Method for recovering lithium and fluorine from electrolytic aluminum solid waste

Abstract

The invention relates to a method for recovering lithium and fluorine from electrolytic aluminum solid waste, which comprises the following steps: firstly mixing the electrolytic aluminum solid waste and an iron-containing acid solution for complexing leaching to obtain a lithium-containing leachate; mixing the lithium-containing leaching solution with a calcium-containing additive for precipitation reaction to obtain a fluorine-removed lithium-containing leaching solution and calcium fluoride; then the fluorine-removed lithium-containing leaching solution is subjected to pH adjustment through alkali liquor, and an iron-removed and aluminum-removed lithium-containing leaching solution is obtained; finally, adsorbing and separating the lithium-containing leaching solution with iron and aluminum removed to obtain a lithium-rich solution; according to the method, lithium and fluorine resources in the electrolytic aluminum solid waste are efficiently recycled, the leaching rate of Li < + > is preferably as high as 87.2% or above, the loss rate of lithium is preferably as low as 1.00% or below, the recovery rate of the fluorine element is preferably as high as 85.9% or above, generation of a large amount of HF gas is avoided, and the method is environmentally friendly and low in recovery cost.

Description

Technical Field

[0001] The present invention relates to the technical field of lithium extraction from electrolytic aluminum solid waste, and in particular to a method for recovering lithium and fluorine from electrolytic aluminum solid waste. Background Art

[0002] Aluminum electrolyte is the molten salt medium in the process of producing aluminum metal by electrolysis of alumina. It is mainly composed of cryosphere and sub-cryosphere, and is also added with aluminum fluoride, calcium fluoride or lithium fluoride. In order to ensure the normal and stable operation of the aluminum electrolytic cell, it needs to be replaced with aluminum electrolyte with lower impurity content after the aluminum electrolytic cell has been running for a period of time. The replaced aluminum electrolyte with higher impurity content is the electrolytic aluminum waste electrolyte, which is one of the electrolytic aluminum solid wastes and contains a certain proportion of fluoride and lithium resources to be recycled.

[0003] At present, the methods for extracting lithium from electrolytic aluminum solid waste mainly include two categories: roasting method and chemical leaching method. Since electrolytic aluminum waste contains a large amount of metal elements, a large number of associated elements including aluminum are also leached into the solution during the acid leaching process. The leaching of these associated elements will seriously affect the subsequent lithium enrichment process. It is often necessary to go through multiple steps of impurity removal to obtain a lithium-containing solution that meets the enrichment requirements, especially the main associated elements aluminum and fluorine. These large amounts of associated metal elements are not thoroughly purified, which will greatly affect the purity of the final lithium product. In addition, a large amount of HF gas will be generated during the acid leaching process, causing environmental pollution and dangerous operations. In order to solve many problems in the process of extracting lithium from electrolytic aluminum solid waste by aluminum hydrochloric acid leaching, roasting processes are currently mainly used, including sulfuric acid roasting or composite salt roasting, but it still has problems such as low lithium recovery rate, low fluorine recovery rate, and pollution.

[0004] For example, CN117735585A discloses a method for enriching and extracting lithium from low-lithium electrolytic aluminum overhaul slag and preparing aluminum fluoride, which uses aluminum salt acid leaching to extract lithium, that is, using sulfuric acid and aluminum sulfate with a concentration of ≥98% for acid leaching. During the leaching process, the aluminum ion content is too high, resulting in a large amount of lithium being adsorbed in the precipitate and unable to be effectively separated. The acid concentration is too high and is not suitable for subsequent steps. In addition, a large amount of HF escapes, resulting in serious pollution, low lithium recovery rate and high energy consumption in the aluminum salt acid leaching method for lithium extraction.

[0005] For example, CN115216645A discloses a method for extracting lithium from electrolytic aluminum waste slag by a mixed salt calcination method. The method uses electrolytic aluminum waste slag as raw material, potassium and calcium salts as auxiliary materials, and adopts calcification roasting, potassium salt re-burning, alkali leaching, solid-liquid separation, cryogenic impurity removal and purification processes. The content of other metal impurity ions in the leaching solution is low, and the lithium salt is easy to separate and extract. However, in order to ensure the extraction rate of lithium, the process requires a roasting process, and there is still a pollution problem. When the aluminum element is precipitated, a large amount of lithium element will be adsorbed, which affects the enrichment process. In addition, there are problems such as lithium loss in the purification process and failure to achieve the recovery and utilization of fluorine element.

[0006] In view of this, how to develop a low-cost and high-efficiency method for recovering lithium and fluorine from electrolytic aluminum solid waste and realize the green resource comprehensive utilization of electrolytic aluminum solid waste has become an urgent problem to be solved. Summary of the invention

[0007] In order to solve the above technical problems, the present invention utilizes an iron-containing acidic solution to carry out complex leaching on electrolytic aluminum solid waste, thereby releasing lithium and other impurity metal elements including iron, aluminum, etc. in the electrolytic aluminum solid waste; combining subsequent precipitation reaction, pH adjustment and adsorption separation process, the fluorine element in the leachate is enriched and recovered, and the impurity metal elements in the leachate are removed to obtain a lithium-rich solution; while improving the lithium leaching rate, a low lithium loss rate is ensured, and the fluorine recovery rate is also improved, thereby realizing efficient recovery and utilization of lithium and fluorine resources in electrolytic aluminum solid waste, avoiding the generation of HF gas, and the entire process is simple, the raw material cost is low and it is environmentally friendly.

[0008] To achieve this object, the present invention adopts the following technical solutions:

[0009] The present invention provides a method for recovering lithium and fluorine from electrolytic aluminum solid waste, the method comprising the following steps:

[0010] (1) mixing electrolytic aluminum solid waste and an iron-containing acidic solution for complex leaching to obtain a lithium-containing leachate;

[0011] (2) mixing the lithium-containing leachate and the calcium-containing additive in step (1) to carry out precipitation reaction to obtain a defluorinated lithium-containing leachate and calcium fluoride;

[0012] (3) The fluorine-removed lithium-containing leachate in step (2) is pH-adjusted with an alkali solution to obtain an iron-removed aluminum-containing lithium-containing leachate;

[0013] (4) The iron-aluminum-removed lithium-containing leaching solution in step (3) is subjected to adsorption separation to obtain a lithium-rich solution.

[0014] The present invention firstly uses an iron-containing acidic solution to carry out complex leaching on electrolytic aluminum solid waste, wherein Fe 3+ Under acidic conditions, it can react with F in electrolytic aluminum solid waste - Complexation to form stable [FeF6] 3- The complex ions are used to destroy the structure of the electrolytic aluminum solid waste, and metal elements such as lithium, aluminum and iron are released from the electrolytic aluminum solid waste into the leachate; then a calcium fluoride product is obtained by adding a calcium-containing additive for precipitation reaction to enrich and recover the fluorine element, and then the iron and aluminum in the leachate are removed through a pH adjustment process; finally, a lithium-rich solution is obtained through adsorption separation; compared with the traditional acid leaching method, the generation of HF gas is avoided, the leaching rate of lithium and the recovery rate of fluorine in the electrolytic aluminum solid waste are improved, and the green resource utilization of the electrolytic aluminum solid waste is realized.

[0015] In addition, the present invention adopts Fe in the iron-containing acidic solution 3+ Complex leaching and precipitation defluorination with calcium-containing additives can not only recover calcium fluoride products, but also the fluoride ion concentration in the iron and aluminum removal process is low and no lithium fluoride, aluminum fluoride, cryolite, or sub-cryosite is generated, thereby reducing lithium loss due to the generation of lithium fluoride and the re-generation of lithium-containing cryolite. Moreover, the iron-aluminum slag produced in the subsequent iron-aluminum removal process is a mixed crystal of iron-aluminum hydroxide, in which the iron ions and aluminum ions are isomorphously replaced, resulting in a change in the structure of aluminum hydroxide that is easily adsorbed for lithium, and a significant reduction in the lithium adsorption performance, thereby reducing lithium loss in the iron-aluminum removal process.

[0016] It is worth noting that, compared with the aluminum-containing acidic solution, the method of the present invention using the iron-containing acidic solution avoids the problem of lithium loss caused by the lithium element and the aluminum element entering the slag phase simultaneously due to the excessive aluminum ion concentration in the subsequent aluminum removal process, thereby achieving the goal of increasing the lithium leaching rate while reducing the lithium loss rate, thereby increasing the recovery rate of lithium resources in electrolytic aluminum solid waste, and at the same time greatly reducing the raw material cost.

[0017] Preferably, the electrolytic aluminum solid waste in step (1) includes any one of electrolytic aluminum waste electrolyte, electrolytic aluminum overhaul slag or electrolytic aluminum carbon slag, or a combination of at least two of them, wherein typical but non-limiting combinations include a combination of electrolytic aluminum waste electrolyte and electrolytic aluminum overhaul slag, a combination of electrolytic aluminum waste electrolyte and electrolytic aluminum carbon slag, or a combination of electrolytic aluminum overhaul slag and electrolytic aluminum carbon slag, etc.

[0018] Preferably, the electrolytic aluminum solid waste includes the following components in terms of mass percentage: Li: 0.2-2.5wt%, Al: 8.1-20wt%, Fe: 0.08-0.89wt%, K: 1.45-2.6wt%, Mg: 0.2-1.1wt%, F: 12.6-33.1wt%, Si: 0.1-8.5wt% and Na: 13.9-18.1wt%.

[0019] Wherein Li: 0.2-2.5wt%, for example, 0.2wt%, 0.4wt%, 0.6wt%, 0.8wt%, 1.0wt%, 1.2wt%, 1.5wt%, 2.0wt%, 2.2wt% or 2.5wt%; Al: 8.1-20wt%, for example, 8.1wt%, 10.0wt%, 12wt%, 15wt%, 18wt% or 20wt%; Fe: 0.08-0.89wt% , for example, it can be 0.08wt%, 0.10wt%, 0.15wt%, 0.20wt%, 0.30wt%, 0.40wt%, 0.50wt%, 0.60wt%, 0.70wt%, 0.80wt% or 0.89wt%, etc.; K: 1.45-2.6wt%, for example, it can be 1.45wt%, 1.50wt%, 1.60wt%, 1.80wt%, 2.0wt%, 2.2wt%, 2.4wt% or 2.6wt%; Mg: 0.2-1.1wt%, for example, 0.2wt%, 0.4wt%, 0.6wt%, 0.8wt% or 1.1wt%; Na: 13.9-18.1wt%, for example, 13.9wt%, 15.0wt%, 15.5wt%, 16.0wt%, 16.5wt%, 17.0wt%, 17.5wt% or 18.1wt%; F: 12.6-33.1wt%, for example For example, it can be 12.6wt%, 15.0wt%, 18.0wt%, 20.0wt%, 23wt%, 25wt%, 28wt%, 30wt% or 33.1wt%, etc.; Si: 0.1-8.5wt%, for example, it can be 0.1wt%, 0.5wt%, 0.8wt%, 1.0wt%, 2.0wt%, 3.0wt%, 4.0wt%, 5.0wt%, 6.0wt%, 7.0wt% or 8.5wt%, etc.

[0020] Preferably, the iron-containing acidic solution in step (1) comprises any one of ferric sulfate solution, ferric chloride solution or ferric nitrate solution, or a combination of at least two of them, wherein typical but non-limiting combinations include a combination of ferric sulfate solution and ferric chloride solution, a combination of ferric chloride solution and ferric nitrate solution, or a combination of ferric sulfate solution and ferric nitrate solution, etc., preferably a ferric sulfate solution.

[0021] The present invention further preferably uses the iron-containing acidic solution as a ferric sulfate solution, which is beneficial for achieving a subsequent full-process circulation process without the influence of volatile corrosive gases (HCl, etc.) on the equipment.

[0022] Preferably, the iron-containing acidic solution contains Fe 3+The concentration is 0.1-3.0 mol / L, for example, it can be 0.1 mol / L, 0.2 mol / L, 0.5 mol / L, 0.8 mol / L, 1.0 mol / L, 1.2 mol / L, 1.5 mol / L, 1.8 mol / L, 2.0 mol / L, 2.2 mol / L, 2.5 mol / L, 2.8 mol / L or 3.0 mol / L, preferably 0.5-2.5 mol / L, and more preferably 1-1.5 mol / L.

[0023] The present invention further preferably contains Fe 3+ The concentration of Fe in the iron-containing acidic solution is 1-1.5 mol / L, which is beneficial to the complex leaching process, fully destroys the structure of the electrolytic aluminum solid waste, and improves the lithium leaching rate and the fluorine recovery rate. 3+ If the concentration of Fe is too low, the complexation will be insufficient, and some lithium elements in the electrolytic aluminum solid waste will not be released into the leaching solution, thereby reducing the leaching rate of lithium elements and the recovery rate of fluorine elements; if the Fe-containing acidic solution contains Fe 3+ The concentration is too high, and there is no obvious improvement in the lithium leaching effect and fluorine recovery effect, resulting in a waste of reagents.

[0024] Preferably, the pH of the iron-containing acidic solution is less than 2, for example, it can be 0, 0.2, 0.5, 0.8, 1.0, 1.2, 1.5 or 1.8, and is preferably 0-1.

[0025] The present invention further preferably has a pH value of less than 2 for the iron-containing acidic solution, which is beneficial to the complexing leaching effect of step (1). 3+ F in electrolytic aluminum solid waste - The lithium element in the electrolytic aluminum solid waste is fully leached out by complexation, thereby improving the lithium leaching rate and fluorine recovery rate; if the pH of the iron-containing acidic solution is ≥ 2, the Fe in the iron-containing acidic solution will 3+ It will hydrolyze and enter the slag phase in advance, which will lead to reduced quality of fluorine products and loss of lithium.

[0026] Preferably, the electrolytic aluminum solid waste is crushed before mixing in step (1).

[0027] Preferably, the particle size of the crushed electrolytic aluminum solid waste is 80-200 mesh, for example, it can be 80 mesh, 90 mesh, 100 mesh, 110 mesh, 120 mesh, 150 mesh, 180 mesh or 200 mesh.

[0028] Preferably, the present invention further preferably crushes the electrolytic aluminum solid waste before mixing, and the particle size of the electrolytic aluminum solid waste after crushing is 80-200 mesh, which is beneficial to increase the contact area between the iron-containing acidic solution and the electrolytic aluminum solid waste, thereby improving the complex leaching efficiency; if the particle size of the electrolytic aluminum solid waste after crushing is too large, the effect is equivalent to that without crushing, and the leaching of lithium element is relatively slow; if the particle size of the electrolytic aluminum solid waste after crushing is too low, the leaching effect is not significantly improved, but requires greater energy consumption.

[0029] Preferably, the liquid-to-solid ratio of the iron-containing acidic solution and the electrolytic aluminum solid waste is (0.5-50):1, for example, it can be 0.5:1, 1.5:1, 3:1, 5:1, 8:1, 10:1, 12:1, 15:1, 18:1, 20:1, 25:1, 30:1, 35:1, 40:1, 45:1 or 50:1, etc., preferably (5-30):1, and more preferably (10-15):1.

[0030] The present invention further preferably has a liquid-to-solid ratio of (10-15):1 to the iron-containing acidic solution and the electrolytic aluminum solid waste, which is beneficial to the leaching of lithium and the enrichment of fluorine in the electrolytic aluminum solid waste, thereby improving the leaching rate of lithium and the recovery rate of fluorine in the electrolytic aluminum solid waste; if the liquid-to-solid ratio of the iron-containing acidic solution and the electrolytic aluminum solid waste is low, the leaching rate of lithium and the recovery rate of fluorine will be reduced; if the liquid-to-solid ratio of the iron-containing acidic solution and the electrolytic aluminum solid waste is high, the effect will not be significantly improved, which will cause waste of reagents and high costs.

[0031] Preferably, the complex leaching in step (1) is multi-stage countercurrent leaching.

[0032] Optionally, the complex leaching in step (1) includes at least one multi-stage countercurrent leaching, for example, 1 time, 2 times, 3 times, 4 times, 5 times, 6 times, 7 times, 8 times, 9 times or 10 times.

[0033] Preferably, the number of stages of the multi-stage countercurrent leaching is 2 to 10, for example, it can be 2, 4, 6, 8 or 10, etc., preferably 3 to 6.

[0034] Preferably, the complex leaching process in step (1) is accompanied by heating.

[0035] Preferably, the heating comprises microwave enhanced heating and / or ultrasound enhanced heating.

[0036] Preferably, the temperature of the complex leaching is 25-90°C, for example, 25°C, 30°C, 50°C, 60°C, 80°C or 90°C, etc., preferably 50-80°C.

[0037] Preferably, the complex leaching time is 10 to 30 minutes, for example, 10 minutes, 15 minutes, 20 minutes, 25 minutes or 30 minutes.

[0038] The present invention further preferably accompanies the complex leaching process with microwave enhanced heating and / or ultrasonic enhanced heating, which is beneficial to promote the complex leaching process, increase the lithium leaching rate and fluorine enrichment rate in the electrolytic aluminum solid waste, thereby shortening the complex leaching time, further avoiding the diffusion of oxygen in the air in the complex leaching system during the heating process, and avoiding the diffusion of HF gas or cyanide, a poisonous component in the electrolytic aluminum overhaul slag.

[0039] Preferably, the calcium-containing additive in step (2) comprises any one of calcium oxide, calcium sulfate or calcium hydroxide, or a combination of at least two thereof, wherein typical but non-limiting combinations include a combination of calcium oxide and calcium sulfate, a combination of calcium oxide and calcium hydroxide, or a combination of calcium sulfate and calcium hydroxide, etc.

[0040] Preferably, the molar ratio of Ca in the calcium-containing additive in step (2) to F in the lithium-containing leachate is (0.25-1.5):1, for example, it can be 0.25:1, 0.50:1, 0.75:1, 0.80:1, 0.85:1, 0.90:1, 0.95:1, 1:1, 1.15:1, 1.25:1, 1.35:1 or 1.5:1, etc., preferably (0.5-1.25):1, and more preferably (0.75-1):1.

[0041] The present invention further preferably has a molar ratio of Ca in the calcium-containing additive to F in the lithium-containing leachate of step (2) of (0.75-1):1, which is conducive to fully precipitating and enriching F in the lithium-containing leachate, thereby improving the recovery rate of fluorine in electrolytic aluminum solid waste; if the molar ratio of Ca in the calcium-containing additive to F in the lithium-containing leachate is low, some fluorine will still remain in the lithium-containing leachate, resulting in a decrease in the recovery rate of fluorine; if the molar ratio of Ca in the calcium-containing additive to F in the lithium-containing leachate is high, the recovery rate of fluorine is not significantly improved, but partial iron hydrolysis leads to lithium loss, and calcium impurity ions are introduced into the leachate, thereby increasing the recovery cost.

[0042] Preferably, the temperature of the precipitation reaction in step (2) is 25-90°C, for example, 25°C, 30°C, 50°C, 60°C, 80°C or 90°C.

[0043] Preferably, the precipitation reaction in step (2) is a multi-stage countercurrent precipitation reaction.

[0044] Preferably, the precipitation reaction in step (2) includes at least one multi-stage countercurrent precipitation reaction, for example, it can be 1 time, 2 times, 3 times, 4 times, 5 times, 6 times, 7 times, 8 times, 9 times or 10 times.

[0045] Preferably, the number of stages of the multi-stage countercurrent precipitation reaction is 2 to 10, for example, it can be 2, 4, 6, 8 or 10, etc., preferably 3 to 6.

[0046] Preferably, the endpoint of the precipitation reaction in step (2) is that the pH of the defluorinated lithium-containing leachate is less than 2.5, for example, it can be 1.0, 1.2, 1.5, 1.8, 2.0, 2.2 or 2.4, preferably 1.5 to 2.2.

[0047] The present invention further prefers that the endpoint of the precipitation reaction in step (2) is that the pH of the defluorinated lithium-containing leachate is 1.5-2.2, which is conducive to the precipitation reaction of fluorine and calcium, and improves the recovery rate of fluorine in electrolytic aluminum solid waste. If the endpoint of the precipitation reaction is that the pH of the defluorinated lithium-containing leachate is too low, F in the lithium-containing leachate cannot be effectively precipitated, resulting in a decrease in the recovery rate of fluorine, and lithium and fluorine combine to form lithium fluoride, resulting in lithium loss; if the endpoint of the precipitation reaction is that the pH of the defluorinated lithium-containing leachate is too high, it will cause Fe to be hydrolyzed and thus lead to lithium loss.

[0048] Preferably, the alkali solution in step (3) comprises sodium hydroxide solution and / or sodium carbonate solution.

[0049] Preferably, the mass concentration of the alkali solution in step (3) is 10-60wt%, for example, it can be 10wt%, 15wt%, 20wt%, 25wt%, 30wt%, 35wt%, 40wt%, 45wt%, 50wt%, 55wt% or 60wt%, etc., preferably 20-50wt%;

[0050] Preferably, the temperature during the pH adjustment process in step (3) is 25-60°C, for example, it can be 25°C, 30°C, 35°C, 40°C, 45°C, 50°C, 55°C or 60°C.

[0051] Preferably, the pH adjustment process in step (3) is accompanied by stirring.

[0052] Preferably, the endpoint of the pH adjustment in step (3) is that the pH of the iron-removing aluminum-containing lithium leachate is 5 to 8, for example, it can be 5.0, 5.3, 5.5, 5.8, 6.2, 6.4, 6.8, 7.0, 7.2, 7.4, 7.6, 7.8 or 8.0, etc., preferably 6.0 to 7.0.

[0053] The present invention further preferably adopts the end point of the pH adjustment in step (3) as the pH value of the lithium-containing leaching solution after iron and aluminum removal to be 5 to 8, so that the impurity iron ions and aluminum ions in the lithium-containing leaching solution fully enter the slag phase, thereby avoiding further decomposition of aluminum hydroxide in an overly alkaline environment, resulting in some Al elements remaining in the iron and aluminum removal leaching solution, which is not conducive to the purity of the lithium product prepared by the subsequent lithium-rich solution.

[0054] Preferably, the adsorbent used for the adsorptive separation in step (4) comprises a titanium-based adsorbent and / or an aluminum-based adsorbent.

[0055] Preferably, the liquid-to-solid ratio of the iron-removing aluminum leaching solution to the adsorbent is (10-20):1, for example, it can be 10:1, 12:1, 15:1, 18:1 or 20:1.

[0056] Preferably, the temperature of the adsorption separation is 20-30°C, for example, 20°C, 22°C, 25°C, 28°C or 30°C.

[0057] Preferably, the adsorption separation time is 5 to 8 hours, for example, 5 hours, 6 hours, 7 hours or 8 hours.

[0058] Preferably, the lithium-rich solution in step (4) + The concentration is 0.5-50 g / L, for example, it can be 0.5 g / L, 1.0 g / L, 1.5 g / L, 2 g / L, 2.5 g / L, 5 g / L, 6.5 g / L, 7 g / L, 8.5 g / L, 10 g / L, 20 g / L, 25 g / L, 30 g / L, 35 g / L, 40 g / L, 45 g / L or 50 g / L, etc., preferably 5-15 g / L.

[0059] As a further preferred technical solution of the present invention, refer to Figure 1 The process is carried out, and the method comprises the following steps:

[0060] (1) Firstly, the electrolytic aluminum solid waste is crushed to a particle size of 80-200 meshes, and then the electrolytic aluminum solid waste is mixed with an iron-containing acidic solution at a liquid-solid ratio of (0.5-50):1; the pH of the iron-containing acidic solution is less than 2 and Fe 3+ The concentration of lithium ions is 0.1-3.0 mol / L; performing 2-10 levels of complex leaching for 10-30 minutes at 25-90° C. and microwave-enhanced heating and / or ultrasonic-enhanced heating to obtain a lithium-containing leaching solution;

[0061] The iron-containing acidic solution includes any one of a ferric sulfate solution, a ferric chloride solution or a ferric nitrate solution, or a combination of at least two thereof;

[0062] (2) mixing the lithium-containing leachate and the calcium-containing additive in step (1) at a molar ratio of Ca in the calcium-containing additive to F in the lithium-containing leachate of (0.25-1.5):1, and performing 2-10 levels of countercurrent precipitation reaction at 25-90° C. to obtain a defluorinated lithium-containing leachate and calcium fluoride; the end point of the precipitation reaction is that the pH of the defluorinated lithium-containing leachate is less than 2.5;

[0063] (3) at 25 to 60° C. and under stirring conditions, the pH of the fluorine-removing lithium-containing leachate in step (2) is adjusted with an alkali solution having a mass concentration of 10 to 60 wt % to obtain an iron-removing aluminum-containing lithium-containing leachate; the end point of the pH adjustment is that the pH of the iron-removing aluminum-containing lithium-containing leachate is 5.0 to 8.0;

[0064] (4) at 20-30° C., the de-ironized aluminum-containing lithium-containing leachate of step (3) is subjected to adsorption separation for 5-8 hours at a liquid-solid ratio of the de-ironized aluminum-containing leachate to the adsorbent of (10-20):1 to obtain Li + The concentration of the lithium-rich solution is 0.5 to 50 g / L; the adsorbent used in the adsorption separation includes a titanium-based adsorbent and / or an aluminum-based adsorbent.

[0065] Optionally, step (1) further comprises treating the leached residue after the complex leaching with H + The acid solution with a concentration of 0.01 to 1 mol / L (for example, 0.01 mol / L, 0.02 mol / L, 0.05 mol / L, 0.1 mol / L, 0.2 mol / L, 0.3 mol / L, 0.4 mol / L, 0.5 mol / L, 0.6 mol / L, 0.8 mol / L or 1 mol / L, etc., preferably 0.02 to 0.1 mol / L) is used for acid leaching to obtain an iron-containing acid solution, wherein the liquid-to-solid ratio of the acid leaching is (0.1 to 5):1 (for example, 0.1:1, 0 .2:1, 0.5:1, 0.7:1, 1.0:1, 1.2:1, 1.5:1, 2:1, 2.5:1, 3:1, 3.5:1, 4:1, 4.5:1 or 5:1, etc., preferably (0.5-1.5):1); the end point of the acid leaching is that the pH of the iron-containing acid solution is 1-2.5 (for example, it can be 1, 1.2, 1.5, 1.8, 2.0, 2.2 or 2.5, etc.), and then the iron-containing acid solution can be reused as the iron-containing acid solution after pH adjustment to continue the complex leaching process of step (1).

[0066] Optionally, step (3) further comprises treating the iron-aluminum slag obtained after pH adjustment with OH at 25 to 60°C (for example, 25°C, 30°C, 35°C, 40°C, 45°C, 50°C, 55°C or 60°C, etc.) -The alkali solution with a concentration of 10 to 60 wt% (for example, 10 wt%, 20 wt%, 30 wt%, 40 wt%, 50 wt% or 60 wt%, etc., preferably 20 to 25 wt%) is used for alkali leaching until the pH of the solution is 10.5 to 16 (for example, 10.5, 12, 12.5, 13, 13.5, 14, 14.5, 15, 15.5 or 16, etc.) to obtain iron hydroxide and aluminum-containing alkali solution; the liquid-to-solid ratio of the alkali leaching is (0.1 to 5):1 (for example, it can be 0.1:1, 0.2:1, 0.5:1, 0.7:1, 1.0:1, 1.2:1, 1.5:1, 2:1, 2.5:1, 3:1, 3.5:1, 4:1, 4.5:1 or 5:1, etc., preferably (0.5-1.5):1); the ferric hydroxide can be further treated with acid to obtain an iron-containing acidic solution for reuse in the complex leaching process of step (1); the aluminum-containing alkali solution can be further used to prepare an aluminum hydroxide product.

[0067] Optionally, before the adsorption separation in step (4), the step further includes evaporating and concentrating the iron-removing aluminum-containing lithium leachate at 60-100°C (for example, 60°C, 65°C, 70°C, 75°C, 80°C, 85°C, 90°C, 95°C or 100°C, etc.) and a vacuum degree of 10-90kPa (for example, 10kPa, 20kPa, 30kPa, 40kPa, 50kPa, 60kPa, 70kPa, 80kPa or 90kPa, etc.).

[0068] Optionally, the adsorption residual liquid obtained after the adsorption separation in step (4) can be further subjected to electrodialysis to obtain corresponding acid solution and alkali solution, which can be used for acid treatment or alkali leaching of intermediate products to obtain iron-containing acidic solution for recycling.

[0069] The waste residue or intermediate product of the method of the present invention can be purified to realize the recycling of the iron-containing acidic solution, the process flow does not generate wastewater, the process is simple and the process is environmentally friendly.

[0070] Compared with the prior art, the present invention has at least the following beneficial effects:

[0071] (1) The method for recovering lithium and fluorine from electrolytic aluminum solid waste provided by the present invention utilizes Fe in the iron-containing acidic solution 3+ F in electrolytic aluminum solid waste - The complexation destroys the structure of the electrolytic aluminum solid waste, thereby transferring the lithium, iron, aluminum and fluorine elements in the electrolytic aluminum solid waste to the leachate, and then adding calcium-containing additives to precipitate and remove fluorine to obtain a calcium fluoride product, and then using alkaline solution to adjust the pH to remove iron and aluminum, and finally obtaining a lithium-rich solution through adsorption separation; the recovery of lithium and fluorine in electrolytic aluminum solid waste is realized.

[0072] (2) The method for recovering lithium and fluorine from electrolytic aluminum solid waste provided by the present invention further optimizes the pH, Fe 3+ The concentration of and its liquid-to-solid ratio with electrolytic aluminum solid waste, the molar ratio of Ca in the calcium-containing additive to F in the lithium-containing leaching solution, the size of the pH adjustment endpoint pH and other process parameters further improve the recovery effect of lithium and fluorine in electrolytic aluminum solid waste. The lithium leaching rate is preferably as high as 87.2% or more, the lithium loss rate is preferably as low as below 1.00%, and the recovery rate of fluorine element is preferably as high as 85.9% or more, thereby realizing the efficient recovery and reuse of lithium and fluorine resources in electrolytic aluminum solid waste. BRIEF DESCRIPTION OF THE DRAWINGS

[0073] Figure 1 It is a process flow chart of the method for recovering lithium and fluorine from electrolytic aluminum solid waste provided by the present invention. DETAILED DESCRIPTION

[0074] The technical solution of the present invention is further described below by specific implementation methods. However, the following examples are only simplified examples of the present invention and do not represent or limit the scope of protection of the present invention. The scope of protection of the present invention shall be subject to the claims.

[0075] The titanium-based adsorbent used in the following embodiments or comparative examples is from Yunnan Gangfeng New Materials Co., Ltd., model G100; the aluminum-based adsorbent is from Zhejiang Zhengguang Industrial Co., Ltd., model ZG LR328; it is worth noting that the present invention does not limit the type of the adsorbent, and any adsorbent that can selectively adsorb lithium can be used.

[0076] Example 1

[0077] This embodiment provides a method for recovering lithium and fluorine from electrolytic aluminum solid waste, the method comprising the following steps:

[0078] (1) First, the electrolytic aluminum solid waste is crushed to a particle size of 100 mesh, and then the electrolytic aluminum solid waste is mixed with a ferric sulfate solution at a liquid-solid ratio of 14:1; the pH of the ferric sulfate solution is 0.5 and Fe 3+ The concentration is 1.2 mol / L; three-level complex leaching is carried out at 60°C and microwave-enhanced heating for 30 minutes to obtain a lithium-containing leaching solution;

[0079] (2) mixing the lithium-containing leachate and CaO in step (1) at a molar ratio of 0.8:1 between Ca in CaO and F in the lithium-containing leachate, and performing a two-stage countercurrent precipitation reaction at 80° C. to obtain a defluorinated lithium-containing leachate and calcium fluoride; the endpoint of the precipitation reaction is that the pH of the defluorinated lithium-containing leachate is 2.0;

[0080] (3) at 40° C. and under stirring conditions, the pH of the fluorine-removing lithium-containing leachate in step (2) is adjusted by a sodium hydroxide solution having a mass concentration of 22.5 wt % to obtain an iron-removing aluminum-containing lithium-containing leachate; the end point of the pH adjustment is that the pH of the iron-removing aluminum-containing lithium-containing leachate is 6.5;

[0081] (4) At 30° C., the iron-aluminum-removing lithium-containing leachate of step (3) was subjected to adsorption separation for 8 hours at a liquid-solid ratio of the iron-aluminum-removing leachate to the titanium-based adsorbent of 20:1 to obtain Li + The concentration of the lithium-rich solution is 8.1 g / L. The electrolytic aluminum solid waste in this embodiment is electrolytic aluminum waste electrolyte, and its main components and their corresponding contents are shown in Table 1;

[0082] Table 1

[0083]

[0084] Example 2

[0085] This embodiment provides a method for recovering lithium and fluorine from electrolytic aluminum solid waste, the method comprising the following steps:

[0086] (1) First, the electrolytic aluminum solid waste is crushed to a particle size of 80 mesh, and then the electrolytic aluminum solid waste is mixed with a ferric sulfate solution at a liquid-solid ratio of 10:1; the pH of the ferric sulfate solution is 1 and Fe 3+ The concentration of 1 mol / L; performing level 3 complex leaching for 20 min at 75°C and ultrasonic enhanced heating to obtain a lithium-containing leaching solution;

[0087] (2) mixing the lithium-containing leachate of step (1) and CaSO4 at a molar ratio of Ca in CaSO4 to F in the lithium-containing leachate of 0.75:1, and performing a three-stage countercurrent precipitation reaction at 90° C. to obtain a defluorinated lithium-containing leachate and calcium fluoride; the endpoint of the precipitation reaction is that the pH of the defluorinated lithium-containing leachate is 1.5;

[0088] (3) at 25° C. and under stirring conditions, the pH of the fluorine-removing lithium-containing leachate in step (2) is adjusted by a sodium hydroxide solution having a mass concentration of 25 wt % to obtain an iron-removing aluminum-containing lithium-containing leachate; the end point of the pH adjustment is that the pH of the iron-removing aluminum-containing lithium-containing leachate is 7.0;

[0089] (4) At 25° C., the iron-aluminum-removing lithium-containing leachate of step (3) was subjected to adsorption separation for 7 hours at a liquid-solid ratio of the iron-aluminum-removing leachate to the titanium-based adsorbent of 10:1 to obtain Li + A lithium-rich solution with a concentration of 14.8 g / L.

[0090] The aluminum electrolysis solid waste in this embodiment is aluminum electrolysis waste electrolyte, and its main components and their corresponding contents are shown in Table 2;

[0091] Table 2

[0092]

[0093] Example 3

[0094] This embodiment provides a method for recovering lithium and fluorine from electrolytic aluminum solid waste, the method comprising the following steps:

[0095] (1) First, the electrolytic aluminum solid waste is crushed to a particle size of 200 mesh, and then the electrolytic aluminum solid waste is mixed with a ferric chloride solution at a liquid-solid ratio of 15:1; the pH of the ferric chloride solution is 0 and Fe 3+ The concentration of lithium ions is 1.5 mol / L; a second-level complex leaching is performed at 80°C and under ultrasonic enhanced heating for 10 minutes to obtain a lithium-containing leaching solution;

[0096] (2) mixing the lithium-containing leachate and calcium hydroxide in step (1) at a molar ratio of Ca in calcium hydroxide to F in the lithium-containing leachate of 0.9:1, and performing a three-stage countercurrent precipitation reaction at 60° C. to obtain a defluorinated lithium-containing leachate and calcium fluoride; the endpoint of the precipitation reaction is that the pH of the defluorinated lithium-containing leachate is 2.2;

[0097] (3) at 60° C. and under stirring conditions, the pH of the fluorine-removing lithium-containing leachate in step (2) is adjusted by a sodium carbonate solution with a mass concentration of 50 wt % to obtain an iron-removing aluminum-containing lithium-containing leachate; the end point of the pH adjustment is that the pH of the iron-removing aluminum-containing lithium-containing leachate is 6.0;

[0098] (4) At 20° C., the iron-removing aluminum-containing lithium leachate of step (3) was subjected to adsorption separation for 6 hours at a liquid-solid ratio of the iron-removing aluminum leachate to the aluminum-based adsorbent of 15:1 to obtain Li + A lithium-rich solution with a concentration of 5.0g / L.

[0099] The electrolytic aluminum solid waste in this embodiment is a combination of electrolytic aluminum overhaul slag and electrolytic aluminum waste electrolyte, and its main components and their corresponding contents are shown in Table 3;

[0100] Table 3

[0101]

[0102] Example 4

[0103] This embodiment provides a method for recovering lithium and fluorine from electrolytic aluminum solid waste, wherein the method removes Fe 3+Except that the concentration of is 0.8 mol / L, the rest are the same as in Example 1.

[0104] Example 5

[0105] This embodiment provides a method for recovering lithium and fluorine from electrolytic aluminum solid waste. The method is the same as that of Example 1 except that the pH of the ferric sulfate solution in step (1) is 2.

[0106] Example 6

[0107] This embodiment provides a method for recovering lithium and fluorine from electrolytic aluminum solid waste. The method is the same as that of Embodiment 1 except that the liquid-to-solid ratio in step (1) is 9:1.

[0108] Example 7

[0109] This embodiment provides a method for recovering lithium and fluorine from electrolytic aluminum solid waste. The method is the same as that of Embodiment 1 except that the end point of the precipitation reaction in step (2) is that the pH of the fluorine-containing lithium leaching solution is 2.5.

[0110] Example 8

[0111] This embodiment provides a method for recovering lithium and fluorine from electrolytic aluminum solid waste. The method is the same as that of Example 1 except that the molar ratio of Ca in CaO to F in the lithium-containing leachate in step (2) is 0.6:1.

[0112] Example 9

[0113] This embodiment provides a method for recovering lithium and fluorine from electrolytic aluminum solid waste. The method is the same as that of Embodiment 1 except that the molar ratio of Ca in CaO to F in the lithium-containing leachate in step (2) is 1.25:1.

[0114] 2. Comparison

[0115] Comparative Example 1

[0116] This comparative example provides a method for recovering lithium and fluorine from electrolytic aluminum solid waste. The method is the same as Example 1 except that the ferric sulfate solution in step (1) of Example 1 is replaced by an aluminum sulfate solution, and the pH of the aluminum sulfate solution is the same as the pH of the ferric sulfate solution, and the aluminum ion concentration is the same as the iron ion concentration.

[0117] Comparative Example 2

[0118] This comparative example provides a method for recovering lithium and fluorine from electrolytic aluminum solid waste. The method is the same as Example 1 except that the ferric sulfate solution in step (1) is replaced by a sulfuric acid solution, and the pH of the sulfuric acid solution is the same as the pH of the ferric sulfate solution, that is, the sulfuric acid solution is used for acid leaching.

[0119] 3. Test and its results

[0120] The lithium leaching rate, lithium loss rate, and fluorine recovery rate in the methods provided in the above embodiments and comparative examples were tested and calculated, and the calculation formula is as follows:

[0121]

[0122] And test the Li in lithium-rich solution + The results are shown in Table 4.

[0123] Table 4

[0124]

[0125]

[0126] It can be seen from Table 4 that:

[0127] (1) From Examples 1 to 3, it can be seen that the present invention utilizes Fe in the iron-containing acidic solution 3+ F in electrolytic aluminum solid waste - Complexation to form stable [FeF6] 3- The complex ions are used to destroy the structure of the electrolytic aluminum solid waste, and the lithium and fluorine elements are transferred to the leaching solution. The lithium leaching rate is as high as 87.2% or more. Combined with the subsequent precipitation reaction, the fluorine element is converted into a fluorine product for recovery, and the fluorine recovery rate is as high as 85.9% or more. Then, the iron and aluminum impurities are removed by pH adjustment. The lithium loss rate in the treatment process after complex leaching is as low as less than 1.00%. Then, Li is separated by adsorption. + The lithium-rich solution with a concentration of more than 5.0g / L realizes the efficient recovery and utilization of lithium and fluorine resources in electrolytic aluminum solid waste.

[0128] (2) Combining Example 1 with Examples 4 to 6, it can be seen that compared with Example 1, the Fe content of the ferric sulfate solution in Example 4 is 3+ The concentration of lithium was low, resulting in a lithium leaching rate of only 85.3% and a fluorine recovery rate of only 84.6% in the complex leaching process. + The concentration of ferrous sulfate solution in Example 5 is only 5.7 g / L; the pH of the ferrous sulfate solution in Example 5 is too high, resulting in a lithium leaching rate of only 90.1% during the complex leaching process, a lithium loss rate in the treatment process after the complex leaching increased to 5.32%, and a fluorine recovery rate of only 87.7%. +The concentration of Fe in the iron-containing acidic solution is only 6.8 g / L; the liquid-solid ratio of the ferric sulfate solution to the electrolytic aluminum solid waste in step (1) of Example 6 is low, resulting in a lithium leaching rate of only 83.4%, a lithium loss rate of up to 1.10%, and a fluorine recovery rate of only 81.2%; This shows that the present invention further prefers Fe in the iron-containing acidic solution 3+ The concentration of the iron-containing acidic solution is 1-1.5 mol / L, or the pH of the iron-containing acidic solution is preferably 0-1, or the liquid-solid ratio of the iron-containing acidic solution to the electrolytic aluminum solid waste is preferably (10-15):1, which further increases the lithium leaching rate in the electrolytic aluminum solid waste and reduces the lithium loss rate, thereby increasing the Li in the obtained lithium-rich solution. + concentration, and also improves the recovery rate of fluorine in electrolytic aluminum solid waste.

[0129] (3) Combining Example 1 with Examples 7 to 9, it can be seen that: compared with Example 1, the end point of the precipitation reaction in Example 7 is that the pH of the defluorinated lithium-containing leachate is too high, resulting in an increase in the lithium loss rate to 5.13%. The molar ratio of Ca in CaO to F in the lithium-containing leachate in Example 8 is too low, resulting in a decrease in the recovery rate of fluorine to 86.1%, and part of the lithium combines with fluorine to form lithium fluoride, increasing the lithium loss rate to 5.72%. The molar ratio of Ca in CaO to F in the lithium-containing leachate in Example 9 is too high. Although the recovery rate of fluorine is slightly increased, the lithium loss rate is increased to 2.70% at the same time. This shows that the present invention further preferably sets the end point of the precipitation reaction to be that the pH of the defluorinated lithium-containing leachate is 1.5 to 2.2, or further preferably sets the molar ratio of Ca in the calcium-containing additive to F in the lithium-containing leachate in step (2) to be (0.75 to 1):1, both of which can further improve the recovery rate of fluorine in electrolytic aluminum solid waste and reduce the lithium loss rate.

[0130] (4) Based on Example 1 and Comparative Examples 1 to 2, it can be seen that: compared with Example 1, since step (1) in Comparative Example 1 is carried out using aluminum sulfate solution, the aluminum ion concentration in the system is too high during the aluminum removal process in step (3), resulting in lithium ions and aluminum ions in the defluorinated lithium-containing leachate co-precipitating into the slag phase and causing lithium loss. Although the lithium leaching rate is slightly higher, its lithium loss rate is as high as 19.80%, and since part of the fluorine will become sub-cryolite and enter the slag phase, the fluorine recovery rate is reduced to 78.1%. Since step (1) in Comparative Example 2 is acid leaching using sulfuric acid solution, the lithium leaching rate is only 63.5%, the lithium loss rate is as high as 15.90%, and the recovery rate of fluorine element is reduced to 60.2%. In addition, a large amount of HF gas will be generated to pollute the environment.

[0131] The applicant declares that the above is only a specific implementation mode of the present invention, but the protection scope of the present invention is not limited thereto. Those skilled in the art should understand that any changes or substitutions that can be easily thought of by those skilled in the art within the technical scope disclosed by the present invention shall fall within the protection scope and disclosure scope of the present invention.

Claims

1. A method for recovering lithium and fluorine from electrolytic aluminum solid waste, characterized in that: The method comprises the following steps: (1) mixing the electrolytic aluminum solid waste and the iron-containing acidic solution for complex leaching to obtain a lithium-containing leachate; (2) mixing the lithium-containing leachate and the calcium-containing additive in step (1) to carry out precipitation reaction to obtain a defluorinated lithium-containing leachate and calcium fluoride; (3) The fluorine-free lithium-containing leachate in step (2) is pH-adjusted with an alkali solution to obtain an iron-free and aluminum-free lithium-containing leachate; (4) The iron-aluminum-removed lithium-containing leaching solution in step (3) is subjected to adsorption separation to obtain a lithium-rich solution.

2. The method according to claim 1, characterized in that The electrolytic aluminum solid waste in step (1) includes any one of electrolytic aluminum waste electrolyte, electrolytic aluminum overhaul slag or electrolytic aluminum carbon slag, or a combination of at least two thereof; Preferably, the electrolytic aluminum solid waste includes the following components in terms of mass percentage: Li: 0.2-2.5wt%, Al: 8.1-20wt%, Fe: 0.08-0.89wt%, K: 1.45-2.6wt%, Mg: 0.2-1.1wt%, Na: 13.9-18.1wt%, F: 12.6-33.1wt% and Si: 0.1-8.5wt%.

3. The method according to claim 1 or 2, characterized in that: The iron-containing acidic solution in step (1) comprises any one of ferric sulfate solution, ferric chloride solution or ferric nitrate solution or a combination of at least two thereof, preferably ferric sulfate solution; Preferably, the iron-containing acidic solution contains Fe 3+ The concentration is 0.1 to 3.0 mol / L, preferably 0.5 to 2.5 mol / L, and more preferably 1 to 1.5 mol / L; Preferably, the pH of the iron-containing acidic solution is less than 2, preferably 0-1.

4. The method according to any one of claims 1 to 3, characterized in that: In step (1), the electrolytic aluminum solid waste is crushed before mixing; Preferably, the particle size of the crushed electrolytic aluminum solid waste is 80-200 mesh; Preferably, the liquid-to-solid ratio of the iron-containing acidic solution to the electrolytic aluminum solid waste is (0.5-50):1, preferably (5-30):1, and further preferably (10-15):

1.

5. The method according to any one of claims 1 to 4, characterized in that: The complex leaching in step (1) is multi-stage countercurrent leaching; Preferably, the number of stages of the multi-stage countercurrent leaching is 2 to 10, preferably 3 to 6; Preferably, the complex leaching process in step (1) is accompanied by heating; Preferably, the heating comprises microwave enhanced heating and / or ultrasound enhanced heating; Preferably, the temperature of the complex leaching is 25 to 90°C, preferably 60 to 80°C; Preferably, the complex leaching time is 10 to 30 minutes.

6. The method according to any one of claims 1 to 5, characterized in that: The calcium-containing additive in step (2) includes any one of calcium oxide, calcium sulfate or calcium hydroxide, or a combination of at least two thereof; Preferably, the molar ratio of Ca in the calcium-containing additive in step (2) to F in the lithium-containing leachate is (0.25-1.5):1, preferably (0.5-1.25):1, and further preferably (0.75-1):

1.

7. The method according to any one of claims 1 to 6, characterized in that: The temperature of the precipitation reaction in step (2) is 25 to 90° C. Preferably, the precipitation reaction in step (2) is a multi-stage countercurrent precipitation reaction; Preferably, the number of stages of the multi-stage countercurrent precipitation reaction is 2 to 10, preferably 3 to 6; Preferably, the endpoint of the precipitation reaction in step (2) is that the pH of the defluorinated lithium-containing leaching solution is less than 2.5, preferably 1.5 to 2.

2.

8. The method according to any one of claims 1 to 7, characterized in that: The alkali solution in step (3) includes sodium hydroxide solution and / or sodium carbonate solution; Preferably, the mass concentration of the alkali solution in step (3) is 10-60wt%, preferably 20-50wt%; preferably, the temperature during the pH adjustment in step (3) is 25-60°C; Preferably, the pH adjustment process in step (3) is accompanied by stirring; Preferably, the endpoint of the pH adjustment in step (3) is that the pH of the iron-removing, aluminum-removing, and lithium-containing leaching solution is 5.0 to 8.0, preferably 6.0 to 7.

0.

9. The method according to any one of claims 1 to 8, characterized in that: The adsorbent used in the adsorption separation in step (4) includes a titanium-based adsorbent and / or an aluminum-based adsorbent; Preferably, the liquid-to-solid ratio of the iron-removing aluminum leaching solution to the adsorbent is (10-20):1; Preferably, the temperature of the adsorption separation is 20 to 30°C; Preferably, the adsorption separation time is 5 to 8 hours; Preferably, the lithium-rich solution in step (4) + The concentration is 0.5 to 50 g / L, preferably 5 to 15 g / L.

10. The method according to any one of claims 1 to 9, characterized in that: The method comprises the following steps: (1) Firstly, the electrolytic aluminum solid waste is crushed to a particle size of 80-200 meshes, and then the electrolytic aluminum solid waste is mixed with an iron-containing acidic solution at a liquid-solid ratio of (0.5-50):1; the pH of the iron-containing acidic solution is less than 2 and Fe 3+ The concentration of the lithium ion battery is 0.1-3.0 mol / L; performing 2-10 levels of complex leaching for 10-30 minutes at 25-90° C. and microwave-enhanced heating and / or ultrasonic-enhanced heating to obtain a lithium-containing leaching solution; The iron-containing acidic solution includes any one of a ferric sulfate solution, a ferric chloride solution or a ferric nitrate solution, or a combination of at least two thereof; (2) mixing the lithium-containing leachate and the calcium-containing additive in step (1) at a molar ratio of Ca in the calcium-containing additive to F in the lithium-containing leachate of (0.25-1.5):1, and performing 2-10 levels of countercurrent precipitation reaction at 25-90° C. to obtain a defluorinated lithium-containing leachate and calcium fluoride; the end point of the precipitation reaction is that the pH of the defluorinated lithium-containing leachate is less than 2.5; (3) at 25 to 60° C. and under stirring conditions, the pH of the fluorine-removing lithium-containing leachate in step (2) is adjusted with an alkali solution having a mass concentration of 10 to 60 wt % to obtain an iron-removing aluminum-containing lithium-containing leachate; the end point of the pH adjustment is that the pH of the iron-removing aluminum-containing lithium-containing leachate is 5.0 to 8.0; (4) at 20-30° C., the de-ironized aluminum-containing lithium-containing leachate of step (3) is subjected to adsorption separation for 5-8 hours at a liquid-solid ratio of the de-ironized aluminum-containing leachate to the adsorbent of (10-20):1 to obtain Li + The concentration of the lithium-rich solution is 0.5 to 50 g / L; the adsorbent used in the adsorption separation includes a titanium-based adsorbent and / or an aluminum-based adsorbent.

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