Method for comprehensively extracting lithium from lepidolite and recovering by-products

Abstract

The invention discloses a method for comprehensively extracting lithium from lepidolite and recovering by-products, and relates to the technical field of comprehensive utilization of lepidolite resources, and the method comprises the following steps: step 1, burdening and conversion roasting; secondly, water leaching and fluorite flotation are conducted; 3, acidizing roasting and secondary leaching, wherein sulfuric acid with the concentration larger than or equal to 98 wt% is added into tailings obtained in the step 2 for acidizing roasting; carrying out secondary leaching on an acidified roasting product by using the filtrate obtained in the step 2; step 4, performing slag-liquid separation and preparing a lithium product; and 5, preparing potassium sulfate and mirabilite. According to the invention, limestone ore blending conversion roasting fluorine fixation is carried out, so that the generation of a large amount of hydrogen fluoride is avoided from the source, fluorine is converted into a high-value product instead of a pollutant through fluorite flotation, lithium precipitation carbon sources are all from primary roasting limestone and fuel gas containing carbon, and final waste residues leached after acidification roasting are used as qualified cement raw materials; and zero emission of solid waste is realized, and the environment-friendly bottleneck of the industry of extracting lithium from lepidolite is thoroughly solved.

Description

Technical Field

[0001] This invention relates to the field of comprehensive utilization technology of lepidolite resources, specifically to a method for comprehensively extracting lithium from lepidolite and recovering by-products. Background Technology

[0002] Lithium, as an energy metal of the 21st century, plays an irreplaceable role in new energy vehicles and energy storage. Lepidolite is one of my country's important lithium resources, but its low grade and complex composition, especially its high fluorine content, make extraction difficult and cause serious environmental pollution. Although the traditional sulfuric acid roasting method for lepidolite has been used for a long time, it generates large amounts of toxic and harmful gases such as hydrogen fluoride and sulfur dioxide during the roasting process, causing serious corrosion and pollution to equipment and the environment. The roasting residue contains soluble fluorides and is classified as hazardous solid waste, requiring storage in warehouses, which poses significant environmental risks and land occupation costs. Furthermore, this process has low or no recovery efficiency for valuable elements associated with lepidolite, such as potassium, rubidium, cesium, and fluorine, resulting in resource waste.

[0003] Currently, most treatments for lepidolite employ either limestone sintering or sulfation roasting. Limestone sintering, through high-temperature roasting with added limestone, fixes fluorine in the slag. While this method reduces harmful gases, the slag after fluorine fixation, due to its complex mineral phases and poor activity, is difficult to use on a large scale in cement production and remains primarily as solid waste. Furthermore, it fails to effectively recover the fixed fluorine as a high-value-added fluorite product. Sulfation roasting, on the other hand, adds a large amount of sodium during roasting, making the slag a high-sodium, fluorine-containing waste that cannot be used as a cement raw material. Large quantities of this high-sodium, fluorine-containing slag are simply discarded and stockpiled. Simultaneously, as potassium sulfate is leached, it forms a potassium-sodium mixed salt with a large amount of sodium sulfate, significantly reducing the quality of the potassium salt. This mixed salt is difficult to sell as a product and has low economic value. Summary of the Invention

[0004] The purpose of this invention is to provide a method for comprehensively extracting lithium from lepidolite and recovering by-products, so as to solve the problems mentioned in the background art.

[0005] To achieve the above objectives, the present invention provides the following technical solution: a method for comprehensively extracting lithium from lepidolite and recovering by-products, comprising the following steps: Step 1, Ingredient preparation and conversion roasting: Lithium mica concentrate and calcium-containing materials are mixed at a fluorine to calcium oxide molar ratio of 1:1 to 1:2, and then converted and roasted at 850℃ to 950℃ to convert fluorine into calcium fluoride and lithium into lithium hydroxide, yielding roasted clinker and CO2-containing flue gas. The CO2 flue gas is then reacted with potassium sodium hydroxide solution after dust removal to generate potassium sodium carbonate solution, which is used for subsequent lithium precipitation. After crushing and grinding, the roasted clinker is controlled to have more than 80% particles smaller than 200 mesh. Step 2, water immersion and fluorite flotation: The roasted clinker obtained in step one is leached with water at a solid-liquid ratio of 1:1.5 to 1:4. Fluorite (CaF2) flotation reagent is added, and the water-leached slurry is subjected to a flotation process of "primary roughing, multi-stage cleaning, and multi-stage scavenging" to obtain fluorite concentrate with a CaF2 grade >90% and a yield >85%. The slurry after flotation is filtered to obtain filtrate and tailings. The filtrate is returned to the leaching process after roasting. Step 3, acidification roasting and secondary leaching: Add sulfuric acid with a concentration of ≥98 wt% to the tailings obtained in step two, and control the amount of sulfuric acid added to be 1.0 to 1.4 times the theoretical molar amount of Li+K in the concentrate. Acid roasting is carried out at 200℃ to 300℃. The acid roasting product is leached a second time with the filtrate obtained in step two to convert the residual lithium, potassium and aluminum into soluble sulfates. Step 4, Slag-liquid separation and lithium product preparation: The secondary leaching slurry from step three is treated with lime to remove iron and aluminum. After filtration, iron-containing silicon-aluminum tailings and a mixed solution of lithium sulfate and potassium sodium sulfate are obtained. The tailings are used as cement raw materials. The mixed solution is evaporated and concentrated to a water content of ≥80%. After deep removal of calcium and magnesium by resin, the potassium sodium carbonate solution obtained in step one is used for carbonation and lithium precipitation to obtain lithium carbonate products and lithium precipitation mother liquor. Step 5, Preparation of potassium sulfate and sodium sulfate: The lithium precipitation mother liquor from step four is subjected to adsorption to recover residual lithium, and rubidium and cesium are extracted to recover rubidium and cesium, resulting in a potassium-rich solution. This potassium-rich solution is mixed with the subsequently returned evaporation mother liquor and then subjected to 0°C freeze crystallization to precipitate potassium-sulfuric acid mixed salt. A certain amount of 0-10°C low-temperature water is added to the mixed salt to wash away K2SO4, yielding pure decahydrate sodium sulfate. The freeze crystallization mother liquor is mixed with potassium-containing elution mother liquor and subsequently returned sodium-containing elution mother liquor and then fed into the evaporation crystallization system to crystallize potassium-sulfuric acid compound salt. The potassium-sulfuric acid compound salt is washed with room temperature water to remove all sodium sulfate and some potassium sulfate, yielding pure K2SO4 product. The sodium-containing elution mother liquor is returned to the fore-liquid recycling system of the aforementioned potassium-sulfuric acid compound salt evaporation crystallization system. All evaporation condensate is recycled for steps two and three. The system does not introduce sodium salt, and all sodium sulfate produced comes from lithium concentrate.

[0006] Further, in step one, the lepidolite concentrate contains Li2O: 1.5-2.5 wt%, F: 3.0-4.5 wt%, Na: 0.5-0.7 wt%, and the calcium-containing material is limestone or quicklime; During the conversion roasting, the roasting time is 45 to 90 minutes.

[0007] Further, in step two, the flotation reagent system is as follows: Roughing stage: collector 300-1000 g / t, inhibitor 300-1500 g / t, frother 15-130 g / t; Fine selection stage: 5-80 g / t of collector, 5-90 g / t of inhibitor, 0-10 g / t of foaming agent, and ≥5 fine selection times; Sweeping stage: The dosage of each reagent is 10% to 50% of that used in the roughing stage, and the number of sweeping operations is ≥1. The collector is selected from one or more of oleic acid, sodium oleate, and oxidized paraffin soap; the inhibitor is selected from one or two of water glass and sodium carbonate; and the foaming agent is selected from one or two of pine oil and No. 2 oil.

[0008] Furthermore, in step three, the acidification roasting equipment is a rotary kiln or a circulating fluidized bed furnace, and the roasting residence time is 30 to 90 minutes; the small amount of HF and SO2 escaping from the roasting tail gas is desulfurized by lime method, and the resulting calcium slag is returned to the flotation system in step two.

[0009] Furthermore, in step three, during the secondary leaching, the solid-liquid ratio before leaching is 1.3:1.5 to 1.2:4, so that the residual lithium is converted into lithium sulfate, which enters the solution together with the potassium sulfate and aluminum sulfate formed at the same time. After leaching, the solid-liquid ratio of the slurry is 0.9:1.9 to 0.8:4.4.

[0010] Further, in step four, the amount of lime added is 1.0 to 1.2 times the amount of Fe in the concentrate; the evaporation and concentration endpoint is controlled to have a solution water content ≥80%; after removing calcium and magnesium from the resin, Ca... 2+ <20 mg / L, Mg 2+ <10 mg / L.

[0011] Further, in step four, the amount of potassium sodium carbonate solution added is 1.0 to 1.2 times the Li content in the concentrate, and the reaction is carried out at 80 to 100°C and at a final pH of 10 to 11 for 20 to 60 minutes, and then filtered to obtain Li2CO3 precipitate.

[0012] Furthermore, in step five, the rubidium and cesium extractant is 4-tert-butyl-2-(α-methylbenzyl)phenol, and a three-stage countercurrent extraction is performed, with a total recovery rate of rubidium and cesium ≥90%; the K2SO4 crystallization temperature is controlled at 25-60℃, and the crystallization rate is ≥80%.

[0013] Furthermore, in step five, a portion of the potassium-sodium sulfate mixed mother liquor reacts with lime to produce potassium-sodium hydroxide and calcium sulfate. Part of the potassium-sodium hydroxide is used to absorb the roasting flue gas in step one to produce potassium-sodium carbonate, and the other part is used as a pH adjuster. The calcium sulfate is washed multiple times to obtain calcium sulfate waste residue. The calcium sulfate waste residue and calcium silicate slag are combined and sold as cement raw materials, realizing a dual-loop cycle of potassium and calcium and zero discharge of waste residue.

[0014] The technical effects and advantages provided by the present invention in the above technical solution are as follows: 1. By converting limestone ore into solid fluorine through roasting, the generation of large amounts of hydrogen fluoride is eliminated from the source. Fluorite flotation converts fluorine into high-value products rather than pollutants. The carbon source for lithium precipitation comes entirely from the carbon-containing limestone and fuel gas from primary roasting. Acid roasting can be carried out using an electric rotary kiln, resulting in almost no carbon emissions. The final waste residue after acid roasting is used as a qualified cement raw material, achieving zero solid waste discharge and completely solving the environmental bottleneck in the lithium extraction industry from lepidolite. 2. This process not only efficiently extracts lithium, but also simultaneously recovers high-grade fluorite (CaF2) and high-purity potassium sulfate and mirabilite, achieving comprehensive recovery of valuable elements such as lithium, fluorine, potassium, sodium, rubidium, and cesium from lepidolite. Gypsum and aluminosilicate slag are used as cement raw materials, turning waste into treasure and greatly improving resource utilization value and economic benefits. 3. Using the leachate from the conversion roasting as the leachate from the acid roasting not only utilizes the alkalinity of the primary leachate to neutralize the excess acid used in the secondary roasting, but also fully recovers the residual lithium in the primary roasting residue, thereby improving the overall lithium recovery rate and simplifying the subsequent neutralization and impurity removal steps. The primary roasting flue gas washing and CO2 removal also produces potassium sodium carbonate as a byproduct for lithium precipitation, making full use of the carbon content of the system. 4. Potassium hydroxide is produced by using the system's own potassium sodium sulfate mother liquor and calcium hydroxide. Gypsum is combined with silicon calcium slag. Rubidium and cesium elements are separated, extracted and refined before potassium extraction. The by-product potassium sulfate and sodium sulfate decahydrate are completely separated. The quality is much better than the unseparated potassium sodium mixed salt produced by the sulfate roasting process. Attached Figure Description

[0015] To more clearly illustrate the technical solutions in the embodiments of this application or the prior art, the drawings used in the embodiments will be briefly introduced below. Obviously, the drawings described below are only some embodiments recorded in this invention. For those skilled in the art, other drawings can be obtained based on these drawings.

[0016] Figure 1 This is a schematic diagram of the process flow for the comprehensive lithium extraction and by-product recovery from lepidolite according to the present invention. Detailed Implementation

[0017] To enable those skilled in the art to better understand the technical solutions of the present invention, the present invention will be further described below through specific embodiments, but the scope of protection of the present invention is not limited by the embodiments.

[0018] Example 1: Take 1000 kg of lithium mica concentrate from a certain place, containing 2.5 wt% Li2O, 3.0 wt% F, 5.0 wt% K2O and 0.5 wt% Na2O.

[0019] Step 1, Batching and Transformation Roasting: Based on the fluorine content, 116 kg of limestone powder is added in excess at a molar ratio of 1.1, and mixed evenly with lepidolite concentrate. The mixture is roasted in a rotary kiln at 900℃ for 90 minutes. After crushing and grinding, the roasted clinker is controlled to have more than 80% of the particles below 200 mesh. After dust removal, the flue gas is reacted with a mixture of KOH and NaOH to generate a potassium sodium carbonate solution for later use.

[0020] Step 2, water immersion and fluorite flotation: Add 3m to the roasted clinker at a solid-liquid ratio of 1:3 3 Water leaching was performed with stirring at 60°C for 1 hour, followed by flotation according to the flotation reagent regime. Crude selection: 700 g / t oleic acid, 1000 g / t water glass, 80 g / t pine oil; Selected 5 times, oleic acid 60 g / t, water glass 60 g / t, pine oil 4 g / t; Two sweeping processes were performed, with the reagent dosage being 30% of that used in the roughing process. Approximately 62 kg of fluorite concentrate was obtained, with a CaF2 grade of 96% and a recovery rate of 97%; the filtrate was returned for leaching.

[0021] Step 3: Acidification roasting and secondary leaching; After filtration, the flotation tailings were mixed with 150 kg of 98% concentrated sulfuric acid (1.1 times the theoretical amount of Li+K), and then acidified and roasted at 250°C for 1.5 hours. Then, the water-leached flotation tailings (approximately 2.8 m³) obtained in step two were used as a buffer. 3 The acidified roasted material was subjected to a second leaching at 80℃ for 1 hour. The second leaching slurry was filtered to obtain approximately 2.7m³. 3 The mixture contains a lithium-potassium solution and approximately 650 kg of cement raw material slag.

[0022] Step 4: Slag-liquid separation and lithium preparation Lime and sodium potassium carbonate slurry were added to the secondary leaching solution in step three to separate the iron slag, magnesium slag, and calcium slag by filtration. The filtrate was then evaporated and concentrated to a water content of 80%. After deep removal of calcium and magnesium by resin, 10% sodium potassium carbonate solution was added to the refined mother liquor for lithium carbonation precipitation. The sodium potassium carbonate solution came from step one, and the pH endpoint was controlled at 11. After filtration, washing, and drying, approximately 62 kg of lithium carbonate (Li2CO3) product was obtained.

[0023] Step 5, Preparation of potassium sulfate and sodium sulfate: The remaining lithium sulfate in the lithium precipitation mother liquor was extracted using a resin adsorbent or extractant. Then, rubidium sulfate and cesium sulfate were extracted using 4-tert-butyl-2-(α-methylbenzyl)phenol as the extractant in a three-stage countercurrent process. The purified potassium-rich solution was then cooled to 0°C to obtain a potassium-sodium sulfate mixture (K₂SO₄:Na₂SO₄.10H₂O = 3.56) and a freezing mother liquor. The potassium-sodium sulfate mixture was washed with water at <10°C to remove K₂SO₄, yielding 25.97 kg of sodium sulfate decahydrate. The potassium-containing eluent was mixed with the freezing mother liquor and then... The evaporation crystallization system continues, concentrating the product to 1 / 3 to 1 / 4 of its original volume during the evaporation crystallization process. Potassium sulfate double salt (3K₂SO₄·Na₂SO₄) and mother liquor are crystallized. The mother liquor is returned to the pre-freezing crystallization circulation preparation. After washing away Na₂SO₄ with water, approximately 85 kg of pure K₂SO₄ product and a sodium-containing eluent are obtained. The sodium-containing eluent is returned to the mother liquor after freezing crystallization for further evaporation crystallization to produce potassium sulfate double salt. All condensate from the secondary evaporation steam is reused. Preparation and recycling of intermediate products: Preparation of potassium sodium carbonate for lithium precipitation: 160 kg of potassium sodium sulfate with a 1.1 ratio is required. 1333.3 kg of a 12% K2SO4 / Na2SO4 (8:1) solution is used, and 68 kg of slaked lime is added to convert it into KOH and NaOH. Then, CO2 from the converted roasting flue gas is absorbed to prepare potassium sodium carbonate solution, yielding 144 kg of dihydrate gypsum. For wet impurity removal, potassium sodium hydroxide is required: 100 kg of a 12% potassium sodium sulfate concentrate is taken, and 5.6 kg of slaked lime is added to prepare it. The washing liquid from the above gypsum is reused for lime slurry preparation and leaching.

[0024] Example 2: Take 1000 kg of lithium mica concentrate from a certain place, containing 1.5 wt% Li2O, 4.0 wt% F, 6.0 wt% K2O and 0.6 wt% Na2O.

[0025] Step 1, Batching and Transformation Roasting: Based on the fluorine content, 155 kg of limestone powder is added in excess at a molar ratio of 1.1, and mixed evenly with lepidolite concentrate. The mixture is roasted in a rotary kiln at 900°C for 90 minutes. After crushing and grinding, the roasted clinker is controlled to have more than 80% of the particles below 200 mesh. After dust removal, the flue gas is reacted with a mixture of KOH and NaOH to generate a potassium sodium carbonate solution for later use.

[0026] Step 2, water immersion and fluorite flotation: The roasted clinker was leached with water at a solid-liquid ratio of 1:4, and the leaching was carried out with stirring at 60°C for 1 hour. Then, flotation was performed according to the flotation reagent system. Crude selection: 900 g / t oleic acid, 1000 g / t water glass, 80 g / t pine oil; Selected 5 times, oleic acid 60 g / t, water glass 50 g / t, pine oil 4 g / t; Two sweeping processes were performed, with the reagent dosage being 30% of that used in the roughing process. Approximately 82.8 kg of fluorite concentrate was obtained, with a CaF2 grade of 97.2% and a recovery rate of 98%; the filtrate was returned for leaching.

[0027] Step 3: Acidification roasting and secondary leaching; After filtration, the flotation tailings were mixed with 125 kg of 98% concentrated sulfuric acid (1.1 times the theoretical amount of Li+K), and then acidified and roasted at 250°C for 1 hour. The resulting water-leached flotation tailings (approximately 3.8 m³) were then used in step two. 3 The acidified roasted material was subjected to a second leaching at 80℃ for 1.5 hours. The second leaching slurry was filtered to obtain approximately 3.7m³. 3 Lithium-potassium solution.

[0028] Step 4: Slag-liquid separation and lithium preparation Lime and potassium sodium carbonate slurry were added to the secondary leaching solution in step three to separate the iron slag, magnesium slag, and calcium slag by filtration. The filtrate was then evaporated and concentrated to a water content of 80%, at which point the content of Li2SO4 was 6.62% and the content of K2SO4 was 13.38%. After further removal of calcium and magnesium by resin, 10% potassium sodium carbonate solution was added to the concentrated mother liquor for lithium carbonation precipitation. The potassium sodium carbonate solution came from step one, and the pH endpoint was controlled at 10. After filtration, washing, and drying, approximately 35.1 kg of lithium carbonate (Li2CO3) product was obtained.

[0029] Step 5, Preparation of potassium sulfate and sodium sulfate: The remaining lithium sulfate in the lithium precipitation mother liquor was extracted using a resin adsorbent or extractant. Then, rubidium sulfate and cesium sulfate were extracted using 4-tert-butyl-2-(α-methylbenzyl)phenol as the extractant in a three-stage countercurrent process. The purified potassium-rich solution was then cooled to 0°C to obtain a potassium-sodium sulfate mixture (K₂SO₄:Na₂SO₄.10H₂O = 3.56) and a freezing mother liquor. The potassium-sodium sulfate mixture was washed with water at <10°C to remove K₂SO₄, yielding 36.35 kg of sodium sulfate decahydrate. The potassium-containing eluent was mixed with the freezing mother liquor for subsequent processing. The evaporation crystallization system concentrates the product to 1 / 5 to 1 / 6 of its original volume during the evaporation crystallization process, crystallizing potassium mirabilite (3K2SO4·Na2SO4) and evaporation mother liquor. The evaporation mother liquor is returned to the pre-freezing crystallization recycling solution. After washing away Na2SO4 with water, approximately 112 kg of pure K2SO4 product and sodium-containing eluent are obtained. The sodium-containing eluent is returned to the mother liquor after freezing crystallization for recycling to produce potassium mirabilite. All the condensate from the secondary evaporation steam is reused.

[0030] Preparation and recycling of intermediate products: Preparation of potassium sodium carbonate for lithium precipitation: 96 kg of potassium sodium sulfate with a 1.1 ratio is required. 800 kg of a 12% K₂SO₄ / Na₂SO₄ (8:1) solution (after lithium precipitation) is used. 41 kg of quicklime is added to convert the solution into KOH and NaOH. Then, CO₂ from the conversion roasting flue gas is absorbed to produce potassium sodium carbonate, yielding 115 kg of dihydrate gypsum. For wet impurity removal, potassium sodium hydroxide is required: 100 kg of 12% potassium sodium sulfate concentrate is taken, and 5.7 kg of quicklime (10% excess) is added to prepare the solution. The washing liquid from the above gypsum is reused for lime slurry preparation and leaching.

[0031] Example 3: Take 1000 kg of lithium mica concentrate from a certain place, containing 2 wt% Li2O, 4.5 wt% F, 7.0 wt% K2O and 0.7 wt% Na2O.

[0032] Step 1, Batching and Transformation Roasting: Based on the fluorine content, 174 kg of limestone powder is added in excess at a molar ratio of 1.1, and mixed evenly with lepidolite concentrate. The mixture is roasted in a rotary kiln at 900℃ for 90 minutes. After crushing and grinding, the roasted clinker is controlled to have more than 80% of the particles below 200 mesh. After dust removal, the flue gas is reacted with a mixture of KOH and NaOH to generate a potassium sodium carbonate solution for later use.

[0033] Step 2, water immersion and fluorite flotation: Add 2.5 mg of roasted clinker at a solid-liquid ratio of 1:2.5. 3 Water leaching was performed with stirring at 60°C for 1 hour, followed by flotation according to the flotation reagent regime. Crude selection: 700 g / t oleic acid, 700 g / t water glass, 80 g / t pine oil; Five selections were made, yielding 60 g / t of oleic acid, 30 g / t of water glass, and 10 g / t of pine oil. Two sweeping processes were performed, with the reagent dosage being 30% of that used in the roughing process. Approximately 92.3 kg of fluorite concentrate was obtained, with a CaF2 grade of 99.2% and a recovery rate of 99.1%; the filtrate was returned for leaching.

[0034] Step 3: Acidification roasting and secondary leaching; After filtration, the flotation tailings were mixed with 198 kg of 98% concentrated sulfuric acid (1.4 times the theoretical amount of Li+K), and then acidified and roasted at 250°C for 1.2 hours. Then, the water-leached flotation tailings (approximately 2.4 m³) obtained in step two were used as a buffer. 3 The acidified roasted material was subjected to a second leaching at 70℃ for 1.5 hours. The second leaching slurry was filtered to obtain approximately 2.7m³. 3 The mixture contains a lithium-potassium solution and approximately 890 kg of cement raw material slag.

[0035] Step 4: Slag-liquid separation and lithium preparation Lime and sodium potassium carbonate slurry were added to the secondary leaching solution in step three to separate the iron slag, magnesium slag, and calcium slag by filtration. The filtrate was then evaporated and concentrated to a water content of 80%, at which point the content of Li2SO4 was 7.23% and the content of K2SO4 was 12.77%. After further removal of calcium and magnesium by resin, 10% sodium potassium carbonate solution was added to the concentrated mother liquor for lithium carbonation precipitation. The sodium potassium carbonate solution came from step one, and the pH endpoint was controlled at 11. After filtration, washing, and drying, approximately 47 kg of lithium carbonate (Li2CO3) product was obtained.

[0036] Step 5, Preparation of potassium sulfate and sodium sulfate: The remaining lithium sulfate in the lithium precipitation mother liquor was extracted using a resin adsorbent or extractant. Then, rubidium sulfate and cesium sulfate were extracted using 4-tert-butyl-2-(α-methylbenzyl)phenol as the extractant in a three-stage countercurrent process. The purified potassium-rich solution was then cooled to 0°C to obtain a potassium-sodium sulfate mixture (K₂SO₄:Na₂SO₄.10H₂O = 3.56) and a freezing mother liquor. The potassium-sodium sulfate mixture was washed with water at <10°C to remove K₂SO₄, yielding 36.35 kg of sodium sulfate decahydrate. The potassium-containing eluent was mixed with the freezing mother liquor for subsequent processing. The evaporation crystallization system concentrates the product to 1 / 3 to 1 / 4 of its original volume during the evaporation crystallization process, crystallizing potassium mirabilite (3K2SO4·Na2SO4) and evaporation mother liquor. The evaporation mother liquor is returned to the pre-freezing crystallization circulation solution. After washing away Na2SO4 with water, approximately 131 kg of pure K2SO4 product and sodium-containing eluent are obtained. The sodium-containing eluent is returned to the mother liquor after freezing crystallization for evaporation crystallization to produce potassium mirabilite. All the condensate from the secondary evaporation steam is reused.

[0037] Preparation and recycling of intermediate products: Preparation of potassium sodium carbonate for lithium precipitation: 127.6 kg of potassium sodium sulfate with a 1.1 ratio is required. 163.3 kg of a 12% K₂SO₄ / Na₂SO₄ (8:1) solution (after lithium precipitation) is used. 54.3 kg of quicklime is added to convert the solution into KOH and NaOH. Then, CO₂ from the converted roasting flue gas is absorbed to produce potassium sodium carbonate, yielding 158 kg of dihydrate gypsum. For wet impurity removal, potassium sodium hydroxide is required: 100 kg of 12% potassium sodium sulfate concentrate is taken, and 5.6 kg of quicklime (10% excess) is added to prepare the solution. The washing liquid from the above gypsum is reused for lime slurry preparation and leaching.

[0038] The foregoing has only described certain exemplary embodiments of the present invention by way of illustration. Undoubtedly, those skilled in the art can modify the described embodiments in various ways without departing from the spirit and scope of the present invention. Therefore, the foregoing drawings and descriptions are illustrative in nature and should not be construed as limiting the scope of protection of the claims of the present invention.

Claims

1. A method for the comprehensive extraction of lithium and recovery of by-products from lepidolite, characterized in that, Includes the following steps: Step 1, Ingredient preparation and conversion roasting: Lithium mica concentrate and calcium-containing materials are mixed at a fluorine to calcium oxide molar ratio of 1:1 to 1:2, and then converted and roasted at 850℃ to 950℃ to convert fluorine into calcium fluoride and lithium into lithium hydroxide, yielding roasted clinker and CO2-containing flue gas. The CO2 flue gas is then reacted with potassium sodium hydroxide solution after dust removal to generate potassium sodium carbonate solution, which is used for subsequent lithium precipitation. After crushing and grinding, the roasted clinker is controlled to have more than 80% particles smaller than 200 mesh. Step 2, water immersion and fluorite flotation: The roasted clinker obtained in step one is leached with water at a solid-liquid ratio of 1:1.5 to 1:

4. Fluorite (CaF2) flotation reagent is added, and the water-leached slurry is subjected to a flotation process of "primary roughing, multi-stage cleaning, and multi-stage scavenging" to obtain fluorite concentrate with a CaF2 grade >90% and a yield >85%. The slurry after flotation is filtered to obtain filtrate and tailings. The filtrate is returned to the leaching process after roasting. Step 3, acidification roasting and secondary leaching: Add sulfuric acid with a concentration of ≥98 wt% to the tailings obtained in step two, and control the amount of sulfuric acid added to be 1.0 to 1.4 times the theoretical molar amount of Li+K in the concentrate. Acid roasting is carried out at 200℃ to 300℃. The acid roasting product is leached a second time with the filtrate obtained in step two to convert the residual lithium, potassium and aluminum into soluble sulfates. Step 4, Slag-liquid separation and lithium product preparation: The secondary leaching slurry from step three is treated with lime to remove iron and aluminum. After filtration, iron-containing silicon-aluminum tailings and a mixed solution of lithium sulfate and potassium sodium sulfate are obtained. The tailings are used as cement raw materials. The mixed solution is evaporated and concentrated to a water content of ≥80%. After deep removal of calcium and magnesium by resin, the potassium sodium carbonate solution obtained in step one is used for carbonation and lithium precipitation to obtain lithium carbonate products and lithium precipitation mother liquor. Step 5, Preparation of potassium sulfate and sodium sulfate: The lithium precipitation mother liquor from step four is subjected to adsorption to recover residual lithium, and rubidium and cesium are extracted to recover rubidium and cesium, resulting in a potassium-rich solution. This potassium-rich solution is mixed with the subsequently returned evaporation mother liquor and then subjected to 0°C freeze crystallization to precipitate potassium-sulfuric acid mixed salt. A certain amount of 0-10°C low-temperature water is added to the mixed salt to wash away K2SO4, yielding pure decahydrate sodium sulfate. The freeze crystallization mother liquor is mixed with potassium-containing elution mother liquor and subsequently returned sodium-containing elution mother liquor and then fed into the evaporation crystallization system to crystallize potassium-sulfuric acid compound salt. The potassium-sulfuric acid compound salt is washed with room temperature water to remove all sodium sulfate and some potassium sulfate, yielding pure K2SO4 product. The sodium-containing elution mother liquor is returned to the fore-liquid recycling system of the aforementioned potassium-sulfuric acid compound salt evaporation crystallization system. All evaporation condensate is recycled for steps two and three. The system does not introduce sodium salt, and all sodium sulfate produced comes from lithium concentrate.

2. The method for comprehensively extracting lithium and recovering by-products from lepidolite according to claim 1, characterized in that: In step one, the lepidolite concentrate contains 1.5-2.5 wt% Li2O, 3.0-4.5 wt% F, and 0.5-0.7 wt% Na, and the calcium-containing material is limestone or quicklime; During the conversion roasting, the roasting time is 45 to 90 minutes.

3. The method for comprehensively extracting lithium and recovering by-products from lepidolite according to claim 1, characterized in that: In step two, the flotation reagent system is as follows: Roughing stage: collector 300-1000 g / t, inhibitor 300-1500 g / t, frother 15-130 g / t; Fine selection stage: 5-80 g / t of collector, 5-90 g / t of inhibitor, 0-10 g / t of foaming agent, and ≥5 fine selection times; Sweeping stage: The dosage of each reagent is 10% to 50% of that used in the roughing stage, and the number of sweeping operations is ≥1. The collector is selected from one or more of oleic acid, sodium oleate, and oxidized paraffin soap; the inhibitor is selected from one or two of water glass and sodium carbonate; and the foaming agent is selected from one or two of pine oil and No. 2 oil.

4. The method for comprehensively extracting lithium and recovering by-products from lepidolite according to claim 1, characterized in that: In step three, the acid roasting equipment is a rotary kiln or a circulating fluidized bed furnace, and the roasting residence time is 30 to 90 minutes. The small amount of HF and SO2 escaping from the roasting tail gas is desulfurized by lime method, and the resulting calcium slag is returned to the flotation system in step two.

5. The method for comprehensively extracting lithium and recovering by-products from lepidolite according to claim 1, characterized in that: In step three, during the secondary leaching, the solid-liquid ratio before leaching is 1.3:1.5 to 1.2:4, so that the residual lithium is converted into lithium sulfate, which enters the solution along with the potassium sulfate and aluminum sulfate formed at the same time. After leaching, the solid-liquid ratio of the slurry is 0.9:1.9 to 0.8:4.

4.

6. The method for comprehensively extracting lithium and recovering by-products from lepidolite according to claim 1, characterized in that: In step four, the amount of lime added is 1.0 to 1.2 times the amount of Fe in the concentrate; the evaporation and concentration endpoint is controlled to have a solution water content ≥80%; after removing calcium and magnesium from the resin, Ca... 2+ <20 mg / L, Mg 2+ <10 mg / L.

7. The method for comprehensively extracting lithium and recovering by-products from lepidolite according to claim 1, characterized in that: In step four, the amount of potassium sodium carbonate solution added is 1.0 to 1.2 times the Li content in the concentrate, and the reaction is carried out at 80 to 100°C and at a final pH of 10 to 11 for 20 to 60 minutes. The precipitate of Li2CO3 is obtained by filtration.

8. The method for comprehensively extracting lithium from lepidolite and recovering by-products according to claim 1, characterized in that: In step five, the rubidium and cesium extractant is 4-tert-butyl-2-(α-methylbenzyl)phenol, and a three-stage countercurrent extraction is performed, with a total recovery rate of rubidium and cesium ≥90%; the K2SO4 crystallization temperature is controlled at 25-60℃, and the crystallization rate is ≥80%.

9. A method for comprehensively extracting lithium from lepidolite and recovering by-products according to claim 1, characterized in that: In step five, a portion of the potassium-sodium sulfate mixed mother liquor reacts with lime to produce potassium-sodium hydroxide and calcium sulfate. Part of the potassium-sodium hydroxide is used to absorb the roasting flue gas from step one to produce potassium-sodium carbonate, and the other part is used as a pH adjuster. The calcium sulfate is washed multiple times to obtain calcium sulfate waste residue. The calcium sulfate waste residue is combined with calcium silicate slag and sold as cement raw material, realizing a dual-loop cycle of potassium and calcium and zero discharge of waste residue.

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