively, the fluorine recovery rate was 99.87%, and the calcium fluoride grade reached 98.31wt.%;

[0080] Step 5b, the leachate (pH = 13.1) of step 4 and an organic solvent extractant (tributyl phosphate organic matter and β-diketone organic reagent, molar concentration ratio of 0.3:0.4, and a mixed reagent configured with sulfonated kerosene) are placed in a separatory funnel at a volume ratio of 40mL:4mL for mixing, and three-stage countercurrent extraction is performed, and then the obtained balanced load organic phase is mixed with 4mol / L hydrochloric acid stripping reagent at a volume ratio of O / A = 20mL:2mL, and two-stage countercurrent stripping is performed. The entire extraction-stripping process, the oscillation time is 10min, the temperature is 25°C, and the stripping solution finally obtained has a lithium concentration of 20.6g / L and a lithium recovery rate of 98.6%.

[0081] The raffinate was treated with Ca(OH)2 in an amount of 1, 1.3, and 1.5 times the theoretical amount, respectively. The reaction was carried out at 60°C and a stirring speed of 300 rpm for 1 hour. After standing for 1 hour, the aluminum was filtered and separated. The aluminum precipitation rates were 86.7%, 96.7%, and 97.9%, respectively.

[0082] The filter residue is pickled with a hydrochloric acid solution at a pH of 4, and the purity of the obtained calcium aluminate product reaches 98.6%; the filtrate is concentrated with alkali solution by adding CaO to obtain 30% alkali solution, and the overall sodium recovery rate reaches 98.5%.

[0083] Example 2

[0084] This embodiment provides a method for efficiently leaching and recovering fluorine, lithium, aluminum and sodium from waste aluminum electrolyte, comprising the following steps:

[0085] Step 1, crushing, grinding and screening the waste aluminum electrolyte blocks to obtain waste aluminum electrolyte powder with a particle size of ≤74 μm;

[0086] Step 2, mixing the reaction auxiliary materials CaO+Ca(OH)2+CaF2 and the waste aluminum electrolyte powder in a mass ratio of 0.8:1;

[0087] Step 3, prepare 100g / L NaOH solution, stir the mixed material and NaOH solution at a liquid-solid ratio of 10mL:1g, put into an alkali-resistant reactor, and react at 150°C and 300rpm for 1.5 hours;

[0088] Step 4: After the reaction is completed, the material is filtered to obtain filter residue (leaching residue) and filtrate (leaching liquid);

[0089] Step 5a, using low concentration hydrochloric acid to pickle the filter residue to remove impurities, the solution pH is maintained at 4, filtering, and drying the filter cake to obtain high-grade calcium fluoride;

[0090] During the leaching process of step 4, the aluminum and lithium leaching rates were 95.34% and 94.27% respectively, the fluorine recovery rate was 99.87%, and the calcium fluoride grade reached 98.31wt.%;

[0091] Step 5b, the leachate (pH = 13.6) of step 4 and an organic solvent extractant (a mixed reagent prepared with tributyl phosphate organic matter and β-diketone organic reagent, with a molar concentration ratio of 0.1:0.4 and sulfonated kerosene) are placed in a separatory funnel at a volume ratio of 40mL:2mL for mixing, and two-stage countercurrent stripping is performed. The entire extraction-stripping process is oscillated for 10min at a temperature of 25°C. The stripping solution finally obtained has a lithium concentration of 25.1g / L and a lithium recovery rate of 98.7%.

[0092] The raffinate was treated with Ca(OH)2 in an amount of 1.3 times the theoretical amount, respectively, and reacted at 20°C, 40°C, 60°C, and 80°C, with a stirring speed of 300 rpm for 1 hour. After standing for 1 hour, the mixture was filtered and separated, and the aluminum precipitation rates were 67.2%, 93.9%, 96.7%, and 96.9%, respectively;

[0093] The filter residue is pickled with a hydrochloric acid solution at a pH of 4, and the purity of the obtained calcium aluminate product reaches more than 96%; the filtrate is concentrated with alkali solution by adding CaO to obtain 30% alkali solution, and the overall recovery rate of sodium reaches more than 98%.

[0094] Example 3

[0095] This embodiment provides a method for efficiently leaching and recovering fluorine, lithium, aluminum and sodium from waste aluminum electrolyte, comprising the following steps:

[0096] Step 1 to step 5a are the same as in Example 1;

[0097] Step 5b, the leachate (pH = 14.0) of step 4 and an organic solvent extractant (tributyl phosphate organic matter and β-diketone organic reagent, molar concentration ratio of 0.2:0.2, and a mixed reagent configured with sulfonated kerosene) are placed in a separatory funnel at a volume ratio of 10mL:2mL for mixing, and three-stage countercurrent extraction is performed, and then the obtained balanced load organic phase is mixed with a 6mol / L hydrochloric acid stripping reagent at a volume ratio of O / A = 20mL:1mL, and two-stage countercurrent stripping is performed. The entire extraction-stripping process, the oscillation time is 10min, the temperature is 25°C, and the stripping solution finally obtained has a lithium concentration of 26.5g / L and a lithium recovery rate of 98.3%.

[0098] The raffinate was treated with Ca(OH)2 in an amount of 1.3 times the theoretical amount, and the reaction was carried out at 60°C and a stirring speed of 300 rpm for 0.5, 1.0, 1.5, and 2.0 hours, respectively. After standing for 1 hour, the aluminum was filtered and separated, and the aluminum precipitation rates were 89.7%, 96.7%, 96.9%, and 97.1%, respectively.

[0099] The filter residue is pickled with a hydrochloric acid solution at a pH of 4, and the purity of the obtained calcium aluminate product reaches more than 96%; the filtrate is concentrated with alkali solution by adding CaO to obtain 30% alkali solution, and the overall recovery rate of sodium reaches more than 98%.

[0100] Example 4

[0101] This embodiment provides a method for efficiently leaching and recovering fluorine, lithium, aluminum and sodium from waste aluminum electrolyte. The method of this embodiment is substantially the same as that of embodiment 1, except that:

[0102] In step 2, the mass ratio of the reaction auxiliary material to the waste aluminum electrolyte powder is 0.6:1;

[0103] Step 3: Prepare 50 g / L NaOH solution, and stir the mixed material and NaOH solution at a liquid-to-solid ratio of 8 mL:1 g.

[0104] During the leaching process of step 4, the aluminum and lithium leaching rates were 78.68% and 86.24% respectively, the fluorine recovery rate was 80.45%, and the calcium fluoride grade reached 83.72%;

[0105] In step 5b, the lithium recovery rate is 82.67%.

[0106] The raffinate was treated with Ca(OH)2 in an amount 1.3 times the theoretical amount, reacted at 60°C and a stirring speed of 300 rpm for 1 hour, allowed to stand for 1 hour, and filtered to separate. The aluminum precipitation rate was 94.82%.

[0107] The filter residue is pickled with a hydrochloric acid solution at a pH of 4, and the purity of the obtained calcium aluminate product reaches 86.21%; the filtrate is concentrated with alkali solution by adding CaO to obtain 30% alkali solution, and the overall sodium recovery rate reaches 85.49%.

[0108] In the method of the present invention, the lithium leaching rate of aluminum during the leaching process can reach more than 78%, for example, 78.68% to 95.34%; the lithium leaching rate can reach more than 86%, for example, 86.24% to 95%; the fluorine recovery rate reaches more than 80%, for example, 80.45% to 99.9%; the calcium fluoride grade reaches more than 83%, for example, 83.72% to 99%; the lithium recovery rate is more than 82%, for example, 82.67% to 99%; the aluminum precipitation rate reaches more than 67.2%, for example, 67.2% to 97.9%; and the overall sodium recovery rate reaches more than 85%, for example, 85.49% to 99%.

[0109] Comparative Example 1

[0110] This comparative example provides a method for leaching and recovering fluorine, lithium, aluminum and sodium from waste aluminum electrolyte, comprising the following steps:

[0111] The waste aluminum electrolyte blocks are crushed, ground and sieved to obtain waste aluminum electrolyte powder with a particle size of ≤74μm, and a 100g / L NaOH solution is prepared. The waste aluminum electrolyte powder is evenly stirred with the NaOH solution at a liquid-solid ratio of 10mL:1g without adding reaction auxiliary materials, and then put into an alkali-resistant reactor, and reacted for 1.5 hours at 150°C and 300rpm. After the reaction, the material is filtered to obtain a filter residue and a filtrate, respectively. The filter residue is pickled and impurities are removed with a hydrochloric acid solution, and the pH of the solution is maintained at 4, filtered, and the filter cake is dried.

[0112] During the leaching process, the leaching rates of aluminum and lithium were 67.61% and 21.08% respectively, and the fluorine recovery rate was 58.02%;

[0113] The filtrate was subjected to three-stage countercurrent extraction using a mixture of tributyl phosphate organic matter: β-diketone organic reagent (molar concentration ratio of 0.3:0.4) and sulfonated kerosene as an organic solvent extractant, and the organic solvent extractant and the filtrate were in a volume ratio of 1:10; the obtained balanced load organic phase was subjected to two-stage stripping using a 4 mol / L HCl solution, the stripping phase ratio O / A=10 / 1, and the oscillation time of the entire extraction-stripping process was controlled at 10 min and the temperature was 25°C; the lithium recovery rate was 18.58%.

[0114] By comparing Example 1 with Comparative Example 1, it can be seen that without adding reaction auxiliary materials, the lithium aluminum leaching rate during the leaching process is low, and the fluoride ion content in the leachate is high; since the cryolite in the waste aluminum electrolyte is not completely dissolved, the filter residue obtained in step 4 is mainly a cryolite phase, and since no reaction auxiliary materials are added during the leaching process, fluorine cannot be converted into calcium fluoride and enter the filter residue, the fluorine recovery rate is low, and no CaF2 product can be obtained. Due to the low lithium leaching rate, the subsequent lithium recovery rate will also be low.

[0115] Comparative Example 2

[0116] This comparative example provides a method for leaching waste aluminum electrolyte and recovering fluorine, lithium, aluminum and sodium. The steps of this comparative example are roughly the same as those of Example 1, except that:

[0117] In step 2, the reaction auxiliary material and the waste aluminum electrolyte powder are evenly mixed in a mass ratio of 0.3:1;

[0118] During the leaching process, the aluminum and lithium leaching rates were 75.64% and 68.97% respectively, the fluorine recovery rate was 75.31%, and the calcium fluoride grade reached 68.19wt.%.

[0119] By comparing Example 1 and Comparative Example 2, it can be seen that when the amount of reaction auxiliary material added is low, the cryolite is not completely dissolved during the leaching process, the aluminum-lithium leaching rate is slightly low, and the fluoride ions in the solution cannot be completely converted into precipitates, resulting in a low fluorine recovery rate and a low purity of the calcium fluoride product; although the impact on indicators such as the lithium extraction rate and the stripping rate is small, due to the low overall lithium leaching rate, the total lithium recovery rate is also low; there is also a certain impact on the process of adding CaO to concentrate the alkali solution, which will result in a low alkali recovery rate.

[0120] The above description is only a preferred specific implementation manner of the present invention, but the protection scope of the present invention is not limited thereto. Any changes or substitutions that can be easily conceived by any technician familiar with the technical field within the technical scope disclosed by the present invention should be covered within the protection scope of the present invention.

Claims

1. A method for efficiently leaching and recovering fluorine, lithium, aluminum and sodium from waste aluminum electrolyte, characterized in that: The method comprises the following steps: Step 1, crushing, grinding and screening the waste aluminum electrolyte blocks to obtain waste aluminum electrolyte powder A; Step 2, mixing the waste aluminum electrolyte powder A and the calcium-containing reaction auxiliary material evenly to obtain material B; Step 3, adding material B to the alkaline solution, stirring evenly, and adding it to the alkali-resistant reaction device for leaching reaction; Step 4: After the reaction is completed, the material is filtered and separated to obtain filter residue and filtrate respectively; Step 5a, using an inorganic acid solution to acid-wash the filter residue obtained in step 4, purify and remove impurities, filter, and dry the filter cake to obtain CaF2; Step 5b, extracting the filtrate obtained in step 4 with an organic solvent extractant to obtain a loaded organic phase and a raffinate; Step 6a, stripping the loaded organic phase obtained in step 5b to obtain a lithium-containing solution; Step 6b, adding reaction auxiliary materials to the raffinate obtained in step 5b, stirring and reacting evenly and standing for a period of time, filtering and separating to obtain calcium aluminate precipitate, and obtaining a high-purity calcium aluminate product after acid washing; Step 7: Continue to add reaction auxiliary materials to the filtrate obtained in step 6b, stir evenly, let stand, and filter and separate.

2. The method according to claim 1, characterized in that: In step 6b, the reaction auxiliary material is at least one of CaO and Ca(OH)2.

3. The method according to claim 1, characterized in that In the step 6b, the reaction temperature is 20-80°C.

4. The method according to claim 3, characterized in that: In the step 6b, the stirring speed is 200 to 500 rpm, and the reaction time is 0.5 to 2 hours.

5. The method according to claim 1, characterized in that In step 6b, the pH value of the solution during the pickling process is in the range of 2 to 5.

6. The method according to claim 1, characterized in that In the step 2, the mass ratio of the calcium-containing reaction auxiliary material to the waste aluminum electrolyte powder A is 0.6-0.9:

1.

7. The method according to claim 1, characterized in that In the step 3, the liquid-to-solid ratio of the alkaline solution to the material B is in the range of 4 to 15 mL: 1 g.

8. The method according to claim 1, characterized in that In step 5a, the pH value of the solution during the pickling process is in the range of 3 to 6.

9. The method according to any one of claims 1 to 8, characterized in that: In the step 5b, the organic solvent extractant is a mixed solvent consisting of a phosphate organic reagent, a ketone organic reagent and sulfonated kerosene.

10. The method according to claim 9, characterized in that In the step 5b, the molar concentration ratio of the phosphate organic reagent to the ketone organic reagent is 0.5:0.1 to 0.1:0.5.

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