A lithium-aluminate integrated extraction method

Abstract

The present application relates to the technical field of ore lithium extraction, and particularly relates to a lithium-aluminate acid-heat integrated extraction method.The lithium-aluminate acid-heat integrated extraction method can simultaneously obtain lithium carbonate, aluminum oxide, potassium sulfate and sulfuric acid products, the potassium sulfate can be used as a raw material to prepare potassium alum, and the obtained sulfuric acid can be used for roasting of lithium-containing materials, thereby significantly reducing the use amount of concentrated sulfuric acid and potassium sulfate.Circulating leaching can improve the lithium ion content in liquid I and reduce the concentration cost.The waste heat generated by roasting can be used for concentration, dissolution of potassium sulfate and dissolution of coarse potassium alum, the waste heat generated by calcination of potassium alum can be used for roasting of lithium-containing materials, thereby achieving full utilization of waste heat and greatly reducing the energy consumption of extraction.In summary, the lithium-aluminate acid-heat integrated method has the characteristics of obtaining multiple products, low energy consumption, low material consumption, low cost and maximum utilization of lithium-containing materials.

Description

Technical Field

[0001] This invention relates to the field of lithium extraction technology from ores, specifically to an integrated thermal extraction method for lithium aluminum ore. Background Technology

[0002] Lithium is an important rare metal raw material, and extracting lithium from lithium-containing ores is an important source of lithium.

[0003] Currently, the main steps of the sulfuric acid process for lithium extraction from lithium-containing materials are as follows: lithium-containing materials are mixed with concentrated sulfuric acid at a mass ratio of 0.5 to 1.5:1 and roasted at a temperature of 200 to 280°C. The resulting roasted material is leached, lime is added to control the pH value to around 11, sodium carbonate is added to remove impurities such as calcium and magnesium, and the solution is filtered to obtain a clear liquid. After evaporation and concentration of the clear liquid, sodium carbonate is added to precipitate lithium carbonate. The existing sulfuric acid process for lithium extraction has the following main disadvantages: (1) the leaching solution has a high content of impurities such as Al and Fe, resulting in a heavy subsequent purification load and a large amount of slag; (2) the recycling of sulfuric acid is not considered, resulting in a large amount of sulfuric acid used in the process and a high cost; (3) the Al content in the leaching solution is high (5 to 20 g / L), and the comprehensive recovery of Al is not considered; (4) the recycling of the leaching solution is not considered, resulting in a high cost for subsequent evaporation and concentration; (5) the reuse of waste heat is not considered. Therefore, it is of great significance to provide an extraction method that can fully recover lithium and aluminum from lithium-containing materials, with low energy consumption and low cost. Summary of the Invention

[0004] In view of this, the purpose of this invention is to provide a method for the integrated thermal extraction of lithium and aluminum, which can simultaneously extract lithium and aluminum with low energy consumption and low production cost.

[0005] To achieve the above-mentioned objectives, the present invention provides the following technical solution:

[0006] This invention provides a method for the integrated thermal extraction of lithium aluminum acid, comprising the following steps:

[0007] (1) The lithium-containing material is mixed with concentrated sulfuric acid and roasted to obtain roasted material and roasting flue gas respectively; the residual heat of roasting is used in steps (3) and (4); the mass fraction of the concentrated sulfuric acid is ≥85%;

[0008] (2) The calcined material is pulverized to obtain calcined pulverized material; the obtained calcined pulverized material is mixed with a leaching agent, leached, and then filtered to obtain filtrate and filter residue respectively; the filter residue is washed with water to obtain washing liquid; the filtrate and washing liquid are combined to obtain leachate; the leaching agent is water; the leaching is carried out under the condition of passing oxygen-containing gas; the solid-liquid ratio of the calcined pulverized material to the water used for washing is 1g:0.5-2mL;

[0009] (3) Adjust the pH of the leachate to 1.5-3, add potassium sulfate to dissolve it, and then filter the resulting mixture after iron removal treatment to obtain filtrate and filter residue. Wash the filter residue with water to obtain washing liquid. Combine the filtrate and washing liquid and concentrate them to 40-60% of the volume of the leachate. After crystallization, separate the solid and liquid to obtain crude potassium alum and liquid I. The pH adjustment uses an alkali including carbonate. The K in the mixture... + With Al 3+ The molar ratio is 0.8 to 1.05:1; the volume ratio of the leachate to the washing water is 1:0.05 to 0.5; the liquid I is recycled as a leaching agent in step (2);

[0010] (4) Dissolve the crude potassium alum II in water, and perform recrystallization and solid-liquid separation in sequence to obtain refined potassium alum and liquid II respectively; liquid II is reused as a leaching agent in step (2);

[0011] (5) Repeat steps (1) to (4) n times. When the lithium ion concentration in liquid I is >3g / L, heat liquid I to 40-100℃, adjust the pH value to 5-7 using carbonate source, and neutralize for 0.5-2h. Add impurity removal agent to adjust the pH value to >9 and remove impurities for 0.5-2h. Filter to obtain filtrate and filter residue respectively. Wash the filter residue with water to obtain washing liquid. Combine the filtrate and washing liquid to obtain crude lithium solution. n≥2. The volume ratio of liquid I to water used for washing is 1:0.1-0.5. In the repeated process, the leaching agent in step (2) is a mixed solution of liquid I in step (3), liquid II in step (4) and water added.

[0012] (6) The crude lithium solution is purified and impurities are removed by using a resin ion exchange column to obtain a refined lithium solution;

[0013] (7) The purified lithium solution is concentrated by homogeneous membrane electrodialysis to obtain a concentrated Li₂SO₄ solution; the concentrated Li₂SO₄ solution is mixed with an aqueous Na₂CO₃ solution, and the resulting mixture is subjected to lithium precipitation followed by solid-liquid separation to obtain a lithium precipitation mother liquor and crude lithium carbonate; the crude lithium carbonate is washed with water and dried to obtain a lithium carbonate product; the Li₂SO₄ concentrate contains Li + The concentration of Na in the mixture is ≥20 g / L; + With Li + The molar ratio is 1 to 1.05:1;

[0014] (8) The lithium precipitation mother liquor is concentrated, sodium is precipitated and solid-liquid separation is performed sequentially to obtain sodium sulfate product and sodium precipitation mother liquor respectively; the sodium precipitation mother liquor is used to replace carbonate source to adjust pH value in step (3) of the next lithium aluminum acid thermal integrated extraction process;

[0015] (9) The refined potassium alum is successively roasted to remove crystallization water and calcined to obtain calcined material and calcination flue gas respectively; the calcined material is soaked in water and then filtered to obtain filtrate and crude alumina; the crude alumina is washed with water to obtain washing liquid and alumina product respectively; the filtrate and washing liquid are combined to obtain water immersion liquid; the solid-liquid ratio of the calcined material to the water used for washing is 1g:0.1~1mL; the residual heat from calcination is recycled to steps (1) and (2);

[0016] (10) The aqueous extract is concentrated, crystallized and separated into solid and liquid components in sequence to obtain potassium sulfate product; the potassium sulfate product is recycled in step (3);

[0017] (11) The roasting flue gas and calcination flue gas are absorbed by water or dilute sulfuric acid to obtain sulfuric acid solution and tail gas respectively; when the mass fraction of the sulfuric acid solution is ≥85%, it is recycled as concentrated sulfuric acid in step (1).

[0018] There is no chronological order between steps (6) to (8) and steps (9) to (10).

[0019] Preferably, in step (1), the lithium-containing material includes one or more of lithium-containing clay ore, lepidolite concentrate and spodumene concentrate heat-insulating roasting pretreatment material;

[0020] The mass ratio of the lithium-containing material to concentrated sulfuric acid is 1:0.2 to 1.2;

[0021] The roasting includes sequential heating and holding roasting; the heating rate is 5-10℃ / min, and the initial temperature of the heating is room temperature; the holding roasting temperature is 300-550℃, and the holding time is 0.5-6h.

[0022] Preferably, in step (2), the solid-liquid ratio of the calcined pulverized material to the leaching agent is 1g:2-5mL; the leaching temperature is 40-100℃ and the leaching time is 0.5-2h.

[0023] Preferably, in step (2), the iron reduction treatment is performed at a temperature of 90–100°C for a time of 0.5–2 hours.

[0024] The carbonate source includes carbonates and / or carbonate minerals.

[0025] Preferably, in steps (3) and (4), the temperatures of the first dissolution and the second dissolution are independently 40–85°C;

[0026] The crystallization and recrystallization are carried out at temperatures ranging from 5 to 25°C and at times ranging from 5 to 24 hours, respectively, and both crystallization and recrystallization are carried out by static crystallization.

[0027] Preferably, in step (5), n is 2 to 10;

[0028] The carbonate source includes carbonates and / or carbonate minerals.

[0029] Preferably, in step (6), the resin in the resin ion exchange column includes ZGC258 macroporous weak acid cation exchange resin and / or LSC-500B ion exchange resin.

[0030] Preferably, in step (7), the temperature of lithium deposition is ≥95℃ and the time is 0.5~1.5h.

[0031] Preferably, in step (8), the sodium precipitation temperature is 0-5°C.

[0032] Preferably, in step (9), the temperature for calcination and decrystallization is 150–300°C, and the time is 0.5–3 hours.

[0033] The calcination temperature is 750–900℃, and the time is 1–5 hours;

[0034] The water immersion temperature is 20–100℃, and the time is 0.5–2 hours.

[0035] Extracting lithium carbonate or alumina separately from lithium-containing materials is costly and cannot achieve simultaneous lithium and aluminum extraction. This invention employs an integrated acid-thermal extraction method for lithium and aluminum, which produces lithium carbonate, alumina, potassium sulfate, and sulfuric acid during the lithium extraction process. Alumina is obtained during lithium carbonate production, and its sale can offset a portion of the lithium carbonate production cost; similarly, lithium carbonate is obtained during alumina production, and its sale can offset a portion of the alumina production cost. Therefore, the integrated lithium and aluminum extraction reduces the overall production cost of extracting lithium and aluminum from lithium-containing materials. The potassium sulfate obtained during the integrated acid-thermal extraction process can be used as a raw material to prepare potassium alum, and the resulting sulfuric acid can be used to calcine lithium-containing materials, resulting in a sulfuric acid recycling rate of over 80% and a potassium sulfate recycling rate of over 95%. This significantly reduces the amount of concentrated sulfuric acid and potassium sulfate used, thus significantly lowering extraction costs. Circulating leaching can increase the lithium-ion content in liquid I, reducing concentration costs. Calcination of potassium alum yields alumina that meets the GB / T24487-2022 standard. The waste heat generated during roasting can be used for concentration, dissolution of potassium sulfate, and dissolution of crude potassium alum. The waste heat generated from the calcination of potassium alum can also be used for roasting, achieving full utilization of waste heat and significantly reducing the energy consumption of extraction. The reuse of liquids II and I increases the lithium-ion concentration and reduces the cost of membrane concentration of refined lithium solution. The reuse of sodium precipitation mother liquor can reduce the reagent cost of carbonate source. In summary, the lithium aluminum acid-thermal integrated method provided by this invention has the characteristics of obtaining multiple products, low energy consumption, low material consumption, low cost, and maximizing the utilization of lithium-containing materials. Attached Figure Description

[0036] Figure 1 The diagram shows the integrated thermal extraction process of lithium aluminum acid in this embodiment. Detailed Implementation

[0037] This invention provides a method for the integrated thermal extraction of lithium aluminum acid, comprising the following steps:

[0038] (1) The lithium-containing material is mixed with concentrated sulfuric acid and roasted to obtain roasted material and roasting flue gas respectively; the residual heat of roasting is used in steps (3) and (4); the mass fraction of the concentrated sulfuric acid is ≥85%;

[0039] (2) The calcined material is pulverized to obtain calcined pulverized material; the obtained calcined pulverized material is mixed with a leaching agent, leached, and then filtered to obtain filtrate and filter residue respectively; the filter residue is washed with water to obtain washing liquid; the filtrate and washing liquid are combined to obtain leachate; the leaching agent is water; the leaching is carried out under the condition of passing oxygen-containing gas; the solid-liquid ratio of the calcined pulverized material to the water used for washing is 1g:0.5-2mL;

[0040] (3) Adjust the pH of the leachate to 1.5-3, add potassium sulfate to dissolve it, and then filter the resulting mixture after iron removal treatment to obtain filtrate and filter residue. Wash the filter residue with water to obtain washing liquid. Combine the filtrate and washing liquid and concentrate them to 40-60% of the volume of the leachate. After crystallization, separate the solid and liquid to obtain crude potassium alum and liquid I. The pH adjustment uses an alkali including carbonate. The K in the mixture... + With Al 3+ The molar ratio is 0.8 to 1.05:1; the volume ratio of the leachate to the washing water is 1:0.05 to 0.5; the liquid I is recycled as a leaching agent in step (2);

[0041] (4) Dissolve the crude potassium alum II in water, and perform recrystallization and solid-liquid separation in sequence to obtain refined potassium alum and liquid II respectively; liquid II is reused as a leaching agent in step (2);

[0042] (5) Repeat steps (1) to (4) n times. When the lithium ion concentration in liquid I is >3g / L, heat liquid I to 40-100℃, adjust the pH value to 5-7 using carbonate source, and neutralize for 0.5-2h. Add impurity removal agent to adjust the pH value to >9 and remove impurities for 0.5-2h. Filter to obtain filtrate and filter residue respectively. Wash the filter residue with water to obtain washing liquid. Combine the filtrate and washing liquid to obtain crude lithium solution. n≥2. The volume ratio of liquid I to water used for washing is 1:0.1-0.5. In the repeated process, the leaching agent in step (2) is a mixed solution of liquid I in step (3), liquid II in step (4) and water added.

[0043] (6) The crude lithium solution is purified and impurities are removed by using a resin ion exchange column to obtain a refined lithium solution;

[0044] (7) The purified lithium solution is concentrated by homogeneous membrane electrodialysis to obtain a concentrated Li₂SO₄ solution; the concentrated Li₂SO₄ solution is mixed with an aqueous Na₂CO₃ solution, and the resulting mixture is subjected to lithium precipitation followed by solid-liquid separation to obtain a lithium precipitation mother liquor and crude lithium carbonate; the crude lithium carbonate is washed with water and dried to obtain a lithium carbonate product; the Li₂SO₄ concentrate contains Li + The concentration of Na in the mixture is ≥20 g / L; + With Li + The molar ratio is 1 to 1.05:1;

[0045] (8) The lithium precipitation mother liquor is concentrated, sodium is precipitated and solid-liquid separation is performed sequentially to obtain sodium sulfate product and sodium precipitation mother liquor respectively; the sodium precipitation mother liquor is used to replace carbonate source to adjust pH value in step (3) of the next lithium aluminum acid thermal integrated extraction process;

[0046] (9) The refined potassium alum is successively roasted to remove crystallization water and calcined to obtain calcined material and calcination flue gas respectively; the calcined material is soaked in water and then filtered to obtain filtrate and crude alumina; the crude alumina is washed with water to obtain washing liquid and alumina product respectively; the filtrate and washing liquid are combined to obtain water immersion liquid; the solid-liquid ratio of the calcined material to the water used for washing is 1g:0.1~1mL; the residual heat from calcination is recycled to steps (1) and (2);

[0047] (10) The aqueous extract is concentrated, crystallized and separated into solid and liquid components in sequence to obtain potassium sulfate product; the potassium sulfate product is recycled in step (3);

[0048] (11) The roasting flue gas and calcination flue gas are absorbed by water or dilute sulfuric acid to obtain sulfuric acid solution and tail gas respectively; when the mass fraction of the sulfuric acid solution is ≥85%, it is recycled as concentrated sulfuric acid in step (1).

[0049] There is no chronological order between steps (6) to (8) and steps (9) to (10).

[0050] Unless otherwise specified, all raw materials used in this invention are commercially available products.

[0051] This invention involves mixing lithium-containing materials with concentrated sulfuric acid and roasting them to obtain roasted feed and roasting flue gas, respectively; the concentrated sulfuric acid has a mass fraction ≥85%. In this invention, the lithium-containing materials preferably include one or more of lithium-containing clay ore, lepidolite concentrate, and spodumene concentrate high-temperature roasting pretreatment materials, more preferably including lithium-containing clay ore, lepidolite concentrate, or spodumene concentrate high-temperature roasting pretreatment materials. In this invention, the content of -0.074mm particle size in the lithium-containing materials is preferably ≥50wt%, more preferably ≥80wt%, where "-" indicates "≤". In this invention, the preparation method of the spodumene concentrate high-temperature roasting pretreatment material preferably includes the following steps: roasting spodumene concentrate to obtain spodumene concentrate high-temperature roasting pretreatment material; the roasting temperature is preferably 900–1100℃, more preferably 1000℃, and the roasting holding time is preferably 1–3h, more preferably 2h. In this invention, the concentrated sulfuric acid has a mass fraction of ≥85%, preferably 90-98%, and more preferably 95-98%. In this invention, the mass ratio of the lithium-containing material to the concentrated sulfuric acid is 1:0.5-1.2, preferably 1:0.5-1.1, and more preferably 1:0.5-1.

[0052] In this invention, the calcination preferably includes sequential heating and holding calcination; the heating rate is preferably 5-10°C / min, more preferably 8-9°C / min, the initial heating temperature is preferably room temperature, and the final heating temperature is the holding calcination temperature; the holding calcination temperature is preferably 300-550°C, more preferably 350-500°C, and even more preferably 380-450°C; the holding calcination time is preferably 1-6 hours, more preferably 1-4 hours, and even more preferably 2-3 hours. In this invention, the residual heat from the calcination is used for concentration in step (3) and in step (4).

[0053] After obtaining the calcined material, the present invention pulverizes the calcined material to obtain calcined pulverized material; mixes the obtained calcined pulverized material with a leaching agent, performs leaching, and then filters to obtain filtrate and filter residue respectively; washes the filter residue with water to obtain washing liquid; combines the filtrate and washing liquid to obtain leachate; the leaching agent is water; the leaching is carried out under the condition of introducing oxygen-containing gas; the solid-liquid ratio of the calcined pulverized material to the washing water is 1g:0.5-2mL. The present invention does not have a special limitation on the pulverization; pulverization can be carried out using a pulverization method well known to those skilled in the art until the content of the -0.15mm particle size of the obtained calcined pulverized material is ≥30wt%, more preferably ≥50wt%, and even more preferably ≥60wt%. In the present invention, the solid-liquid ratio of the calcined pulverized material to the leaching agent is preferably 1g:2-5mL, more preferably 1g:3-4mL. In this invention, the leaching temperature is preferably 40–100°C, more preferably 60–80°C; the leaching time is preferably 0.5–2 h, more preferably 1–1.5 h. In this invention, the oxygen-containing gas preferably includes air or oxygen, more preferably air; this invention does not have a particular limitation on the amount of oxygen-containing gas introduced, as long as it is sufficient to oxidize ferrous ions to ferric ions. In this invention, the solid-liquid ratio of the calcined pulverized material to the washing water is 1 g: 0.5–2 mL, preferably 1 g: 1–1.5 mL.

[0054] After obtaining the leachate, the present invention adjusts the pH value of the leachate to 1.5-3, adds potassium sulfate for initial dissolution, and then filters the resulting mixture after iron reduction treatment to obtain filtrate and filter residue. The filter residue is washed with water to obtain washing liquid. The filtrate and washing liquid are combined and concentrated to 40-60% of the leachate volume, crystallized, and then separated into solid and liquid components to obtain crude potassium alum and liquid I, respectively. The pH value is adjusted using an alkali including carbonate. The K in the mixture... + With Al 3+The molar ratio is 0.8–1.05:1; the volume ratio of the leachate to the washing water is 1:0.05–0.5; liquid I is reused as a leaching agent in step (2). In this invention, the pH adjustment preferably includes sequentially adjusting the pH to 1.5–3 (more preferably 1.5–2.5) and maintaining the temperature at 90–100°C (more preferably 95°C) for 0.5–2 h (more preferably 1–1.5 h). In this invention, the carbonate source preferably includes carbonates and / or carbonate minerals; the carbonate preferably includes one or more of sodium carbonate, potassium carbonate, and calcium carbonate; the carbonate mineral preferably includes one or more of limestone, calcite, carbonate-bearing rocks, calcite concentrate, and dolomite; this invention does not specifically limit the amount of carbonate source used, but adjusts the pH to 1.5–3. In this invention, the K in the mixture + With Al 3+ The molar ratio is 0.8–1.05:1, preferably 0.9–1.05:1, and more preferably 1–1.05:1; the dissolution is preferably achieved by heating the leachate and then adding potassium sulfate while maintaining the temperature and stirring until dissolved, the heating temperature (i.e., the dissolution temperature) is preferably 40–85°C, and more preferably 40–80°C; the heat source for dissolution preferably includes the residual heat from the roasting. In this invention, the iron reduction treatment temperature is preferably 90–100°C, and more preferably 95°C; the iron reduction treatment time is preferably 0.5–2 h, and more preferably 1–1.5 h. In this invention, the volume ratio of the leachate to the washing water is 1:0.05–0.5, preferably 1:0.1–0.4, and more preferably 1:0.2–0.3. This invention does not have a special limitation on the concentration, and concentration methods well known to those skilled in the art can be used; preferably, it is concentrated to 45–55% of the leachate volume, and more preferably to 50% of the leachate volume. In this invention, the crystallization temperature is preferably 5–25°C, more preferably 5–10°C, and the crystallization time is preferably 5–24 h, more preferably 5–15 h; the crystallization is preferably static crystallization. In this invention, the solid-liquid separation method preferably includes filtration, vacuum filtration, or centrifugation.

[0055] After obtaining crude potassium alum, the present invention dissolves the crude potassium alum in water, and performs recrystallization and solid-liquid separation sequentially to obtain refined potassium alum and liquid II, respectively; liquid II is reused as a leaching agent in step (2). In the present invention, the mass ratio of crude potassium alum to water volume is preferably 1g:0.2-1mL, more preferably 1g:0.25-0.5mL. In the present invention, the dissolution temperature is preferably 40-85℃, more preferably 40-80℃; the heat source for dissolution preferably includes the residual heat from roasting. In the present invention, the recrystallization temperature is preferably 5-25℃, more preferably 15-25℃, and the recrystallization time is preferably 5-24h, more preferably 5-15h; the recrystallization is preferably static recrystallization. In the present invention, the solid-liquid separation method preferably includes filtration, vacuum filtration, or centrifugation.

[0056] This invention repeats steps (1) to (4) n times. When the lithium ion concentration in liquid I is >3 g / L, liquid I is heated to 40-100°C, the pH is adjusted to 5-7 using carbonate source, and the neutralization reaction is carried out for 0.5-2 h. An impurity removal agent is added to adjust the pH to >9, and the impurity removal treatment is carried out for 0.5-2 h. After filtration, filtrate and filter residue are obtained respectively. The filter residue is washed with water to obtain washing liquid. The filtrate and washing liquid are combined to obtain crude lithium solution. n≥2. The volume ratio of liquid I to washing water is 1:0.1-0.5. In the repeated process, the leaching agent in step (2) is a mixed solution of liquid I in step (3), liquid II in step (4), and water. In this invention, n≥2 is preferably 2-10. In this invention, the lithium ion concentration in liquid I obtained in the last repetition is preferably 3-20 g / L, more preferably 4-15 g / L. In this invention, the carbonate source preferably includes carbonates and / or carbonate minerals; the carbonate preferably includes one or more of sodium carbonate, potassium carbonate, and calcium carbonate; the carbonate mineral preferably includes one or more of limestone, calcite, calcite concentrate, and dolomite; this invention does not have a special limitation on the amount of carbonate source used, but it is based on adjusting the pH value to 5-7 after the liquid I is heated to 40-100°C, and the pH value is more preferably 5-6; the temperature of the neutralization reaction is preferably 40-90°C, more preferably 40-50°C, and the time of the neutralization reaction is preferably 0.5-1.5h, more preferably 0.5-1h. In this invention, the impurity removal agent preferably includes one or more of lime, Ca(OH)2, and carbide slag. The amount of the impurity removal agent is not particularly limited, but the pH value of the system is preferred to be >9, more preferably 10-11, and even more preferably 10. The temperature of the impurity removal treatment is preferably 40-90℃, more preferably 40-50℃, and the time of the impurity removal treatment is preferably 0.5-1.5 h, more preferably 0.5-1 h. In this invention, the solid-liquid separation method preferably includes filtration, vacuum filtration, or centrifugation.

[0057] After obtaining the crude lithium solution, this invention utilizes a resin ion exchange column for deep purification and impurity removal to obtain a refined lithium solution. In this invention, the resin in the resin ion exchange column preferably includes ZGC258 macroporous weak acid cation exchange resin and / or LSC-500B ion exchange resin. In this invention, the resin is preferably pretreated before use. The pretreatment preferably includes: loading the resin into the ion exchange column, eluting with deionized water until the effluent is clear, eluting with NaOH aqueous solution, and eluting with deionized water until the pH of the effluent is 7-9, thus obtaining a pretreated resin ion exchange column. In this invention, the mass concentration of the NaOH aqueous solution is preferably 2-5%, more preferably 2-3%; the elution flow rate of the NaOH aqueous solution is preferably 3-5 m / h, more preferably 3 m / h, 4 m / h, or 5 m / h. In this invention, the flow rate of the crude lithium solution during the deep purification and impurity removal process is preferably 1-2 BV / h, more preferably 1.5-2 BV / h.

[0058] After obtaining a refined lithium solution, the present invention uses homogeneous membrane electrodialysis to concentrate the refined lithium solution to obtain a Li₂SO₄ concentrate; the Li₂SO₄ concentrate is mixed with a Na₂CO₃ aqueous solution, and the resulting mixture is subjected to lithium precipitation followed by solid-liquid separation to obtain a lithium precipitation mother liquor and crude lithium carbonate; the crude lithium carbonate is washed with water and dried to obtain the lithium carbonate product; the Li₂SO₄ concentrate contains Li + The concentration of Na in the mixture is ≥20 g / L; + With Li + The molar ratio is 1 to 1.05:1. In this invention, the concentration of Li+ in the Li2SO4 concentrate is ≥20 g / L, preferably 20 to 25 g / L. In this invention, the concentration of the Na2CO3 aqueous solution is preferably 200 to 350 g / L, more preferably 300 g / L. In this invention, the mixing temperature is preferably ≥95℃, more preferably 95 to 100℃, and even more preferably 97 to 98℃. In this invention, the Na+ concentration in the mixture is... + With Li +The molar ratio is 1–1.05:1, preferably 1.03–1.05:1. In this invention, the lithium precipitation temperature is preferably ≥95℃, more preferably 95–100℃, and even more preferably 97–98℃; the lithium precipitation time is preferably 0.5–1.5 h, more preferably 1 h. In this invention, the solid-liquid separation method preferably includes filtration, vacuum filtration, or centrifugation. In this invention, the temperature of the washing water is preferably ≥95℃, more preferably 95–100℃, and even more preferably 97–98℃; the water is preferably deionized water; the number of washings is preferably 1–3 times, more preferably 2 times; the volume ratio of the dry weight of crude lithium carbonate to the water used for a single washing is preferably 1 g: 1.5–3 mL, more preferably 1 g: 2–2.5 mL. The volume ratio of the dry weight of crude lithium carbonate to the water used for a single washing is preferably 1 g: 0.5–2 mL, more preferably 1 g: 1–1.5 mL. In this invention, the drying temperature is preferably 75-100°C, more preferably 80-90°C.

[0059] After obtaining the lithium precipitation mother liquor, the present invention sequentially concentrates, precipitates sodium, and separates the solid and liquid components to obtain sodium sulfate product and sodium precipitation mother liquor, respectively. The sodium precipitation mother liquor is used to replace carbonate source in step (3) of the next lithium aluminum acid thermal extraction process to adjust the pH value. In the present invention, the volume of the concentrated lithium precipitation solution obtained by concentration is preferably 1 / 2 to 2 / 3 of the volume of the lithium precipitation mother liquor, preferably 1 / 2 to 3 / 5, and the concentration is preferably evaporation concentration. In the present invention, the sodium precipitation temperature is preferably 0 to 5°C, more preferably 3 to 5°C, and the present invention does not have a special limitation on the sodium precipitation time, as long as the sodium precipitation reaches the point where sodium sulfate no longer needs to be added. In the present invention, the solid-liquid separation method preferably includes filtration, vacuum filtration, or centrifugation.

[0060] After obtaining refined potassium alum, the present invention sequentially roasts and dehydrates the refined potassium alum to obtain calcined material and calcination flue gas; the calcined material is then filtered after water immersion to obtain filtrate and crude alumina; the crude alumina is washed with water to obtain washing liquid and alumina product; the filtrate and washing liquid are combined to obtain a water immersion liquid; the solid-liquid ratio of the calcined material to the water used for washing is 1g:0.1-1mL; the residual heat from calcination is recycled to steps (1) and (2). In the present invention, the roasting and dehydration temperature is preferably 150-300℃, more preferably 150-200℃, and the roasting and dehydration time is preferably 0.5-3h, more preferably 1-2h. In the present invention, the calcination temperature is preferably 750-900℃, more preferably 800-850℃, and the calcination time is preferably 1-5h, more preferably 2-3h. In this invention, the solid-liquid ratio of the calcined material to the washing water is 1g:0.1-1mL, preferably 1g:0.2-0.8mL, more preferably 1g:0.3-0.5mL; the water immersion temperature is preferably 20-100℃, more preferably 50-60℃, and the water immersion time is preferably 0.5-2h, more preferably 1-1.5h. In this invention, the preferred solid-liquid ratio of the calcined material to the washing water is 1g:0.1-1mL, more preferably 1g:0.2-1mL.

[0061] After obtaining the aqueous extract, the present invention sequentially concentrates, crystallizes, and separates the solid and liquid components to obtain potassium sulfate product; the potassium sulfate product is recycled in step (3). In the present invention, the K2SO4 mass concentration in the concentrated aqueous extract is preferably 15-18%, more preferably 17-18%. In the present invention, the crystallization temperature is 0-10℃, more preferably 0-5℃; the crystallization time is preferably 5-24h, more preferably 10-15h; the crystallization is static crystallization. In the present invention, the solid-liquid separation method preferably includes filtration, vacuum filtration, or centrifugation.

[0062] After obtaining the roasting flue gas and the calcination flue gas, the present invention uses water or dilute sulfuric acid to absorb the roasting flue gas and the calcination flue gas, respectively, to obtain sulfuric acid solution and tail gas; when the mass fraction of the sulfuric acid solution is ≥85%, it is recycled as concentrated sulfuric acid in step (1). In the present invention, the mass fraction of the sulfuric acid is preferably 90-98 wt%, more preferably 95-98 wt%.

[0063] The present invention preferably further includes purifying the exhaust gas. The purifying agent used for purification preferably includes a carbonate aqueous solution, a carbonate mineral slurry, or an alkaline aqueous solution. The carbonate in the carbonate aqueous solution preferably includes one or more of sodium carbonate, potassium carbonate, and calcium carbonate. The carbonate mineral in the carbonate mineral slurry preferably includes one or more of limestone, calcite, calcite concentrate, and dolomite. The present invention does not have a special limitation on the concentration of the carbonate aqueous solution, carbonate mineral slurry, or alkaline aqueous solution, as long as it can achieve the purification of the exhaust gas.

[0064] The technical solutions of this invention will be clearly and completely described below with reference to the embodiments thereof. Obviously, the described embodiments are only a part of the embodiments of this invention, and not all of them. All other embodiments obtained by those skilled in the art based on the embodiments of this invention without creative effort are within the scope of protection of this invention.

[0065] Example 1

[0066] according to Figure 1 The process flow diagram shown illustrates the integrated thermal extraction of lithium aluminum acid. The specific steps are as follows:

[0067] (1) The lithium-containing clay mineral powder (-0.074mm particle size content 75.33wt%) shown in Table 1 and 98wt% concentrated sulfuric acid are thoroughly mixed in a mass ratio of 1:1. The mixture is heated from room temperature to 400℃ at a heating rate of 10℃ / min and then kept at the temperature for 2 hours to obtain the roasted material and roasting flue gas. The residual heat obtained from the roasting is used in steps (3) and (4).

[0068] Table 1. Multi-element analysis results of lithium-bearing clay deposits

[0069] Components <![CDATA[Li2O]]> <![CDATA[Al2O3]]> <![CDATA[SiO2]]> CaO MgO <![CDATA[TFe2O3]]> <![CDATA[TiO2]]> <![CDATA[K2O]]> <![CDATA[Na2O]]> <![CDATA[MnO2 <!-- 7 -->]]> Content / wt% 0.55 21.20 52.49 5.05 0.35 7.15 0.88 1.32 0.08 0.10

[0070] (2) The calcined material is pulverized to a particle size of -0.15mm with a content of ≥60wt% to obtain calcined pulverized material; the calcined pulverized material is mixed with an extractant at a ratio of 1g:3mL, and after leaching at 80℃ for 1h under air circulation, it is filtered to obtain filtrate and filter residue respectively; the filter residue is washed with water to obtain washing liquid; the filtrate and washing liquid are combined to obtain leachate; wherein, the extractant is water, and the solid-liquid ratio of the calcined pulverized material to the water used for washing is 1g:2mL; the amount of air introduced is based on the complete oxidation of ferrous ions to ferric ions;

[0071] (3) The pH of the leachate was adjusted to 2 using carbonate source, heated to 60°C, and potassium sulfate was added to dissolve it. The resulting mixture was subjected to iron removal treatment at 92°C for 1 hour, and then filtered to obtain filtrate and filter residue. The filter residue was washed with water to obtain washing liquid. The filtrate and washing liquid were combined and concentrated to 50% of the leachate volume, cooled to 5°C, and allowed to stand for crystallization for 12 hours. After centrifugation, crude potassium alum and liquid I were obtained, respectively. Among them, the K in the mixture... + With Al 3+ The molar ratio is 1.02:1; the volume ratio of the leachate to the washing water is 1:0.1; the liquid I is recycled as a leaching agent in step (2);

[0072] (4) Mix the crude potassium alum with water at a ratio of 1g:1mL, heat to 80°C to dissolve, then cool to 5°C and let stand for 12 hours to crystallize. Separate the solid and liquid to obtain refined potassium alum and liquid II respectively; liquid II is reused in step (2);

[0073] (5) Steps (1) to (4) are repeated 3 times. The lithium ion concentration in the liquid I obtained in the last repetition is 3.25 g / L. The liquid I is heated to 40°C, limestone is added to adjust the pH to 5.5, and the reaction is carried out at 40°C with stirring for 1 h. Then, Ca(OH)2 is added to adjust the pH to 10, and the reaction is carried out at 40°C with stirring for 1 h to remove impurities. The mixture is filtered to obtain filtrate and filter residue. The filter residue is washed with water to obtain washing liquid. The filtrate and washing liquid are combined to obtain crude lithium solution. The volume ratio of liquid I to water used for washing is 1:0.2. In the repetition process, the leaching agent in step (2) is a mixed solution of liquid I in step (3), liquid II in step (4) and water. The average leaching rate of Li2O is 90.28% after 3 repetitions.

[0074] (6) ZGC258 macroporous weak acid cation exchange resin is packed into an ion exchange column. The resulting ion exchange resin column is rinsed with deionized water until the effluent is clear. A 2wt% NaOH aqueous solution is passed through the resin layer at a flow rate of 3 m / h. The resin is then rinsed with deionized water until the pH of the effluent is 8, thus obtaining a pretreated resin ion exchange column. The crude lithium solution is then passed through the pretreated resin ion exchange column at a flow rate of 2 BV / h for deep purification and impurity removal, thus obtaining a refined lithium solution. The volume ratio of NaOH aqueous solution to ZGC258 macroporous weak acid cation exchange resin is 3:1.

[0075] (7) The purified lithium solution was concentrated to Li₂ using homogeneous membrane electrodialysis. +With a concentration ≥20 g / L, the obtained Li₂SO₄ concentrate was heated to 97°C, and a 300 g / L Na₂CO₃ aqueous solution was slowly added and mixed thoroughly. The resulting lithium-sodium mixture was stirred to precipitate lithium for 1 hour, followed by solid-liquid separation to obtain lithium precipitation mother liquor and crude lithium carbonate. The crude lithium carbonate was washed twice with deionized water at 97°C and dried to constant weight at 80°C to obtain the lithium carbonate product. The content of each component is shown in Table 2. The Na₂SO₄ in the lithium-sodium mixture... + :Li + The molar ratio was 1.05:1; the solid-liquid ratio of the crude lithium carbonate to the total amount of deionized water used for washing was 1 g: 1 mL; and the Li2O recovery rate was 83.49%.

[0076] Table 2 shows the content of each component in the lithium carbonate product obtained and the YST582-2013 standard product.

[0077]

[0078] As shown in Table 2, the lithium carbonate product prepared by this invention meets the YST582-2013 standard.

[0079] (8) After evaporating and concentrating the lithium precipitation mother liquor to a Na2SO4 mass concentration of 25%, the resulting concentrate is cooled to 5°C for sodium precipitation, centrifuged to separate sodium sulfate product and sodium precipitation mother liquor respectively; the sodium precipitation mother liquor is reused in step (3) to replace carbonate source for adjusting pH value;

[0080] (9) The refined potassium alum was calcined at 170°C for 1 hour to remove crystallization water, and the resulting dehydrated potassium alum was calcined at 800°C for 3 hours to obtain calcined material and calcination flue gas, respectively; the calcined material was soaked in water at 60°C for 1 hour and then filtered to obtain filtrate and crude alumina; the crude alumina was washed with water to obtain washing liquid and alumina product, respectively; the filtrate and washing liquid were combined to obtain water immersion liquid; the solid-liquid ratio of the calcined material to the water for washing was 1 g: 0.1-1 mL; the residual heat of calcination was recycled to steps (1) and (2); wherein the solid-liquid ratio of the calcined material to the water for washing was 2 g: 1 mL; the solid-liquid ratio of the calcined material to the water for washing was 1 g: 0.1-1 mL; the content of each component of the alumina product (dry basis) is shown in Table 3, and the Al2O3 recovery rate is 81.27%.

[0081] Table 3 shows the content of each component in Al2O3 and the YST582-2013 standard product.

[0082]

[0083] As shown in Table 3, the main chemical composition of the Al2O3 product prepared by this invention meets the GB / T24487-2022 standard AO-2 grade.

[0084] (10) After concentrating the aqueous extract to a K2SO4 mass concentration of 17-18%, cool it to a temperature of 5°C and let it stand for crystallization for 12 hours. Then filter it to obtain potassium sulfate product. The potassium sulfate product is recycled in step (3).

[0085] (11) The roasting flue gas and calcination flue gas are absorbed by water or 5% dilute sulfuric acid to obtain sulfuric acid solution and tail gas; when the concentration of the sulfuric acid solution is 98wt%, it is recycled as concentrated sulfuric acid in step (1).

[0086] (12) The exhaust gas is purified by using limestone slurry to obtain purified exhaust gas that meets the GB26132-2010 emission standard.

[0087] Example 2

[0088] (1) The lithium mica concentrate powder (-0.074mm particle size content 60.12wt%) shown in Table 4 and 98wt% concentrated sulfuric acid are thoroughly mixed in a mass ratio of 1:1.1. The mixture is heated from room temperature to 380℃ at a heating rate of 8℃ / min and then kept at the temperature for 3h to obtain roasted material and roasting flue gas respectively. The residual heat obtained from roasting is used in steps (3) and (4).

[0089] Table 4. Multi-element analysis results of lepidolite concentrate powder

[0090] Components <![CDATA[Li2O]]> <![CDATA[Al2O3]]> <![CDATA[SiO2]]> CaO <![CDATA[TFe2O3]]> <![CDATA[K2O]]> <![CDATA[Na2O]]> F Content / wt% 2.39 23.56 37.23 0.28 0.37 7.08 1.33 4.15

[0091] (2) The roasted material is pulverized to a particle size of -0.15mm with a content of ≥50wt% to obtain roasted pulverized material; the roasted pulverized material is mixed with the leaching agent at a ratio of 1g:3mL, and after leaching at 80℃ for 1h under air circulation, it is filtered to obtain filtrate and filter residue respectively; the filter residue is washed with water to obtain washing liquid; the filtrate and washing liquid are combined to obtain leachate; wherein, the leaching agent is water, and the solid-liquid ratio of the roasted pulverized material to the water used for washing is 1g:2mL; the amount of air introduced is based on the complete oxidation of ferrous ions to ferric ions;

[0092] (3) The pH of the leachate was adjusted to 1.8 using carbonate source, heated to 80°C, and potassium sulfate was added to dissolve it. The resulting mixture was subjected to iron reduction treatment at 95°C for 1 hour, followed by filtration to obtain filtrate and filter residue. The filter residue was washed with water to obtain washing liquid. The filtrate and washing liquid were combined and concentrated to 50% of the leachate volume, cooled to 5°C, and allowed to stand for crystallization for 12 hours. After centrifugation, crude potassium alum and liquid I were obtained, respectively. The mixture contained K... + With Al 3+The molar ratio is 1:1; the volume ratio of the leachate to the washing water is 1:0.1; the liquid I is recycled as a leaching agent in step (2);

[0093] (4) Mix the crude potassium alum with water at a ratio of 1g:1mL, heat to 80°C to dissolve, then cool to 5°C and let stand for 12 hours to crystallize. Filter to obtain refined potassium alum and liquid II respectively; liquid II is reused in step (2);

[0094] (5) Steps (1) to (4) are repeated 3 times. The lithium ion concentration in the liquid I obtained in the last repetition is 8.87 g / L. The liquid I is heated to 40°C, limestone is added to adjust the pH to 5.3, and the reaction is carried out at 40°C with stirring for 1 h. Then, Ca(OH)2 is added to adjust the pH and the mixture is stirred at 40°C for 1 h to remove impurities. The mixture is filtered to obtain filtrate and filter residue. The filter residue is washed with water to obtain washing liquid. The filtrate and washing liquid are combined to obtain crude lithium solution. The volume ratio of liquid I to water used for washing is 1:0.5. In the repetition process, the leaching agent in step (2) is a mixed solution of liquid I in step (3), liquid II in step (4) and water. The average leaching rate of Li2O is 94.73% after 3 repetitions.

[0095] (6) ZGC258 macroporous weak acid cation exchange resin is packed into an ion exchange column. The resulting ion exchange resin column is rinsed with deionized water until the effluent is clear. A 2wt% NaOH aqueous solution is passed through the resin layer at a flow rate of 5 m / h. The resin is then rinsed with deionized water until the pH of the effluent is 8, thus obtaining a pretreated resin ion exchange column. The crude lithium solution is then passed through the pretreated resin ion exchange column at a flow rate of 2 BV / h for deep purification and impurity removal, thus obtaining a refined lithium solution. The volume ratio of NaOH aqueous solution to ZGC258 macroporous weak acid cation exchange resin is 3:1.

[0096] (7) The purified lithium solution was concentrated to Li₂ using homogeneous membrane electrodialysis. + With a concentration ≥20 g / L, the obtained Li₂SO₄ concentrate was heated to 97°C, and a 300 g / L Na₂CO₃ aqueous solution was slowly added and mixed thoroughly. The resulting lithium-sodium mixture was stirred to precipitate lithium for 1 hour, followed by solid-liquid separation to obtain lithium precipitation mother liquor and crude lithium carbonate. The crude lithium carbonate was washed twice with deionized water at 97°C and dried to constant weight at 80°C to obtain the lithium carbonate product. The content of each component is shown in Table 2. The Na₂SO₄ in the lithium-sodium mixture... + :Li + The molar ratio was 1.05:1; the solid-liquid ratio of the crude lithium carbonate to the total amount of deionized water used for washing was 1 g: 1 mL; and the Li2O recovery rate was 86.27%.

[0097] Table 5 shows the content of each component in the lithium carbonate product obtained and the YST582-2013 standard product.

[0098]

[0099]

[0100] As shown in Table 5, the lithium carbonate product prepared by this invention meets the YST582-2013 standard.

[0101] (8) After evaporating and concentrating the lithium precipitation mother liquor to a Na2SO4 mass concentration of 25%, the resulting concentrate is cooled to 5°C for sodium precipitation, centrifuged to separate sodium sulfate product and sodium precipitation mother liquor respectively; the sodium precipitation mother liquor is reused in step (3) to replace carbonate source for adjusting pH value;

[0102] (9) The refined potassium alum was calcined at 150°C for 1.5 h to remove crystallization water, and the dehydrated potassium alum was calcined at 850°C for 2 h to obtain calcined material and calcination flue gas, respectively; the calcined material was soaked in water at 60°C for 1 h and then filtered to obtain filtrate and crude alumina; the crude alumina was washed with water to obtain washing liquid and alumina product, respectively; the filtrate and washing liquid were combined to obtain water immersion liquid; the solid-liquid ratio of the calcined material to the water for washing was 1 g: 0.1-1 mL; the residual heat of calcination was recycled to steps (1) and (2); wherein the solid-liquid ratio of the calcined material to the water for washing was 2 g: 1 mL; the solid-liquid ratio of the calcined material to the water for washing was 1 g: 0.2 mL; the content of each component of the alumina product (dry basis) is shown in Table 6, and the Al2O3 recovery rate is 80.77%.

[0103] Table 6 shows the content of each component in Al2O3 and the YST582-2013 standard product.

[0104]

[0105] As shown in Table 6, the main chemical composition of the Al2O3 product prepared by this invention meets the GB / T24487-2022 standard AO-2 grade.

[0106] (10) After concentrating the aqueous extract to a K2SO4 mass concentration of 17-18%, cool it to a temperature of 5°C and let it stand for crystallization for 12 hours. Then filter it to obtain potassium sulfate product. The potassium sulfate product is recycled in step (3).

[0107] (11) The roasting flue gas and calcination flue gas are absorbed by water or 5% dilute sulfuric acid to obtain sulfuric acid solution and tail gas; when the concentration of the sulfuric acid solution is 98wt%, it is recycled as concentrated sulfuric acid in step (1).

[0108] (12) The exhaust gas is purified by using limestone slurry to obtain purified exhaust gas that meets the GB26132-2010 emission standard.

[0109] The above description is only a preferred embodiment of the present invention. It should be noted that for those skilled in the art, several improvements and modifications can be made without departing from the principle of the present invention, and these improvements and modifications should also be considered within the scope of protection of the present invention.

Claims

1. A method for thermal extraction of lithium aluminum acid, characterized in that, Includes the following steps: (1) The lithium-containing material is mixed with concentrated sulfuric acid and roasted to obtain roasted material and roasting flue gas respectively; The residual heat from the roasting is used in steps (3) and (4); the concentrated sulfuric acid has a mass fraction of ≥85%; (2) The calcined material is pulverized to obtain calcined pulverized material; the obtained calcined pulverized material is mixed with a leaching agent, leached, and then filtered to obtain filtrate and filter residue respectively; the filter residue is washed with water to obtain washing liquid; the filtrate and washing liquid are combined to obtain leachate; the leaching agent is water; the leaching is carried out under the condition of passing oxygen-containing gas; the solid-liquid ratio of the calcined pulverized material to the water used for washing is 1g:0.5-2mL; (3) Adjust the pH of the leachate to 1.5-3, add potassium sulfate to dissolve it, and then filter the resulting mixture after iron reduction treatment to obtain filtrate and filter residue. Wash the filter residue with water to obtain washing liquid. Combine the filtrate and washing liquid and concentrate them to 40-60% of the volume of the leachate. After crystallization, separate the solid and liquid to obtain crude potassium alum and liquid I. The pH adjustment uses an alkali including carbonate. The K in the mixture... + With Al 3+ The molar ratio is 0.8 to 1.05:1; the volume ratio of the leachate to the washing water is 1:0.05 to 0.5; the liquid I is recycled as a leaching agent in step (2); (4) Dissolve the crude potassium alum II in water, and perform recrystallization and solid-liquid separation in sequence to obtain refined potassium alum and liquid II respectively; liquid II is reused as a leaching agent in step (2); (5) Repeat steps (1) to (4) n times. When the lithium ion concentration in liquid I is >3g / L, heat liquid I to 40-100℃, adjust the pH value to 5-7 using carbonate source, and neutralize for 0.5-2h. Add impurity removal agent to adjust the pH value to >9 and remove impurities for 0.5-2h. Filter to obtain filtrate and filter residue respectively. Wash the filter residue with water to obtain washing liquid. Combine the filtrate and washing liquid to obtain crude lithium solution. n≥2. The volume ratio of liquid I to water used for washing is 1:0.1-0.

5. In the repeated process, the leaching agent in step (2) is a mixed solution of liquid I in step (3), liquid II in step (4) and water added. (6) The crude lithium solution is purified and impurities are removed by using a resin ion exchange column to obtain a refined lithium solution; (7) The purified lithium solution is concentrated by homogeneous membrane electrodialysis to obtain a concentrated Li₂SO₄ solution; the concentrated Li₂SO₄ solution is mixed with an aqueous Na₂CO₃ solution, and the resulting mixture is subjected to lithium precipitation followed by solid-liquid separation to obtain a lithium precipitation mother liquor and crude lithium carbonate; the crude lithium carbonate is washed with water and dried to obtain a lithium carbonate product; the Li₂SO₄ concentrate contains Li + The concentration of Na in the mixture is ≥20 g / L; + With Li + The molar ratio is 1 to 1.05:1; (8) The lithium precipitation mother liquor is concentrated, sodium is precipitated and solid-liquid separation is performed sequentially to obtain sodium sulfate product and sodium precipitation mother liquor respectively; the sodium precipitation mother liquor is used to replace carbonate source to adjust pH value in step (3) of the next lithium aluminum acid thermal integrated extraction process; (9) The refined potassium alum is successively roasted to remove crystallization water and calcined to obtain calcined material and calcination flue gas respectively; the calcined material is soaked in water and then filtered to obtain filtrate and crude alumina; the crude alumina is washed with water to obtain washing liquid and alumina product respectively; the filtrate and washing liquid are combined to obtain water immersion liquid; the solid-liquid ratio of the calcined material to the water used for washing is 1g:0.1~1mL; the residual heat from calcination is recycled to steps (1) and (2); (10) The aqueous extract is concentrated, crystallized and separated into solid and liquid components in sequence to obtain potassium sulfate product; the potassium sulfate product is recycled in step (3); (11) The roasting flue gas and calcination flue gas are absorbed by water or dilute sulfuric acid to obtain sulfuric acid solution and tail gas respectively; when the mass fraction of the sulfuric acid solution is ≥85%, it is recycled as concentrated sulfuric acid in step (1). There is no chronological order between steps (6) to (8) and steps (9) to (10).

2. The method for integrated thermal extraction of lithium aluminum acid according to claim 1, characterized in that, In step (1), the lithium-containing material includes one or more of lithium-containing clay ore, lepidolite concentrate and spodumene concentrate heat-insulating roasting pretreatment material; The mass ratio of the lithium-containing material to concentrated sulfuric acid is 1:0.2 to 1.2; The roasting includes sequential heating and holding roasting; the heating rate is 5-10℃ / min, and the initial temperature of the heating is room temperature; the holding roasting temperature is 300-550℃, and the holding time is 0.5-6h.

3. The method for integrated thermal extraction of lithium aluminum acid according to claim 1, characterized in that, In step (2), the solid-liquid ratio of the calcined pulverized material to the leaching agent is 1g:2-5mL; the leaching temperature is 40-100℃ and the leaching time is 0.5-2h.

4. The method for integrated thermal extraction of lithium aluminum acid according to claim 1, characterized in that, In step (2), the temperature of the iron reduction treatment is 90-100℃ and the time is 0.5-2h; The carbonate source includes carbonates and / or carbonate minerals.

5. The method for integrated thermal extraction of lithium aluminum acid according to claim 1, characterized in that, In steps (3) and (4), the temperatures of the first dissolution and the second dissolution are independently 40–85°C; The crystallization and recrystallization are carried out at temperatures ranging from 5 to 25°C and at times ranging from 5 to 24 hours, respectively, and both crystallization and recrystallization are carried out by static crystallization.

6. The method for integrated thermal extraction of lithium aluminum acid according to claim 1, characterized in that, In step (5), n is 2 to 10; The carbonate source includes carbonates and / or carbonate minerals.

7. The method for integrated thermal extraction of lithium aluminum acid according to claim 1, characterized in that, In step (6), the resin in the resin ion exchange column includes ZGC258 macroporous weak acid cation exchange resin and / or LSC-500B ion exchange resin.

8. The method for integrated thermal extraction of lithium aluminum acid according to claim 1, characterized in that, In step (7), the temperature of lithium deposition is ≥95℃ and the time is 0.5~1.5h.

9. The method for integrated thermal extraction of lithium aluminum acid according to claim 1, characterized in that, In step (8), the sodium precipitation temperature is 0-5℃.

10. The method for integrated thermal extraction of lithium aluminum acid according to claim 1, characterized in that, In step (9), the temperature for calcination and decrystallization is 150–300°C, and the time is 0.5–3 hours. The calcination temperature is 750–900℃, and the time is 1–5 hours; The water immersion temperature is 20–100℃, and the time is 0.5–2 hours.

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