Method for step-by-step acidification lithium extraction and multi-element comprehensive recovery of lepidolite ore

Abstract

The invention discloses a method for step-by-step acidification lithium extraction and multi-element comprehensive recovery of lepidolite ore to solve the problems that the leaching rate of a large number of impurity elements in the acidification stage is high, and the difficulty and cost of subsequent lithium purification are increased. Comprising the following steps: (1) roasting lepidolite ore at 700-900 DEG C for defluorination to obtain defluorinated lepidolite ore; (2) the defluorinated lepidolite ore is subjected to primary acidification leaching, primary acidification leaching liquid and water leaching residues are obtained through the primary acidification leaching, and the primary acidification leaching enables the leaching rate of the lithium element to be larger than or equal to 90%, the leaching rate of the sodium element to be 50%-70%, the leaching rate of the potassium element to be 10%-30% and the leaching rate of the aluminum element to be 15%-30%; and (3) separating the primary acidification leaching solution from the water leaching residues, and carrying out subsequent treatment on the separated lithium-containing leaching solution to prepare a battery-grade lithium carbonate product, the primary acidification leaching comprises the steps that defluorinated lepidolite ore and concentrated sulfuric acid are mixed for acidification, the mass ratio of the concentrated sulfuric acid to the defluorinated lepidolite ore is (0.4-0.6): 1, and size mixing leaching is conducted after acidification.

Description

Technical Field

[0001] This invention relates to a method for stepwise acidification of lithium mica ore for lithium extraction and comprehensive recovery of multiple elements. Background Technology

[0002] Lepidolite is an important lithium resource, containing not only lithium but also valuable elements such as potassium, rubidium, cesium, aluminum, and silicon. With the rapid development of the new energy industry and the increasing demand for lithium, how to efficiently, economically, and environmentally extract lithium from lepidolite while comprehensively recovering other valuable elements has become a crucial issue in lithium resource development. Currently, the main process for lithium extraction from lepidolite is the sulfuric acid process, which is widely used due to its simplicity and low cost.

[0003] Existing sulfuric acid processes typically involve reacting high-concentration sulfuric acid with defluorinated lepidolite pretreated by acidification. This allows for ion exchange between the H⁺ in the sulfuric acid and the Li⁺ and K⁺ cations between the lepidolite layers. Simultaneously, the sulfuric acid disrupts the aluminosilicate structure, allowing metal elements to enter the solution as sulfates, thus leaching lithium and other metals. Currently, defluorination pretreatment usually involves high-temperature roasting at around 1000℃ to remove fluorine from the lepidolite as HF or SiF₄, primarily to reduce the corrosion of equipment by hydrofluoric acid during subsequent acidification. However, existing defluorination processes often involve excessively high temperatures, leading to sintering of lepidolite particles, reduced specific surface area, and decreased reactivity. Alternatively, adding additives such as CaCO₃ or CaO can lower the defluorination temperature, but this introduces impurities such as calcium, increasing the burden on subsequent impurity removal processes. In the acidification stage, in order to achieve a high lithium leaching rate (over 90%), high acid leaching (the acid-to-ore ratio is usually above 1:1) must be used. This also leads to the simultaneous leaching of a large number of impurity elements, such as potassium, aluminum, and iron, with leaching rates reaching over 80%. This not only increases the difficulty and cost of subsequent lithium purification but also makes it difficult to separate and recover other valuable elements. Summary of the Invention

[0004] The purpose of this invention is to provide a method for stepwise acidification and lithium extraction from lepidolite ore, as well as comprehensive recovery of multiple elements, to solve the problem of high leaching rates of a large number of impurity elements during the acidification stage, which increases the difficulty and cost of subsequent lithium purification.

[0005] A method for stepwise acid leaching of lithium from lepidolite and comprehensive recovery of multiple elements includes: (1) roasting lepidolite at 700℃-900℃ to remove fluoride, obtaining defluorinated lepidolite; (2) subjecting the defluorinated lepidolite to a primary acid leaching process, obtaining a primary acid leaching solution and a water leaching residue through the primary acid leaching process, wherein the primary acid leaching process enables the following leaching rates: lithium ≥90%, sodium 50%-70%, potassium 10%-30%, and aluminum 15%-30%; (3) separating the primary acid leaching solution from the water leaching residue, and further processing the separated lithium-containing leaching solution to obtain battery-grade lithium carbonate products; (4) further processing the water leaching solution. The leaching residue is subjected to secondary acid leaching or chlorination treatment. Secondary acid leaching solution and silica slag are obtained through secondary acid leaching. Secondary acid leaching can allow most of the metal elements to enter the secondary acid leaching solution; or metal chloride and silica slag are obtained through chlorination treatment; (5) the secondary acid leaching solution and silica slag are separated, or the metal chloride and silica slag are separated, and then valuable products are further recovered from the secondary acid leaching solution or metal chloride; wherein, primary acid leaching includes: acidifying defluorinated lithium mica ore with concentrated sulfuric acid, the mass ratio of concentrated sulfuric acid to defluorinated lithium mica ore is (0.4-0.6):1, and slurry leaching is carried out after acidification.

[0006] As an optimization and / or instance of the above-mentioned method for stepwise acid leaching of lithium and comprehensive recovery of multiple elements in lepidolite ore, the single acid leaching also results in: rubidium leaching rate of 10%-30%, cesium leaching rate of 10%-30%, calcium leaching rate of ≤10%, magnesium leaching rate of ≤15%, manganese leaching rate of 60%-80%, and iron leaching rate of 20%-40%.

[0007] As an optimization and / or instance of the above-mentioned stepwise acid leaching method for lithium extraction and multi-element comprehensive recovery of lithium mica ore, the secondary acid leaching includes: mixing water leaching residue with concentrated sulfuric acid for acidification, wherein the mass ratio of concentrated sulfuric acid to water leaching residue is (0.6-1.2):1, and then performing slurry leaching after acidification.

[0008] As an optimization and / or instance of the above-mentioned method for stepwise acidification and lithium extraction and multi-element comprehensive recovery of lithium mica ore, the chlorination treatment includes: chlorinating and roasting the water-leached residue at a temperature below 1400°C to chlorinate the metal elements in the water-leached residue except for silicon. During the chlorination roasting, ferric chloride and aluminum chloride escape as gases. After the chlorination roasting, the remaining chlorides are mixed with silicon dioxide. Subsequently, these chlorides are separated from the silicon dioxide by adding water leaching, and then the chlorides are further purified.

[0009] As an optimization and / or instance of the above-mentioned method for stepwise acidification and lithium extraction and multi-element comprehensive recovery of lithium mica ore, the chlorination treatment includes: chlorinating and roasting the water-leached residue at a temperature of ≥1400℃ to chlorinate silicon and other metal elements in the water-leached residue. During chlorination and roasting, silicon tetrachloride, ferric chloride, and aluminum chloride escape in gaseous form and are recovered by staged cooling using the different sublimation temperatures of silicon tetrachloride, ferric chloride, and aluminum chloride.

[0010] As an optimization and / or instance of the above-mentioned stepwise acid leaching lithium extraction and multi-element comprehensive recovery method for lithium mica ore, the silica slag obtained by separating the secondary acid leaching solution from the silica slag has a SiO2 content ≥ 85wt%.

[0011] As an optimization and / or instance of the above-mentioned method for stepwise acid leaching of lithium and comprehensive recovery of multiple elements from lepidolite ore, the valuable products recovered from the secondary acid leaching solution include: alum formation: by controlling temperature and concentration conditions, aluminum, potassium, rubidium, and cesium in the secondary acid leaching solution react with sulfate ions to form potassium alum and rubidium-cesium alum, which are then filtered to obtain alum products; roasting leaching: the alum products are roasted and decomposed, and the roasting products are separated by water leaching to obtain alumina; cooling crystallization: potassium sulfate is obtained from the roasting leaching solution through cooling crystallization; and fractional extraction: rubidium and cesium are recovered separately from the cooling crystallization mother liquor through fractional extraction.

[0012] As an optimization and / or instance of the above-mentioned stepwise acidification lithium extraction and multi-element comprehensive recovery method for lithium mica ore, the subsequent treatment of lithium-containing leachate to obtain battery-grade lithium carbonate products includes: alum formation: controlling temperature and concentration conditions to cause aluminum, potassium and sulfate ions in the lithium-containing leachate to form potassium alum precipitate, filtering to separate potassium alum and reduce the potassium and aluminum content in the leachate; impurity removal: adding sodium hydroxide to the alum-formed solution to adjust the pH to 11-12, causing impurities such as iron, aluminum, calcium, and magnesium to form hydroxide precipitates, filtering to obtain purified solution; MVR evaporation and concentration: evaporating and concentrating the purified solution through MVR to achieve a lithium ion concentration of 20-30 g / L; lithium precipitation: adding sodium carbonate to the concentrated solution, reacting at 90℃-95℃ to generate lithium carbonate precipitate, filtering, washing, and drying to obtain battery-grade lithium carbonate products.

[0013] This invention breaks through the conventional technical approach that "high lithium leaching rate requires high acid content, and high acid content inevitably leads to high impurities." By defluorinating and roasting lepidolite within a specific temperature range of 700℃-900℃, fluorine is effectively removed to prevent equipment corrosion, while maintaining the loose and porous structure of the defluorinated lepidolite, significantly improving its reactivity. The crystal structure of the lepidolite ore after defluorination treatment within this temperature range is optimized. In a single acid leaching process, a moderate acid-ore ratio of (0.4-0.6):1 is required to achieve a lithium leaching rate of ≥90%, while simultaneously controlling the potassium leaching rate at 10%-30% and the aluminum leaching rate at 15%-30%, achieving highly efficient and selective lithium leaching.

[0014] Specifically, during defluorination roasting in a specific temperature range of 700℃-900℃, lepidolite undergoes partial dehydroxylation and interlayer structure relaxation, but its layered silicate framework is not completely destroyed. Lithium ions, due to their small ionic radius and high charge density, migrate from their originally tightly packed octahedral coordination to more easily exchangeable interlayer positions or structural defects during defluorination, forming "activated lithium." Potassium and aluminum ions, on the other hand, remain primarily in their original lattice positions. Therefore, when a medium acid ratio of 0.4-0.6:1 is used for a single acidification, sulfuric acid primarily reacts rapidly with these "activated lithium" ions through ion exchange, achieving over 90% lithium leaching without completely destroying the aluminosilicate framework. Simultaneously, due to the controlled acid ratio, the destruction of the layered structure by sulfuric acid is limited; most potassium remains bound within the mica layers, and aluminum remains within the octahedral sheets, thus controlling the leaching rates of potassium and aluminum to low levels of 10%-30% and 15%-30%, respectively.

[0015] Based on this selective leaching effect, a stepwise acidification strategy is further adopted to separate the extraction of lithium from the recovery of other valuable elements. Lithium is extracted preferentially in the first acidification, while valuable elements such as potassium, aluminum, rubidium, and cesium enriched in the water leaching residue are recovered through secondary acidification leaching or chlorination treatment. This stepwise acidification strategy not only results in a low impurity content in the first acidification leaching solution, simplifying the subsequent lithium purification process and reducing production costs, but also avoids the problem of difficult separation of mixed elements in traditional high-acid one-step leaching. The resulting silica slag has high purity, truly realizing the comprehensive recovery and utilization of multiple elements in lithium mica ore.

[0016] The present invention will be further described below with reference to specific embodiments. Additional aspects and advantages of the invention will be set forth in part in the description which follows, and in part will be obvious from the description, or may be learned by practice. Detailed Implementation

[0017] The present invention will now be clearly and completely described. Those skilled in the art will be able to implement the present invention based on this description. It should be particularly noted that:

[0018] The technical solutions and features provided in the various sections, including the following description, can be combined with each other without conflict. Furthermore, where possible, these technical solutions, features, and related combinations can be given specific technical subject matter and protected by relevant patents.

[0019] The embodiments of the present invention described below are generally only some embodiments and not all embodiments. Based on these embodiments, all other embodiments obtained by those skilled in the art without creative effort should fall within the scope of patent protection.

[0020] Regarding terminology and units in this specification: The term "comprising" and any variations thereof in this specification, the corresponding claims, and related sections are intended to cover a non-exclusive inclusion. Furthermore, other related terms and units can be reasonably interpreted based on the relevant information provided in this specification.

[0021] Example 1

[0022] This embodiment uses lepidolite ore from a certain area in Jiangxi Province as raw material. The main chemical components of the lepidolite ore are (wt%): Li2O 3.2%, K2O 10.5%, Al2O3 18.2%, SiO2 48.5%, F 4.8%, Rb2O 0.35%, Cs2O 0.08%, Fe2O3 2.1%, Na2O 0.8%, CaO 0.5%, MgO 0.6%, MnO 0.15%.

[0023] Process Flow

[0024] The process flow of this embodiment includes the following steps:

[0025] (1) The lithium mica ore was roasted at 800℃ to remove fluorine, and defluorinated lithium mica ore was obtained.

[0026] Lithium mica ore was crushed to a particle size of less than 2 mm and fed into a rotary kiln for roasting and defluorination. The rotary kiln temperature was controlled at 800℃, and the material residence time in the kiln was 2 hours. During roasting, the fluorine in the lepidolite was mainly removed by volatilization in the form of HF. The roasted defluorinated lepidolite ore exhibited a loose and porous structure, and its color changed from the grayish-white of the original ore to a light pink.

[0027] (2) The defluorinated lithium mica ore is subjected to a first acid leaching process to obtain a first acid leaching solution and a water leaching residue.

[0028] The first acid leaching process includes: mixing defluorinated lithium mica ore with concentrated sulfuric acid for acidification, with a mass ratio of concentrated sulfuric acid to defluorinated lithium mica ore of 0.5:1, followed by slurry conditioning and leaching.

[0029] The specific procedure is as follows: Add defluorinated lithium mica ore to the acidification reactor, and slowly add 98wt% concentrated sulfuric acid, controlling the acid addition rate to avoid an overly vigorous reaction. During the mixing process, the material temperature naturally rises to approximately 150℃, and acidification is maintained at this temperature for 1 hour. The acidified material is a dark brown paste.

[0030] The acidified material was transferred to a leaching tank, and water was added at a liquid-to-solid ratio of 4:1 to prepare the slurry. The leaching temperature was controlled at 80℃, and the leaching was carried out with stirring for 2 hours. Samples were taken periodically during the leaching process to analyze the leaching status of each element.

[0031] The test results were as follows: lithium leaching rate 92%, sodium leaching rate 60%, potassium leaching rate 20%, aluminum leaching rate 22%, rubidium leaching rate 18%, cesium leaching rate 15%, calcium leaching rate 8%, magnesium leaching rate 12%, manganese leaching rate 70%, and iron leaching rate 30%.

[0032] (3) Separate the primary acid leaching solution from the water leaching residue, and further process the separated lithium-containing leaching solution to obtain battery-grade lithium carbonate products.

[0033] After leaching, a plate and frame filter press is used for solid-liquid separation to obtain primary acidified leachate (lithium-containing leachate) and water-leached residue.

[0034] The first acidification leachate was light green and transparent, with a pH of approximately 1.5. The main component analysis results were: Li + 10.5g / L, K + 8.2g / L, Al 3+ 12.5g / L, Na + 2.1 g / L, Fe 3+ 1.8g / L, Mn 2+ 0.35 g / L. The water-leached residue has a moisture content of about 25% and its main components are unreacted aluminosilicates, which still contain a large amount of unleached valuable elements such as potassium, aluminum, rubidium, and cesium.

[0035] Battery-grade lithium carbonate products obtained through further processing of lithium-containing leachates include:

[0036] Alum formation: By controlling temperature and concentration conditions, aluminum, potassium, and sulfate ions in the lithium-containing leachate are reacted to form potassium alum precipitate. The potassium alum is then separated by filtration, reducing the potassium and aluminum content in the leachate. Specifically, the lithium-containing leachate is heated to 60°C. Utilizing the existing aluminum and potassium in the solution, the mixture is stirred for 2 hours and then cooled to 25°C. Potassium alum crystallizes out and is then separated by filtration.

[0037] Impurity removal: Sodium hydroxide is added to the alum-forming solution to adjust the pH to 11-12, causing impurities such as iron, aluminum, calcium, and magnesium to precipitate as hydroxides. The purified solution is then obtained by filtration. Specifically, sodium hydroxide solution is added to the alum-forming solution, and the mixture is stirred at 60°C for 1 hour. The purified solution is then obtained by filtration.

[0038] MVR evaporation and concentration: The purified solution is concentrated by MVR evaporation to achieve a lithium-ion concentration of 20-30 g / L. Specifically, the purified solution is fed into an MVR evaporator for evaporation and concentration to obtain a concentrated solution containing Li... + The concentration reached 25 g / L.

[0039] Lithium precipitation: Sodium carbonate is added to the concentrate, and the reaction is carried out at 90℃-95℃ to form lithium carbonate precipitate. After filtration, washing, and drying, battery-grade lithium carbonate product is obtained. The specific operation is as follows: the concentrate is heated to 92℃, anhydrous sodium carbonate is added at 1.05 times the theoretical amount of Li2CO3, the reaction is stirred for 2 hours, filtered, washed 3 times, and dried at 180℃ for 4 hours to obtain battery-grade lithium carbonate product with a Li2CO3 purity of 99.6wt%, which meets the battery-grade lithium carbonate standard.

[0040] (4) The water-leached residue is subjected to secondary acid leaching to obtain secondary acid leaching solution and silica residue.

[0041] The water-leached residue was acidified with 98wt% concentrated sulfuric acid at a mass ratio of 1:0.8, under the same acidification conditions as the first acidification leaching. After acidification, the slurry was leached.

[0042] Secondary acid leaching allows most metal elements to enter the secondary acid leaching solution. The leaching rates of each element (based on defluorinated lithium mica ore) were as follows: potassium 85%, aluminum 88%, rubidium 82%, cesium 80%, and residual lithium 98%.

[0043] (5) Separate the secondary acidification leachate from the silica slag, and then further recover valuable products from the secondary acidification leachate.

[0044] The secondary acidification leaching slurry was subjected to solid-liquid separation using a plate and frame filter press to obtain 380L of secondary acidification leaching solution and 35kg of silica slag. The silica slag obtained by separating the secondary acidification leaching solution and silica slag has a SiO2 content of 88wt% and can be sold as a building material raw material.

[0045] Valuable products recovered from secondary acidification leachate include:

[0046] Alum formation: By controlling the temperature and concentration conditions, aluminum, potassium, rubidium, and cesium in the secondary acidification leachate are reacted with sulfate ions to form potassium alum and rubidium-cesium alum, which are then filtered to obtain the alum product. The specific operation is as follows: the secondary acidification leachate is concentrated, heated to 80°C, an appropriate amount of sulfuric acid is added, and after stirring for 2 hours, it is slowly cooled to 30°C to obtain a mixed alum precipitate.

[0047] Calcination and Leaching: Alum products are calcined and decomposed, and the calcined product is separated by water leaching to obtain alumina. Specifically, the alum product is calcined at 850℃ for 3 hours, decomposing the alum into alumina and sulfate. The calcined product is then leached with water at 80℃, and alumina is obtained through solid-liquid separation.

[0048] Cooling crystallization: Potassium sulfate is obtained from the roasting leachate by cooling crystallization. The specific operation is as follows: the roasting leachate is concentrated and then cooled to 10°C, potassium sulfate crystals out, and the product is obtained by filtration.

[0049] Fractional extraction: Rubidium and cesium are recovered separately from the cooled crystallization mother liquor by fractional extraction.

[0050] The technical solution in Example 1 breaks through the conventional technical approach that "high lithium leaching rate requires high acid content, and high acid content inevitably leads to high impurities." By defluorinating and roasting lepidolite in a specific temperature range of 700℃-900℃, fluorine is effectively removed to avoid equipment corrosion, while maintaining the loose and porous structure of the defluorinated lepidolite, significantly improving its reactivity. The crystal structure of the lepidolite after defluorination treatment in this temperature range is optimized. In a single acid leaching process, a medium acid-to-ore ratio of (0.4-0.6):1 is required to achieve a lithium leaching rate of ≥90%, while controlling the potassium leaching rate at 10%-30% and the aluminum leaching rate at 15%-30%, achieving highly efficient and selective lithium leaching.

[0051] Specifically, during defluorination roasting in a specific temperature range of 700℃-900℃, lepidolite undergoes partial dehydroxylation and interlayer structure relaxation, but its layered silicate framework is not completely destroyed. Lithium ions, due to their small ionic radius and high charge density, migrate from their originally tightly packed octahedral coordination to more easily exchangeable interlayer positions or structural defects during defluorination, forming "activated lithium." Potassium and aluminum ions, on the other hand, remain primarily in their original lattice positions. Therefore, when a medium acid ratio of 0.4-0.6:1 is used for a single acidification, sulfuric acid primarily reacts rapidly with these "activated lithium" ions through ion exchange, achieving over 90% lithium leaching without completely destroying the aluminosilicate framework. Simultaneously, due to the controlled acid ratio, the destruction of the layered structure by sulfuric acid is limited; most potassium remains bound within the mica layers, and aluminum remains within the octahedral sheets, thus controlling the leaching rates of potassium and aluminum to low levels of 10%-30% and 15%-30%, respectively.

[0052] Based on this selective leaching effect, a stepwise acidification strategy is further adopted to separate the extraction of lithium from the recovery of other valuable elements. Lithium is extracted preferentially in the first acidification, while valuable elements such as potassium, aluminum, rubidium, and cesium enriched in the water leaching residue are recovered through secondary acidification leaching or chlorination treatment. This stepwise acidification strategy not only results in a low impurity content in the first acidification leaching solution, simplifying the subsequent lithium purification process and reducing production costs, but also avoids the problem of difficult separation of mixed elements in traditional high-acid one-step leaching. The resulting silica slag has high purity, truly realizing the comprehensive recovery and utilization of multiple elements in lithium mica ore.

[0053] Example 2

[0054] This embodiment uses the same lithium mica ore from a certain place in Jiangxi as the raw material as in Example 1, and the main chemical composition is the same as in Example 1.

[0055] The process flow is basically the same as in Example 1, except that:

[0056] (1) The lithium mica ore was roasted at 700℃ to remove fluorine, and defluorinated lithium mica ore was obtained.

[0057] Lithium mica ore was crushed to a particle size of less than 2 mm and then fed into a rotary kiln for roasting and defluorination. The rotary kiln temperature was controlled at 700℃, and the material residence time in the kiln was 2.5 hours. The roasted defluorinated lithium mica ore was in a loose state, and its color changed from the grayish-white of the original ore to light yellow. Testing showed a defluorination rate of 89%.

[0058] (2) The defluorinated lithium mica ore was subjected to a single acid leaching process. The mass ratio of concentrated sulfuric acid to defluorinated lithium mica ore was 0.5:1, the acidification temperature was 150℃, and the acidification time was 1h. After acidification, water was added to adjust the slurry at a liquid-to-solid ratio of 4:1, and leaching was carried out at 80℃ for 2h.

[0059] The test results were as follows: lithium leaching rate 90%, sodium leaching rate 58%, potassium leaching rate 26%, aluminum leaching rate 28%, rubidium leaching rate 25%, cesium leaching rate 23%, calcium leaching rate 9%, magnesium leaching rate 14%, manganese leaching rate 66%, and iron leaching rate 36%.

[0060] Subsequent steps (3)-(5) are the same as in Example 1. A battery-grade lithium carbonate product with a Li2CO3 purity of 99.3 wt% was finally obtained. The SiO2 content in the silica slag was 85 wt%.

[0061] Example 3

[0062] This embodiment uses the same lithium mica ore from a certain place in Jiangxi as the raw material as in Embodiment 1.

[0063] The process flow is basically the same as in Example 1, except that:

[0064] (1) The lepidolite was roasted at 900℃ to remove fluoride. The rotary kiln temperature was controlled at 900℃, and the material residence time in the kiln was 1.5 hours. The roasted fluorinated lepidolite turned pink. The defluorination rate was tested to be 96%.

[0065] (2) The defluorinated lithium mica ore was subjected to a single acid leaching process, with a mass ratio of concentrated sulfuric acid to defluorinated lithium mica ore of 0.5:1. The acidification and leaching conditions were the same as in Example 1.

[0066] The test results were as follows: lithium leaching rate 91%, sodium leaching rate 61%, potassium leaching rate 24%, aluminum leaching rate 26%, rubidium leaching rate 22%, cesium leaching rate 20%, calcium leaching rate 8%, magnesium leaching rate 13%, manganese leaching rate 69%, and iron leaching rate 34%.

[0067] Subsequent processing was the same as in Example 1. The final product obtained was battery-grade lithium carbonate with a Li₂CO₃ purity of 99.4 wt%. The SiO₂ content in the silica slag was 87 wt%.

[0068] Example 4

[0069] This embodiment uses the same lithium mica ore from a certain place in Jiangxi as the raw material as in Embodiment 1.

[0070] The process flow is basically the same as in Example 1, except that:

[0071] (1) The roasting temperature is 800℃, the same as in Example 1.

[0072] (2) The defluorinated lepidolite was subjected to a single acid leaching process, with a mass ratio of concentrated sulfuric acid to defluorinated lepidolite of 0.4:1. The acidification temperature was 145℃, and the acidification time was 1.5h. The slurry preparation and leaching conditions were the same as in Example 1.

[0073] The test results were as follows: lithium leaching rate 90%, sodium leaching rate 52%, potassium leaching rate 15%, aluminum leaching rate 18%, rubidium leaching rate 16%, cesium leaching rate 14%, calcium leaching rate 7%, magnesium leaching rate 11%, manganese leaching rate 63%, and iron leaching rate 26%.

[0074] Subsequent processing was the same as in Example 1. The final product obtained was battery-grade lithium carbonate with a Li₂CO₃ purity of 99.5 wt%. The SiO₂ content in the silica slag was 88 wt%.

[0075] Example 5

[0076] This embodiment uses the same lithium mica ore from a certain place in Jiangxi as the raw material as in Embodiment 1.

[0077] The process flow is basically the same as in Example 1, except that:

[0078] (1) The roasting temperature is 800℃, the same as in Example 1.

[0079] (2) The defluorinated lepidolite was subjected to a single acid leaching process, with a mass ratio of concentrated sulfuric acid to defluorinated lepidolite of 0.6:1. The acidification temperature was 155℃, and the acidification time was 1 hour. The slurry preparation and leaching conditions were the same as in Example 1.

[0080] The test results were as follows: lithium leaching rate 93%, sodium leaching rate 65%, potassium leaching rate 28%, aluminum leaching rate 29%, rubidium leaching rate 27%, cesium leaching rate 25%, calcium leaching rate 10%, magnesium leaching rate 15%, manganese leaching rate 74%, and iron leaching rate 38%.

[0081] In subsequent processing, due to the slightly higher impurity content, the amount of sodium hydroxide used in the impurity removal process was increased by approximately 15%. The final product obtained was battery-grade lithium carbonate with a Li₂CO₃ purity of 99.4 wt%. The SiO₂ content in the silica slag was 85 wt%.

[0082] Example 6 (Chlorination Treatment Route)

[0083] This embodiment uses the same lithium mica ore from a certain place in Jiangxi as the raw material as in Embodiment 1.

[0084] Steps (1)-(3) are the same as in Example 1, and water-leached residue is obtained.

[0085] (4) Chlorination treatment of the water-leached residue. The water-leached residue and calcium chloride were mixed at a mass ratio of 1:0.3 and chlorinated and roasted at 1200℃ for 2 hours. During the chlorination and roasting process, the potassium, aluminum, rubidium, cesium and other metal elements in the water-leached residue were converted into chlorides, and the chlorides of iron and aluminum were volatilized in gaseous form and recovered through a condensation system.

[0086] (5) The chlorinated roasting product was leached with water at 60°C with a liquid-to-solid ratio of 3:1 for 1 hour. Filtration separated the product to obtain a chloride solution containing potassium, rubidium, and cesium, and silica slag. The silica slag contained 90 wt% SiO2 and could be sold as a high-purity silicon source. Potassium chloride, rubidium chloride, and cesium chloride were recovered from the chloride solution by stepwise crystallization and extraction.

[0087] Comparative Example 1

[0088] This comparative example uses the same lithium mica ore from a certain place in Jiangxi Province as the raw material as in Example 1.

[0089] The process flow includes the following steps:

[0090] (1) Lithium mica ore was roasted at 1050℃ to remove fluoride. The rotary kiln temperature was controlled at 1050℃, and the material residence time in the kiln was 1.5h. After roasting, the material sintered, the particles hardened, and the color turned dark red. The defluorination rate was tested to be 99%.

[0091] (2) Acid leaching of defluorinated lithium mica ore was carried out, with a mass ratio of concentrated sulfuric acid to defluorinated lithium mica ore of 1.1:1. The sintered material was crushed and then acidified at a temperature of 170℃ for 2.5 hours. After acidification, water was added at a liquid-to-solid ratio of 4:1 to form a slurry, and leaching was carried out at 85℃ for 3 hours.

[0092] The test results are as follows: lithium leaching rate 92%, sodium leaching rate 82%, potassium leaching rate 78%, aluminum leaching rate 76%, rubidium leaching rate 74%, cesium leaching rate 72%, calcium leaching rate 42%, magnesium leaching rate 48%, manganese leaching rate 85%, and iron leaching rate 70%.

[0093] (3) Due to the high impurity content in the leachate, multiple impurity removal steps are required: first, the pH is adjusted to 4.5 to remove iron and aluminum, then sodium carbonate is added to remove calcium and magnesium, and finally potassium is removed by ion exchange. The impurity removal process consumes a large amount of reagents, and the amount of waste residue generated is about three times that of Example 1. After a complex purification process, a lithium carbonate product with a purity of 99.1 wt% Li2CO3 is finally obtained, with a total lithium recovery rate of 75%.

[0094] Because the one-step leaching process mixes all elements, valuable elements such as potassium, aluminum, rubidium, and cesium are dispersed in the various impurity removal residues, making them difficult to recover. The SiO2 content in the silica residue is only 72 wt%.

[0095] Comparative Example 2

[0096] This comparative example uses the same lithium mica ore from a certain place in Jiangxi Province as the raw material as in Example 1.

[0097] The process flow includes the following steps:

[0098] (1) Lithium mica ore was roasted at 650℃ to remove fluoride. The rotary kiln temperature was controlled at 650℃, and the material residence time in the kiln was 3 hours. The appearance of the roasted lithium mica ore changed little. The defluorination rate was 72%, and the residual fluorine content was 1.34%.

[0099] (2) The defluorinated lepidolite was subjected to a single acid leaching process, with a mass ratio of concentrated sulfuric acid to defluorinated lepidolite of 0.5:1. HF gas was released during the acid leaching process.

[0100] The test results were as follows: lithium leaching rate 78%, sodium leaching rate 48%, potassium leaching rate 20%, and aluminum leaching rate 22%. The leaching effect was unsatisfactory due to incomplete defluorination.

[0101] (3) The high fluorine content in the leachate affected subsequent processing. The final purity of Li2CO3 was 98.6 wt%, which did not meet the battery-grade standard.

[0102] Comparative Example 3

[0103] This comparative example uses the same lithium mica ore from a certain place in Jiangxi Province as the raw material as in Example 1.

[0104] The process flow is basically the same as in Example 1, except that:

[0105] (1) The roasting temperature is 800℃, the same as in Example 1.

[0106] (2) The defluorinated lithium mica ore is subjected to a single acid leaching process, with the mass ratio of concentrated sulfuric acid to defluorinated lithium mica ore being 0.3:1 (below the scope of this invention).

[0107] The acidification reaction was incomplete, and the test results were as follows: lithium leaching rate 82%, sodium leaching rate 45%, potassium leaching rate 10%, and aluminum leaching rate 14%.

[0108] The lithium leaching rate did not reach 90%, which does not meet the requirements of industrial production. To improve the lithium leaching rate, additional lithium needs to be leached during secondary acidification, which increases the complexity and cost of the process.

[0109] The present invention has been described above. Those skilled in the art will be able to implement the present invention based on these descriptions. All other embodiments obtained by those skilled in the art based on the foregoing content of this specification without inventive effort should fall within the scope of patent protection.

Claims

1. A method for stepwise acidification lithium extraction and multi-element comprehensive recovery from lepidolite ore, characterized in that, include: (1) The lithium mica ore was roasted at 700℃-900℃ to remove fluoride, and defluorinated lithium mica ore was obtained. (2) The defluorinated lithium mica ore is subjected to a single acid leaching process to obtain a single acid leaching solution and a water leaching residue. The single acid leaching process can achieve the following results: lithium leaching rate ≥90%, sodium leaching rate 50%-70%, potassium leaching rate 10%-30%, and aluminum leaching rate 15%-30%. (3) Separate the primary acid leaching solution from the water leaching residue, and further process the separated lithium-containing leaching solution to obtain battery-grade lithium carbonate products. (4) The water-leached residue is subjected to secondary acid leaching or chlorination treatment. Secondary acid leaching yields secondary acid leaching solution and silica slag. Secondary acid leaching allows most of the metal elements to enter the secondary acid leaching solution; or chlorination treatment yields metal chlorides and silica slag. (5) Separate the secondary acid leaching solution from the silica slag, or separate the metal chloride from the silica slag, and then further recover valuable products from the secondary acid leaching solution or the metal chloride. The first acid leaching process includes: mixing defluorinated lithium mica ore with concentrated sulfuric acid for acidification, with the mass ratio of concentrated sulfuric acid to defluorinated lithium mica ore being (0.4-0.6):1, followed by slurry leaching.

2. The method for stepwise acidification lithium extraction and multi-element comprehensive recovery of lepidolite ore as described in claim 1, characterized in that: A single acid leaching process also results in the following leaching rates: rubidium leaching rate of 10%-30%, cesium leaching rate of 10%-30%, calcium leaching rate of ≤10%, magnesium leaching rate of ≤15%, manganese leaching rate of 60%-80%, and iron leaching rate of 20%-40%.

3. The method for stepwise acidification lithium extraction and multi-element comprehensive recovery of lithium mica ore as described in claim 1, characterized in that: Secondary acid leaching includes: mixing the water-leached residue with concentrated sulfuric acid for acidification, with a mass ratio of concentrated sulfuric acid to water-leached residue of (0.6-1.2):1, followed by slurry preparation and leaching.

4. The method for stepwise acid extraction of lithium from lepidolite ore and comprehensive recovery of multiple elements as described in claim 1, characterized in that: The chlorination process includes: chlorinating and roasting the water-leached residue at a temperature below 1400°C to chlorinate the metal elements in the residue other than silicon. During the chlorination roasting, ferric chloride and aluminum chloride escape as gases. After the chlorination roasting, the remaining chlorides are mixed with silicon dioxide. Subsequently, these chlorides are separated from the silicon dioxide by adding water leaching, and then the chlorides are further purified.

5. The method for stepwise acid extraction of lithium from lepidolite ore and comprehensive recovery of multiple elements as described in claim 1, characterized in that: The chlorination process includes: chlorinating and roasting the water-leached residue at a temperature of ≥1400℃ to chlorinate silicon and other metal elements in the residue. During the chlorination and roasting, silicon tetrachloride, ferric chloride, and aluminum chloride escape in gaseous form and are then cooled and recovered in stages using the different sublimation temperatures of silicon tetrachloride, ferric chloride, and aluminum chloride.

6. The method for stepwise acid extraction of lithium from lepidolite ore and comprehensive recovery of multiple elements as described in claim 1, characterized in that: The silica slag obtained by separating the secondary acidification leachate from the silica slag has a SiO2 content of ≥85wt%.

7. The method for stepwise acidification and lithium extraction from lepidolite ore and comprehensive recovery of multiple elements as described in claim 1, characterized in that: Valuable products recovered from secondary acidification leachate include: Alum formation: By controlling the temperature and concentration conditions, aluminum, potassium, rubidium, and cesium in the secondary acidification leachate react with sulfate ions to form potassium alum and rubidium-cesium alum, which are then filtered to obtain the alum product; Calcination and leaching: Alum products are decomposed by calcination, and the calcined products are separated by water leaching to obtain aluminum oxide; Cooling crystallization: Potassium sulfate is obtained from the roasting leachate by cooling crystallization; Fractional extraction: Rubidium and cesium are recovered separately from the cooled crystallization mother liquor by fractional extraction.

8. The method for stepwise acid extraction of lithium from lepidolite ore and comprehensive recovery of multiple elements as described in claim 1, characterized in that: Battery-grade lithium carbonate products obtained through further processing of lithium-containing leachates include: Alum formation: By controlling temperature and concentration conditions, aluminum, potassium and sulfate ions in lithium-containing leachate are precipitated to form potassium alum, which is then separated by filtration and the potassium and aluminum content in the leachate is reduced. Impurity removal: Sodium hydroxide is added to the alum-forming solution to adjust the pH to 11-12, so that impurities such as iron, aluminum, calcium, and magnesium form hydroxide precipitates, and the purified solution is obtained by filtration. MVR evaporation and concentration: The purified liquid is concentrated by MVR evaporation to achieve a lithium ion concentration of 20-30 g / L; Lithium precipitation: Sodium carbonate is added to the concentrate and reacted at 90℃-95℃ to form lithium carbonate precipitate. After filtration, washing and drying, battery-grade lithium carbonate product is obtained.

9. The method for stepwise acid extraction of lithium from lepidolite ore and comprehensive recovery of multiple elements as described in claim 1, characterized in that: The calcination defluorination was carried out at 750℃-850℃.

10. The method for stepwise acidification lithium extraction and multi-element comprehensive recovery of lepidolite ore as described in claim 1, characterized in that: During the first acid leaching, the mass ratio of concentrated sulfuric acid to defluorinated lithium mica ore is (0.45-0.55):1.