Method for extracting lithium from complex lithium ore or lithium-containing waste through pretreatment and low-temperature transformation roasting

Abstract

The invention discloses a complex lithium ore or lithium-containing waste pretreatment low-temperature transformation roasting lithium extraction method which comprises the following steps: uniformly mixing one or more of petalite concentrate, spodumene concentrate, lithium chinastone and lithium-containing glass waste with aluminum electrolysis overhaul slag, fluoride and concentrated sulfuric acid, and then carrying out low-temperature calcination pretreatment; pre-treated materials are obtained; carrying out transformation roasting on the pretreated material to obtain a transformed material; then, uniformly mixing the transformed material with concentrated sulfuric acid, and performing low-temperature acidification calcination to obtain an acidified material; and finally, placing the acidified material in water, carrying out water leaching, and carrying out solid-liquid separation to obtain a lithium leachate, according to the technical scheme, collaborative consumption of multiple lithium resources can be achieved, the comprehensive utilization rate of lithium is high, transformation roasting energy consumption is low, the process is simple, cost is low, the method is suitable for industrial application, the important significance is achieved for efficient development and utilization of natural hard rock type lithium ore and secondary lithium resources, and the application and popularization prospects are great.

Description

Technical Field

[0001] The present invention relates to the field of lithium extraction from natural hard rock lithium ore and secondary lithium resources, and in particular to a method for extracting lithium by pre-treating complex lithium ore or lithium-containing waste materials through low-temperature transformation roasting. Background Art

[0002] Lithium is the foundation for the development of my country's new energy industry, and the demand for lithium products such as lithium carbonate and lithium hydroxide is also increasing. Petalite is an important hard rock lithium mineral resource with a chemical composition of LiAlSi4O 10, its theoretical Li2O content is about 4.91%, and it is usually associated with minerals such as spodumene and mica; compared with spodumene (LiAlSi2O6), petalite contains more SiO2 components in its crystal structure, and more chemical additives are usually required to destroy its crystal structure. The efficient utilization of petalite is of great significance to ensure the stable supply of lithium resources in my country and support the development of emerging industries. The methods for extracting lithium from lithium-containing minerals include sulfate method, sulfuric acid method, high-pressure alkali boiling method, chlorination method, etc. For the extraction of lithium from petalite, patent CN118360497A reported a method of roasting sulfate and fluoride salt and then leaching lithium with water, but the amount of leached slag is higher than the lithium ore entering the leaching, which increases the environmental burden, and the roasting time is too long, and the operation efficiency is low. Patent CN116121560B reports a process of roasting and transforming sulfate and sodium aluminosilicate, acidifying with sulfuric acid and then leaching lithium with water, but the roasting temperature of this method is relatively high, greater than 1100°C, and the amount of transformation aid used is relatively large, which is costly. Patent CN118256739A discloses a method of extracting lithium from lithium porcelain stone ore, which can achieve lithium extraction by combining strong acid and weak acid (weak acid plays the role of catalyst) at 70-90°C for ultra-long leaching for 6-12h, but the acid consumption of this process is 4-6 times that of the spodumene sulfuric acid process, and excessive acid consumption will inevitably result in a high leaching rate of impurity elements such as aluminum, increasing the purification burden and cost. Patent CN117602652A discloses a method for extracting lithium from lithium-containing glass waste. The process needs to be carried out in a pressure reactor, and the amount of leaching agent (sodium hydroxide) and crystallization agent (sodium chloride and potassium chloride) is relatively high, about 20% of the mass of the glass waste. The reagent cost is relatively high, and the leachate must be acidified with hydrochloric acid and causticized with sodium hydroxide in turn. The entire reaction process is from alkali to acid and then to alkali, and the reagent consumption is large. Overall, there is no mature process reported for extracting lithium from petalite. The present invention uses the traditional sulfuric acid method to treat it, which can achieve efficient extraction of lithium, but the roasting temperature needs to be above 1100°C, and the energy consumption is relatively high. Although the high-temperature process can be avoided in extracting lithium from lithium porcelain stone and lithium-containing glass waste, the lithium extraction rate is still insufficient. Due to the large amount of acid and alkali used, the amount of impurity ions leached is high, and the purification process is more complicated than the spodumene sulfuric acid method, and the lithium loss rate is higher. Therefore, there is an urgent need to develop a lithium extraction technology from composite lithium resources (petalite, spodumene, lithium porcelain stone and lithium-containing glass waste) that is economically feasible, easy to practice, has a simple leaching solution composition, is easy to purify, and has a high lithium recovery rate. Summary of the invention

[0003] The purpose of the present invention is to overcome the shortcomings of the prior art and provide a method for pre-treating complex lithium ore or lithium-containing waste materials through low-temperature transformation roasting and lithium extraction. The method comprises the following steps: uniformly mixing one or more of petalite concentrate, spodumene concentrate, lithium porcelain stone and lithium-containing glass waste materials with aluminum electrolysis overhaul slag, fluoride and concentrated sulfuric acid, and then performing low-temperature calcination pretreatment to obtain a pretreated material; performing transformation roasting on the pretreated material to obtain a transformation material; then, uniformly mixing the transformation material with concentrated sulfuric acid, performing low-temperature acidification calcination to obtain an acidified material; finally, placing the acidified material in water, performing water immersion and obtaining a lithium leaching solution through solid-liquid separation; the technical scheme of the present invention can realize the coordinated consumption of multiple lithium resources, has a high comprehensive utilization rate of lithium, low energy consumption of transformation roasting, a simple process, low cost, is suitable for industrial application, has important significance for the efficient development and utilization of natural hard rock lithium ore and secondary lithium resources, and has great prospects for application and promotion.

[0004] In order to achieve the above technical effects, the following technical solutions are adopted: A method for extracting lithium by pre-treating complex lithium ore or lithium-containing waste materials through low-temperature transformation roasting, comprising the following steps: Step S1: uniformly mixing the lithium-containing composite material with aluminum electrolysis overhaul slag, fluoride and concentrated sulfuric acid and then performing low-temperature calcination pretreatment to obtain a pretreated material; the lithium-containing composite material is one or more of petalite concentrate, spodumene concentrate, lithium porcelain stone and lithium-containing glass waste; Step S2: subjecting the pretreated material in step S1 to transformation roasting to obtain transformation material; Step S3: uniformly mixing the transformation material in step S2 with concentrated sulfuric acid, and performing low-temperature acidification and calcination to obtain an acidified material; Step S4: placing the acidified material in step S3 in water for water immersion, and then performing solid-liquid separation to obtain a lithium leaching solution, thereby achieving efficient lithium extraction.

[0005] Furthermore, the lithium oxide content of the petalite concentrate is 1.0~4.5wt%; the lithium oxide content of the spodumene concentrate is 3.0~6.5wt%; the lithium oxide content of the lithium porcelain stone is 0.5~4.0wt%; the lithium oxide content of the lithium-containing glass waste is 1.0~6.0wt%; and the lithium oxide content of the aluminum electrolysis overhaul slag is 0.5~3.5wt%.

[0006] Furthermore, the F content of the aluminum electrolysis overhaul slag is 3.0-15.0wt%, and the fluoride includes calcium fluoride.

[0007] Furthermore, the mass ratio of the lithium-containing composite material, the aluminum electrolysis overhaul slag and the fluoride in step S1 is 5-40:1-2:0.5-1; the molar amount of concentrated sulfuric acid added in step S1 is 1 to 4 times the molar amount of F in the aluminum electrolysis overhaul slag and the fluoride.

[0008] The dosage of aluminum electrolysis overhaul slag is too low, which is not enough to continuously support the transformation of composite lithium ore during the pretreatment and transformation roasting process. Excessive dosage will cause the composite lithium ore to be converted into a dense lithium sodium aluminum silicate solid solution phase during the transformation roasting process, which is difficult to achieve ion exchange with sulfuric acid and has a low lithium extraction rate.

[0009] Furthermore, in the step S1, the low-temperature calcination pretreatment temperature is 200-320° C.; and the low-temperature activation calcination time in the step S1 is 30-60 min.

[0010] Furthermore, the transition calcination temperature in step S2 is 800-950° C.; the transition calcination time in step S2 is 20-100 min.

[0011] Only by controlling the transformation roasting temperature and time within a reasonable range can a high lithium extraction rate be guaranteed. When the roasting temperature is too low or the time is too short, the synergistic transformation of lithium-containing composite materials such as petalite concentrate, spodumene concentrate, lithium porcelain stone, lithium-containing glass waste, aluminum electrolysis overhaul slag and fluoride cannot be completed. If the roasting temperature is too high or the time is too long, the roasting energy consumption will increase.

[0012] Furthermore, in step S3, the amount of sulfuric acid mixed with the conversion material is 1.2 to 1.4 times the theoretical acid consumption when all the lithium elements in the lithium-containing composite material are converted into lithium sulfate, and the sulfuric acid is concentrated sulfuric acid.

[0013] If the amount of sulfuric acid used is too low, the lithium replacement rate is low, and water leaching cannot achieve full lithium extraction; if the amount of sulfuric acid used is too high, the impurity leaching rate is high, the purification burden is heavy, the lithium inclusion loss rate is high, equipment corrosion is aggravated, and the economy is insufficient.

[0014] Furthermore, the temperature of the low-temperature acidification and calcination in step S3 is 150-300° C.; the time of the low-temperature acidification and calcination in step S3 is 10-60 min.

[0015] Controlling the acidification reaction temperature and time within a reasonable range can improve the Li + With H + To ensure a high lithium extraction rate, the replacement efficiency must be improved. Too high an acidification reaction temperature or too long a reaction time will increase energy consumption.

[0016] Furthermore, the water immersion reaction conditions in step S4 are: liquid-solid ratio of 1-5 mL:1 g, temperature of 25-90° C., and time of 5-60 min.

[0017] Too low liquid-to-solid ratio, leaching temperature and too short leaching time cannot ensure the full dissolution of lithium sulfate, and the lithium extraction rate is insufficient; too high liquid-to-solid ratio will increase the subsequent evaporation and concentration costs; too high leaching temperature will increase energy consumption costs; too long leaching time will reduce production efficiency.

[0018] The beneficial effects of the present invention are: (1) Coordinated consumption of multiple lithium resources. Lithium can be extracted from lithium-containing primary minerals such as petalite, spodumene, and lithium porcelain stone, as well as lithium-containing secondary resources such as lithium-containing glass waste and aluminum electrolysis overhaul slag.

[0019] (2) High comprehensive utilization rate of lithium and low energy consumption for transition roasting. The traditional sulfuric acid method requires a high-temperature transition roasting process of 1100°C to process petalite, spodumene, etc. alone, and the energy consumption is more than half of the entire lithium salt production process. The lithium extraction rate of lithium porcelain stone and lithium-containing secondary resources such as lithium-containing glass waste and aluminum electrolysis overhaul slag alone is insufficient for sulfuric acid leaching. Based on the low-temperature activation of sulfuric acid, fluoride activation calcination is used to achieve pre-attack and attachment of complex minerals and raw material structures. During the transition roasting, fluoride first quickly removes the surface of complex ores and difficult-to-attack sites. The activated aluminum electrolysis overhaul slag continuously supplies a small amount of fluorine to continuously attack the complex ores and raw materials that have been pre-attacked by fluoride, and disintegrate the structure of complex ores and raw materials; lithium-containing composite materials such as lithium feldspar concentrate, spodumene concentrate, lithium porcelain stone and lithium-containing glass waste are transformed with aluminum electrolysis overhaul slag and fluoride at 800~950°C to achieve synergistic transformation of mineral phases, and combined with subsequent low-temperature acid calcination, Li + With H + The lithium extraction rate is significantly improved in the water leaching stage, and the lithium extraction rate is maintained at more than 96%. The energy consumption of the present invention is better than the traditional sulfuric acid method of single treatment of petalite concentrate, spodumene concentrate and lithium porcelain stone, and the lithium extraction rate is also better than the single treatment process of petalite concentrate, spodumene concentrate, lithium porcelain stone and lithium-containing glass waste and aluminum electrolysis overhaul slag.

[0020] (3) The scheme of the present invention is safe and easy to implement, and can be used as a basis for technical innovation and industrial application based on the traditional sulfuric acid method in industry. BRIEF DESCRIPTION OF THE DRAWINGS

[0021] In order to more clearly illustrate the embodiments of the present invention or the technical solutions in the prior art, the drawings required for use in the embodiments or the description of the prior art will be briefly introduced below. The drawings in the following description are merely embodiments of the present invention. For ordinary technicians in this field, other drawings can be obtained based on the provided drawings without paying any creative work.

[0022] Figure 1 It is a process schematic diagram of a method for pre-treating complex lithium ore or lithium-containing waste by low-temperature transformation roasting and lithium extraction provided by an embodiment of the present invention. DETAILED DESCRIPTION

[0023] In order to make the purpose, technical solution and advantages of the present invention more clearly understood, the present invention is further described in detail below in conjunction with the embodiments. It should be understood that the specific embodiments described herein are only used to explain the present invention and are not used to limit the present invention.

[0024] It should be noted that the following detailed descriptions are exemplary and are intended to provide further explanation of the present invention. Unless otherwise specified, all technical and scientific terms used herein have the same meanings as those commonly understood by those skilled in the art to which the present invention belongs.

[0025] It should be noted that the terms used herein are only for describing specific embodiments and are not intended to limit exemplary embodiments according to the present invention. As used herein, unless the context clearly indicates otherwise, the singular form is also intended to include the plural form. In addition, it should be understood that when the terms "comprising" and / or "including" are used in this specification, it indicates the presence of features, steps, operations and / or combinations thereof.

[0026] Embodiment 1: according to Figure 1 As shown, the lithium extraction operation of petalite concentrate is carried out: Take 50g of petalite concentrate with a lithium oxide content of 3.45wt.%, 10g of aluminum electrolysis overhaul slag with a fluorine content of 7.08% (lithium oxide content of 1.5wt.%), 5g of calcium fluoride and 0.331 mol of concentrated sulfuric acid (i.e., 2 times the molar amount of F in the composite raw material) and mix them evenly, and then place them in a muffle furnace for low-temperature sulfuric acid activation calcination, the calcination temperature is set to 300°C, and the calcination time is 40min. Then continue to heat up for transformation roasting, the roasting temperature is set to 900°C, the roasting time is set to 60min, and after the roasting material is cooled to room temperature, it is ball-milled to less than 200 mesh, and then concentrated sulfuric acid is added according to 1.3 times the theoretical acid consumption and evenly mixed, and the mixed acid material is placed in a muffle furnace for low-temperature acidification calcination, the calcination temperature is set to 200°C, and the calcination time is 20min to obtain an acidified material. Then, the acidified material was placed in water for leaching at a liquid-to-solid ratio of 2 mL:1 g, and the leaching temperature was set to 30° C. for 20 min. After leaching, the lithium leaching solution and lithium leaching tailings were obtained by filtration.

[0027] After testing and analysis, the leaching rate of Li in the leaching solution obtained in Example 1 was 97.44%, indicating that the technical solution can extract lithium from the composite lithium resources of petalite concentrate and aluminum electrolysis overhaul slag at a lower energy consumption (900°C).

[0028] Embodiment 2: according to Figure 1 As shown, the lithium extraction operation of spodumene concentrate is carried out: Take 40g of spodumene concentrate with a lithium oxide content of 5.00wt.%, 1g of aluminum electrolysis overhaul slag with a fluorine content of 7.08% (lithium oxide content of 1.5wt.%), 0.5g of calcium fluoride and 0.0662 mol of concentrated sulfuric acid (i.e. 4 times the molar amount of F in the composite raw material) and mix them evenly, and then place them in a muffle furnace for low-temperature sulfuric acid activation calcination, the calcination temperature is set to 200°C, and the calcination time is 60min. Then continue to heat up for transformation roasting, the roasting temperature is set to 950°C, the roasting time is set to 100min, and after the roasting material is cooled to room temperature, it is ball-milled to less than 200 mesh, and then concentrated sulfuric acid is added according to 1.4 times the theoretical acid consumption and evenly mixed, and the mixed acid material is placed in a muffle furnace for low-temperature acidification calcination, the calcination temperature is set to 300°C, and the calcination time is 40min to obtain an acidified material. Then, the acidified material was placed in water for leaching at a liquid-to-solid ratio of 2 mL:1 g, and the leaching temperature was set to 30° C. for 20 min. After leaching, the lithium leaching solution and lithium leaching tailings were obtained by filtration.

[0029] After testing and analysis, the leaching rate of Li in the leaching solution obtained in Example 1 was 96.99%, indicating that the technical solution can extract lithium from spodumene concentrate and aluminum electrolysis overhaul slag composite lithium resources at a relatively low energy consumption (900°C).

[0030] Embodiment 3: according to Figure 1 As shown, the lithium extraction operation of lithium porcelain stone concentrate is carried out: Take 40g of lithium porcelain stone concentrate with a lithium oxide content of 1.00wt.%, 2g of aluminum electrolysis overhaul slag with a fluorine content of 7.08% (lithium oxide content of 1.5wt.%), 1g of calcium fluoride and 0.132 mol of concentrated sulfuric acid (i.e. 4 times the molar amount of F in the composite raw material) and mix them evenly, then place them in a muffle furnace for low-temperature sulfuric acid activation calcination, the calcination temperature is set to 300°C, and the calcination time is 40min. Then continue to heat up for transformation roasting, the roasting temperature is set to 800°C, the roasting time is set to 60min, after the roasting material is cooled to room temperature, it is ball-milled to less than 200 mesh, and then concentrated sulfuric acid is added according to 1.2 times the theoretical acid consumption and evenly mixed, and the mixed acid material is placed in a muffle furnace for low-temperature acidification calcination, the calcination temperature is set to 200°C, and the calcination time is 60min to obtain an acidified material. Then, the acidified material was placed in water for leaching at a liquid-to-solid ratio of 3 mL:1 g, and the leaching temperature was set to 40° C. for 40 min. After leaching, the lithium leaching solution and lithium leaching tailings were obtained by filtration.

[0031] After testing and analysis, the leaching rate of Li in the leaching solution obtained in Example 1 was 97.49%, indicating that the technical solution can extract lithium from lithium porcelain stone concentrate and aluminum electrolysis overhaul slag composite lithium resources at a lower energy consumption (900°C).

[0032] Embodiment 4: according to Figure 1 As shown, the lithium extraction operation of lithium-containing glass waste is carried out: Take 40g of lithium-containing glass waste with a lithium oxide content of 1.00wt.%, 2g of aluminum electrolysis overhaul slag with a fluorine content of 7.08% (lithium oxide content of 1.5wt.%), 1g of calcium fluoride and 0.132 mol of concentrated sulfuric acid (i.e. 4 times the molar amount of F in the composite raw material) and mix them evenly, and then place them in a muffle furnace for low-temperature sulfuric acid activation calcination, the calcination temperature is set to 320°C, and the calcination time is 30min. Then continue to heat up for transformation roasting, the roasting temperature is set to 900°C, the roasting time is set to 80min, after the roasting material is cooled to room temperature, ball milled to less than 200 mesh, and then concentrated sulfuric acid is added according to 1.3 times the theoretical acid consumption and evenly mixed, the mixed acid material is placed in a muffle furnace for low-temperature acidification calcination, the calcination temperature is set to 200°C, and the calcination time is 20min, to obtain an acidified material. Then, the acidified material was placed in water for leaching at a liquid-to-solid ratio of 4 mL:1 g, and the leaching temperature was set to 30 ° C for 30 min. After leaching, the lithium leaching solution and lithium leaching tailings were obtained by filtration.

[0033] After testing and analysis, the leaching rate of Li in the leaching solution obtained in Example 1 was 98.17%, indicating that the technical solution can extract lithium from lithium-containing glass waste and aluminum electrolysis overhaul slag composite lithium resources at a lower energy consumption (900°C).

[0034] Embodiment 5: according to Figure 1 As shown, the lithium extraction operation of petalite concentrate and lithium porcelain stone concentrate is carried out: Take 20g of petalite concentrate with a lithium oxide content of 3.45wt.%, take 20g of lithium porcelain stone concentrate with a lithium oxide content of 1.00wt.%, 2g of aluminum electrolysis overhaul slag with a fluorine content of 7.08% (lithium oxide content of 1.5wt.%), 1g of calcium fluoride and 0.132 mol of concentrated sulfuric acid (i.e. 4 times the molar amount of F in the composite raw material) and mix them evenly, and then place them in a muffle furnace for low-temperature sulfuric acid activation calcination, the calcination temperature is set to 300°C, and the calcination time is 40min. Then continue to heat up for transformation roasting, the roasting temperature is set to 900°C, the roasting time is set to 60min, after the roasting material is cooled to room temperature, ball mill to less than 200 mesh, then add concentrated sulfuric acid according to 1.4 times the theoretical acid consumption and mix evenly, the mixed acid material is placed in a muffle furnace for low-temperature acidification calcination, the calcination temperature is set to 150°C, and the calcination time is 60min, to obtain an acidified material. Then, the acidified material was placed in water for leaching at a liquid-to-solid ratio of 5 mL:1 g, and the leaching temperature was set to 80° C. for 40 min. After leaching, the lithium leaching solution and lithium leaching tailings were obtained by filtration.

[0035] After testing and analysis, the leaching rate of Li in the leaching solution obtained in Example 1 was 96.47%, indicating that the technical solution can extract lithium from the lithium feldspar concentrate, lithium porcelain stone concentrate and aluminum electrolysis overhaul slag composite lithium resources at a relatively low energy consumption (900°C).

[0036] Embodiment 6: according to Figure 1 As shown, the lithium extraction operation of petalite concentrate and lithium-containing glass waste is carried out: Take 20g of petalite concentrate with a lithium oxide content of 3.45wt.%, take 20g of lithium-containing glass waste with a lithium oxide content of 1.00wt.%, 2g of aluminum electrolysis overhaul slag with a fluorine content of 7.08% (lithium oxide content of 1.5wt.%), 1g of calcium fluoride and 0.132 mol of concentrated sulfuric acid (i.e. 4 times the molar amount of F in the composite raw material) and mix them evenly, and then place them in a muffle furnace for low-temperature sulfuric acid activation calcination, the calcination temperature is set to 300°C, and the calcination time is 60min. Then continue to heat up for transformation roasting, the roasting temperature is set to 900°C, the roasting time is set to 80min, after the roasting material is cooled to room temperature, it is ball-milled to less than 200 mesh, and then concentrated sulfuric acid is added according to 1.3 times the theoretical acid consumption and evenly mixed, and the mixed acid material is placed in a muffle furnace for low-temperature acidification calcination, the calcination temperature is set to 200°C, and the calcination time is 60min to obtain an acidified material. Then, the acidified material was placed in water for leaching at a liquid-to-solid ratio of 1 mL:1 g, and the leaching temperature was set at 30° C. for 60 min. After leaching, the lithium leaching solution and lithium leaching tailings were obtained by filtration.

[0037] After testing and analysis, the leaching rate of Li in the leaching solution obtained in Example 1 is 98.69%, indicating that the technical solution can extract lithium from the lithium resources of petalite concentrate, lithium-containing glass waste and aluminum electrolysis overhaul slag at a relatively low energy consumption (900°C).

[0038] Embodiment 7: according to Figure 1 As shown, lithium extraction operation of spodumene concentrate and lithium-containing glass waste is carried out: Take 10g of spodumene concentrate with a lithium oxide content of 5.00wt.%, take 30g of lithium-containing glass waste with a lithium oxide content of 1.00wt.%, 2g of aluminum electrolysis overhaul slag with a fluorine content of 7.08% (lithium oxide content of 1.5wt.%), 1g of calcium fluoride and 0.132 mol of concentrated sulfuric acid (i.e. 4 times the molar amount of F in the composite raw material) and mix them evenly, and then place them in a muffle furnace for low-temperature sulfuric acid activation calcination, the calcination temperature is set to 300°C, and the calcination time is 40min. Then continue to heat up for transformation roasting, the roasting temperature is set to 900°C, the roasting time is set to 60min, after the roasting material is cooled to room temperature, it is ball-milled to less than 200 mesh, and then concentrated sulfuric acid is added according to 1.3 times the theoretical acid consumption and evenly mixed, and the mixed acid material is placed in a muffle furnace for low-temperature acidification calcination, the calcination temperature is set to 200°C, and the calcination time is 40min to obtain an acidified material. Then, the acidified material was placed in water for leaching at a liquid-to-solid ratio of 3 mL:1 g, and the leaching temperature was set at 90° C. for 40 min. After leaching, the lithium leaching solution and lithium leaching tailings were obtained by filtration.

[0039] After testing and analysis, the leaching rate of Li in the leaching solution obtained in Example 1 is 98.37%, indicating that the technical solution can extract lithium from spodumene concentrate, lithium-containing glass waste and aluminum electrolysis overhaul slag composite lithium resources at a relatively low energy consumption (900°C).

[0040] Embodiment 8: according to Figure 1 As shown, lithium extraction operations are carried out on lithium porcelain stone concentrate and lithium-containing glass waste: Take 10g of lithium porcelain stone concentrate with a lithium oxide content of 1.00wt.%, take 30g of lithium-containing glass waste with a lithium oxide content of 1.00wt.%, take 2g of aluminum electrolysis overhaul slag with a fluorine content of 7.08% (lithium oxide content of 1.5wt.%), 1g of calcium fluoride and 0.132 mol of concentrated sulfuric acid (i.e. 4 times the molar amount of F in the composite raw material) and mix them evenly, and then place them in a muffle furnace for low-temperature sulfuric acid activation calcination, the calcination temperature is set to 300°C, and the calcination time is 40min. Then continue to heat up for transformation roasting, the roasting temperature is set to 950°C, the roasting time is set to 60min, after the roasting material is cooled to room temperature, ball mill to less than 200 mesh, then add concentrated sulfuric acid according to 1.3 times the theoretical acid consumption and mix evenly, the mixed acid material is placed in a muffle furnace for low-temperature acidification calcination, the calcination temperature is set to 300°C, and the calcination time is 20min, to obtain an acidified material. Then, the acidified material was placed in water for leaching at a liquid-to-solid ratio of 2 mL:1 g, and the leaching temperature was set to 25° C. for 10 min. After leaching, the lithium leaching solution and lithium leaching tailings were obtained by filtration.

[0041] After testing and analysis, the leaching rate of Li in the leaching solution obtained in Example 1 is 96.18%, indicating that the technical solution can extract lithium from lithium porcelain stone concentrate, lithium-containing glass waste and aluminum electrolysis overhaul slag composite lithium resources at a relatively low energy consumption (900°C).

[0042] Embodiment 9: according to Figure 1 As shown, lithium extraction operations are carried out on petalite concentrate, lithium porcelain stone concentrate and lithium-containing glass waste: 10 g of petalite concentrate with a lithium oxide content of 3.45 wt.%, 10 g of lithium porcelain stone concentrate with a lithium oxide content of 1.00 wt.%, 20 g of lithium-containing glass waste with a lithium oxide content of 1.00 wt.%, 2 g of aluminum electrolysis overhaul slag with a fluorine content of 7.08% (lithium oxide content of 1.5 wt.%), 1 g of calcium fluoride and 0.132 mol of concentrated sulfuric acid (i.e., 4 times the molar amount of F in the composite raw material) are uniformly mixed, and then placed in a muffle furnace for low-temperature sulfuric acid activation calcination, the calcination temperature is set to 320°C, and the calcination time is 60 min. Then continue to heat up and carry out transition roasting, roasting temperature is set at 950 DEG C, roasting time is set at 60min, after roasting material is cooled to normal temperature, ball mill is to below 200 mesh, then according to 1.4 times of theoretical acid consumption, concentrated sulfuric acid is added and uniformly mixed, mixed acid material is placed in muffle furnace and carries out low temperature acidification calcination, calcining temperature is set at 300 DEG C, calcining time is 20min, obtains acidifying material.It is subsequently that 3mL:10g is placed in water and carries out water immersion according to liquid-solid ratio, leaching temperature is set at 30 DEG C, and time is 20min.After leaching, filter and obtain lithium leachate and lithium leaching tailings.

[0043] After testing and analysis, the leaching rate of Li in the leachate obtained in Example 1 is 97.88%, indicating that the technical solution can extract lithium from lithium feldspar concentrate, lithium porcelain stone concentrate, lithium-containing glass waste and aluminum electrolysis overhaul slag composite lithium resources at a relatively low energy consumption (900°C).

[0044] Embodiment 10: according to Figure 1 As shown, lithium extraction operations are carried out on petalite concentrate, spodumene concentrate, lithium porcelain stone concentrate and lithium-containing glass waste: 10 g of petalite concentrate with a lithium oxide content of 3.45 wt.%, 10 g of spodumene concentrate with a lithium oxide content of 5.00 wt.%, 10 g of lithium porcelain stone concentrate with a lithium oxide content of 1.00 wt.%, 10 g of lithium-containing glass waste with a lithium oxide content of 1.00 wt.%, 2 g of aluminum electrolysis overhaul slag with a fluorine content of 7.08% (lithium oxide content of 1.5 wt.%), 1 g of calcium fluoride and 0.132 mol of concentrated sulfuric acid (i.e., 4 times the molar amount of F in the composite raw material) are uniformly mixed, and then placed in a muffle furnace for low-temperature sulfuric acid activation calcination, the calcination temperature is set to 300°C, and the calcination time is 60 min. Then continue to heat up and carry out transition roasting, roasting temperature is set at 950 DEG C, roasting time is set at 100min, after roasting material is cooled to normal temperature, ball mill is to below 200 mesh, then according to 1.3 times of theoretical acid consumption, concentrated sulfuric acid is added and uniformly mixed, mixed acid material is placed in muffle furnace and carries out low temperature acidification calcination, calcining temperature is set at 300 DEG C, calcining time is 50min, obtains acidifying material.It is subsequently that 4mL:10g is placed in water and carries out water immersion according to liquid-solid ratio, leaching temperature is set at 50 DEG C, and time is 50min.After leaching, filter and obtain lithium leachate and lithium leaching tailings.

[0045] After testing and analysis, the leaching rate of Li in the leachate obtained in Example 1 is 97.53%, indicating that the technical solution can extract lithium from petalite concentrate, spodumene concentrate, lithium porcelain stone concentrate, lithium-containing glass waste and aluminum electrolysis overhaul slag composite lithium resources at a relatively low energy consumption (900°C).

[0046] Comparative Example 1: Taking Example 1 as a benchmark, the first step of low-temperature sulfuric acid activation calcination was removed, and the others remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 1 was 62.37%.

[0047] Comparative Example 2: Taking Example 1 as a benchmark, calcium fluoride was not added in the first step of low-temperature sulfuric acid activation calcination, and calcium fluoride was added in the second step of transition roasting, while other factors remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 2 was 68.46%.

[0048] Comparative Example 3: Taking Example 1 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, aluminum electrolysis overhaul slag is not added, and in the second step of transition roasting, aluminum electrolysis overhaul slag is added, and the other remains unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 3 is 69.86%.

[0049] Comparative Example 4: Taking Example 1 as a benchmark, petalite concentrate was not added in the first step of low-temperature sulfuric acid activation calcination, and petalite concentrate was added in the second step of transition roasting, and the others remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 4 was 64.55%.

[0050] Comparative Example 5: Taking Example 1 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, aluminum electrolysis overhaul slag and calcium fluoride are not added, and in the second step of transition roasting, aluminum electrolysis overhaul slag and calcium fluoride are added, and the other conditions remain unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 5 is 62.41%.

[0051] Comparative Example 6: Taking Example 1 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, neither petalite concentrate nor calcium fluoride was added, and in the second step of transition roasting, petalite concentrate and calcium fluoride were added, and the others remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 6 was 64.95%.

[0052] Comparative Example 7: Taking Example 1 as a benchmark, the third step of low-temperature acidification and calcination was omitted, and water leaching was performed directly, while other conditions remained unchanged. After testing and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 7 was 52.48%.

[0053] Comparative Example 8: Based on Example 1, the first step of low-temperature sulfuric acid activation calcination was not performed, and the same concentrated sulfuric acid was used for immersion instead, and the temperature was not increased for activation calcination. Other factors remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 8 was 63.89%.

[0054] Comparative Example 9: Based on Example 1, the first step of low-temperature sulfuric acid activation calcination was not performed. Instead, the raw materials were pre-grinded in a ball mill for 60 minutes and then subjected to transition roasting. Other factors remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 9 was 63.11%.

[0055] Comparative Example 10: Based on Example 1, the first step of low-temperature sulfuric acid activation calcination was not performed, and the raw materials were mixed with the same equivalent of sulfuric acid and ultrasonically treated for 30 minutes. The other conditions remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 10 was 64.67%.

[0056] Comparative Example 11: Taking Example 1 as a benchmark, the raw material petalite concentrate was removed, and the others remained unchanged. After testing and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 11 was 72.53%. When the aluminum electrolysis overhaul slag was treated alone, the lithium extraction rate was lower than that of the composite lithium resource. This is mainly because the aluminum electrolysis overhaul slag alone is more easily melted to form a dense phase glass during the transformation roasting; while when the composite lithium resource is synergistically treated, the amount of aluminum electrolysis overhaul slag added is relatively low (mainly serving as a fluorine source), which can be loaded on the surface of petalite and spodumene to prevent the appearance of an amorphous dense phase.

[0057] Comparative Example 12: Taking Example 3 as a benchmark, the first step of low-temperature sulfuric acid activation calcination was removed, and the others remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 12 was 64.97%.

[0058] Comparative Example 13: Taking Example 3 as a benchmark, calcium fluoride was not added in the first step of low-temperature sulfuric acid activation calcination, and calcium fluoride was added in the second step of transition roasting, while other factors remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 13 was 69.33%.

[0059] Comparative Example 14: Taking Example 3 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, aluminum electrolysis overhaul slag is not added, and in the second step of transition roasting, aluminum electrolysis overhaul slag is added, and the other remains unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 14 is 69.24%.

[0060] Comparative Example 15: Taking Example 3 as a benchmark, lithium porcelain stone concentrate was not added in the first step of low-temperature sulfuric acid activation calcination, and lithium porcelain stone concentrate was added in the second step of transition roasting, while the others remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 15 was 65.39%.

[0061] Comparative Example 16: Taking Example 3 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, aluminum electrolysis overhaul slag and calcium fluoride are not added, and in the second step of transition roasting, aluminum electrolysis overhaul slag and calcium fluoride are added, and the other conditions remain unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 16 is 65.47%.

[0062] Comparative Example 17: Taking Example 3 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, lithium porcelain stone concentrate and calcium fluoride are not added, lithium porcelain stone and calcium fluoride are added in the second step of transition roasting, and the others remain unchanged. After detection and analysis, the leaching rate of Li in the leachate obtained in Comparative Example 17 is 66.39%.

[0063] Comparative Example 18: Taking Example 3 as a benchmark, the third step of low-temperature acid calcination was not performed, and water leaching was performed directly, while other conditions remained unchanged. After testing and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 18 was 54.89%.

[0064] Comparative Example 19: Based on Example 3, the first step of low-temperature sulfuric acid activation calcination was not performed, and the same concentrated sulfuric acid immersion was used instead, and the temperature was not increased for activation calcination. Other factors remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 19 was 65.32%.

[0065] Comparative Example 20: Based on Example 3, the first step of low-temperature sulfuric acid activation calcination is not performed. Instead, the raw materials are pre-grinded in a ball mill for 60 minutes and then subjected to transition roasting. Others remain unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 20 is 65.22%.

[0066] Comparative Example 21: Based on Example 3, the first step of low-temperature sulfuric acid activation calcination was not performed. Instead, the raw materials were mixed with the same equivalent of sulfuric acid and subjected to ultrasonic treatment for 30 minutes. The other parameters remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 21 was 66.18%.

[0067] Comparative Example 22: Taking Example 4 as a benchmark, the first step of low-temperature sulfuric acid activation calcination was removed, and the others remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 22 was 60.36%.

[0068] Comparative Example 23: Taking Example 4 as a benchmark, calcium fluoride was not added in the first step of low-temperature sulfuric acid activation calcination, and calcium fluoride was added in the second step of transition roasting, while other factors remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 23 was 64.67%.

[0069] Comparative Example 24: Taking Example 4 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, aluminum electrolysis overhaul slag is not added, and in the second step of transition roasting, aluminum electrolysis overhaul slag is added, and the other remains unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 24 is 66.74%.

[0070] Comparative Example 25: Taking Example 4 as a benchmark, lithium-containing glass waste is not added in the first step of low-temperature sulfuric acid activation calcination, lithium-containing glass waste is added in the second step of transition roasting, and the others remain unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 25 is 62.59%.

[0071] Comparative Example 26: Taking Example 4 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, aluminum electrolysis overhaul slag and calcium fluoride were not added, and in the second step of transition roasting, aluminum electrolysis overhaul slag and calcium fluoride were added, and the other factors remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 26 was 60.68%.

[0072] Comparative Example 27: Taking Example 4 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, lithium-containing glass waste and calcium fluoride are not added, and lithium-containing glass waste and calcium fluoride are added in the second step of transition roasting, and the other conditions remain unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 27 is 62.39%.

[0073] Comparative Example 28: Taking Example 4 as a benchmark, the third step of low-temperature acid calcination was not performed, and water leaching was performed directly, while other conditions remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 28 was 48.67%.

[0074] Comparative Example 29: Based on Example 4, the first step of low-temperature sulfuric acid activation calcination was not performed, and the same concentrated sulfuric acid immersion was used instead, and the temperature was not increased for activation calcination. Other factors remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 29 was 61.29%.

[0075] Comparative Example 30: Based on Example 4, the first step of low-temperature sulfuric acid activation calcination is not performed. Instead, the raw materials are pre-grinded in a ball mill for 60 minutes and then subjected to transition roasting. Other factors remain unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 30 is 62.37%.

[0076] Comparative Example 31: Based on Example 4, the first step of low-temperature sulfuric acid activation calcination was not performed. Instead, the raw materials were mixed with the same equivalent of sulfuric acid and subjected to ultrasonic treatment for 30 minutes. The other parameters remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 31 was 62.83%.

[0077] Comparative Example 32: Taking Example 9 as a benchmark, the first step of low-temperature sulfuric acid activation calcination was removed, and the others remained unchanged. After testing and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 32 was 58.64%.

[0078] Comparative Example 33: Taking Example 9 as a benchmark, calcium fluoride was not added in the first step of low-temperature sulfuric acid activation calcination, and calcium fluoride was added in the second step of transition roasting, while other factors remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 33 was 62.33%.

[0079] Comparative Example 34: Taking Example 9 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, aluminum electrolysis overhaul slag is not added, and in the second step of transition roasting, aluminum electrolysis overhaul slag is added, and the other parts remain unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 34 is 61.59%.

[0080] Comparative Example 35: Taking Example 9 as a benchmark, petalite concentrate was not added in the first step of low-temperature sulfuric acid activation calcination, and petalite concentrate was added in the second step of transition roasting, and the others remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 35 was 59.66%.

[0081] Comparative Example 36: Taking Example 9 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, aluminum electrolysis overhaul slag and calcium fluoride were not added, and in the second step of transition roasting, aluminum electrolysis overhaul slag and calcium fluoride were added, and the other factors remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 36 was 58.61%.

[0082] Comparative Example 37: Taking Example 9 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, neither petalite concentrate nor calcium fluoride was added, and in the second step of transition roasting, petalite concentrate and calcium fluoride were added, and the others remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 37 was 60.27%.

[0083] Comparative Example 38: Taking Example 9 as a benchmark, the third step of low-temperature acidification and calcination is omitted, and water leaching is performed directly, while other conditions remain unchanged. After testing and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 38 is 44.36%.

[0084] Comparative Example 39: Taking Example 9 as a benchmark, the first step of low-temperature sulfuric acid activation calcination is not performed, and the same concentrated sulfuric acid immersion is used instead, and the temperature is not increased for activation calcination. Other things remain unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 39 is 58.93%.

[0085] Comparative Example 40: Based on Example 9, the first step of low-temperature sulfuric acid activation calcination was not performed. Instead, the raw materials were pre-grinded in a ball mill for 60 minutes and then subjected to transition roasting. Other factors remained unchanged. After testing and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 40 was 59.66%.

[0086] Comparative Example 41: Based on Example 9, the first step of low-temperature sulfuric acid activation calcination was not performed. Instead, the raw materials were mixed with the same equivalent of sulfuric acid and subjected to ultrasonic treatment for 30 minutes. The other parameters remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 41 was 60.18%.

[0087] Comparative Example 42: Taking Example 10 as a benchmark, the first step of low-temperature sulfuric acid activation calcination was removed, and the others remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 42 was 52.34%.

[0088] Comparative Example 43: Taking Example 10 as a benchmark, calcium fluoride was not added in the first step of low-temperature sulfuric acid activation calcination, and calcium fluoride was added in the second step of transition roasting, while other factors remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 43 was 59.64%.

[0089] Comparative Example 44: Taking Example 10 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, aluminum electrolysis overhaul slag is not added, and in the second step of transition roasting, aluminum electrolysis overhaul slag is added, and the other parts remain unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 44 is 57.46%.

[0090] Comparative Example 45: Taking Example 10 as a benchmark, petalite concentrate was not added in the first step of low-temperature sulfuric acid activation calcination, and petalite concentrate was added in the second step of transition roasting, and the others remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 45 was 55.65%.

[0091] Comparative Example 46: Taking Example 10 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, aluminum electrolysis overhaul slag and calcium fluoride were not added, and in the second step of transition roasting, aluminum electrolysis overhaul slag and calcium fluoride were added, and the other factors remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 46 was 52.91%.

[0092] Comparative Example 47: Taking Example 10 as a benchmark, in the first step of low-temperature sulfuric acid activation calcination, neither petalite concentrate nor calcium fluoride was added, and in the second step of transition roasting, petalite concentrate and calcium fluoride were added, and the others remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 47 was 55.33%.

[0093] Comparative Example 48: Taking Example 10 as a benchmark, the third step of low-temperature acidification and calcination is omitted, and water leaching is performed directly, while other conditions remain unchanged. After testing and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 48 is 40.31%.

[0094] Comparative Example 49: Taking Example 10 as a benchmark, the first step of low-temperature sulfuric acid activation calcination is not performed, and the same concentrated sulfuric acid immersion is used instead, and the temperature is not increased for activation calcination. Others remain unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 49 is 53.17%.

[0095] Comparative Example 50: Based on Example 10, the first step of low-temperature sulfuric acid activation calcination is not performed. Instead, the raw materials are pre-grinded in a ball mill for 60 minutes and then subjected to transition roasting. Others remain unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 50 is 53.38%.

[0096] Comparative Example 51: Based on Example 10, the first step of low-temperature sulfuric acid activation calcination was not performed. Instead, the raw materials were mixed with the same equivalent of sulfuric acid and subjected to ultrasonic treatment for 30 minutes. The other parameters remained unchanged. After detection and analysis, the leaching rate of Li in the leaching solution obtained in Comparative Example 51 was 54.19%.

[0097] From the above data, it can be seen that in order to process one or more lithium-containing raw materials such as petalite concentrate, spodumene concentrate, lithium porcelain stone and lithium-containing glass waste, it is necessary to first perform low-temperature activation calcination with sulfuric acid, then perform transformation roasting, and finally perform low-temperature acid calcination to significantly improve the lithium extraction effect of difficult-to-extract lithium minerals and lithium-containing raw materials. The principle may be: Too much aluminum electrolytic overhaul slag introduces too many other impurity ions, which will cause the ore and lithium-containing raw materials to transform into an amorphous dense phase and reduce the lithium extraction rate. Therefore, it is impossible to improve the attack effect by simply increasing the amount of electrolytic overhaul slag.

[0098] The low-temperature activation calcination of the composite raw material with sulfuric acid of the present invention is to activate the activity of F in the aluminum electrolysis overhaul slag, provide sufficient F source for the transformation roasting process to promote the transformation of the crystal structure of complex ores and lithium-containing raw materials, so that a small amount of fluorine can be continuously supplied during the transformation roasting process to continuously attack the complex ores and raw materials, and at the same time, the fluoride activation calcination realizes the pre-attack and attachment of the complex mineral and raw material structure. During the transformation roasting, the fluoride first quickly removes the surface of the complex ore and the hard-to-attack sites, and the activated aluminum electrolysis overhaul slag continuously supplies a small amount of fluorine to continuously attack the complex ores and raw materials that have been pre-attacked by the fluoride, and the structure of the complex ores and raw materials is disintegrated. This effect can utilize the effect of the pretreatment activation calcination and the synergistic effect of the transformation roasting process to disintegrate the structure of the complex ores and raw materials to a state that can be treated by subsequent acidification calcination; the final low-temperature acidification calcination of the transformation material is to further loosen the complex ores and raw materials that have been treated and disintegrated, and fully replace Li from the loose complex ore and raw material crystal structure. +; Treat it to a state where it can be directly immersed in water to achieve a significant improvement in the lithium extraction effect. Ultimately, it is possible to complete the efficient development and utilization of complex natural hard rock lithium ores and secondary lithium resources at around 900°C, reducing the traditional sulfuric acid method at 1100°C by 200°C, saving significant energy consumption.

[0099] In general, the technical solution of the present invention uses sulfuric acid low-temperature activation calcination to ensure that a continuous and sufficient fluorine source is provided with a relatively low amount of aluminum electrolysis overhaul slag, and at the same time utilizes the dual synergistic effect of low-temperature activation of fluoride. After the fluoride and overhaul slag are pre-treated with low-temperature activation of ore and lithium-containing raw materials, the fluoride pre-attack cleaning is combined with the overhaul slag slow release to continuously provide a small amount of fluorine attack, and a very small amount of fluorine is used to force the ore and lithium-containing raw materials to undergo crystal transformation at a relatively low temperature during the transformation stage, and the transformation efficiency is extremely high. Combined with subsequent low-temperature acidification and calcination, efficient extraction of lithium is achieved. Compared with the traditional high-temperature transformation sulfuric acid method, this solution has the advantages of lower energy consumption and a wider range of lithium resources. The entire process solution is low-cost and suitable for industrial application. It plays an extremely important role in the efficient development and utilization of lithium resources and has excellent application and promotion prospects.

[0100] At this point, those skilled in the art recognize that, although the embodiments of the present invention have been shown and described in detail herein, many other variations or modifications that conform to the principles of the present invention can still be directly determined or derived based on the contents disclosed in the present invention without departing from the spirit and scope of the present invention. Therefore, the scope of the present invention should be understood and recognized as covering all such other variations or modifications.

Claims

1. A method for extracting lithium by pre-treating complex lithium ore or lithium-containing waste materials through low-temperature transformation roasting, characterized in that: The lithium extraction method comprises the following steps: Step S1: uniformly mixing the lithium-containing composite material with aluminum electrolysis overhaul slag, fluoride and concentrated sulfuric acid, and then performing low-temperature calcination pretreatment to obtain a pretreated material; The lithium-containing composite material is one or more of petalite concentrate, spodumene concentrate, lithium porcelain stone and lithium-containing glass waste; Step S2: subjecting the pretreated material in step S1 to transformation roasting to obtain transformation material; Step S3: uniformly mixing the transformation material in step S2 with concentrated sulfuric acid, and performing low-temperature acidification and calcination to obtain an acidified material; Step S4: placing the acidified material in step S3 in water for water immersion, and then performing solid-liquid separation to obtain a lithium leaching solution, thereby achieving efficient lithium extraction.

2. A method for extracting lithium by pre-treating complex lithium ore or lithium-containing waste by low-temperature transformation roasting as described in claim 1, characterized in that: The lithium oxide content of the petalite concentrate is 1.0~4.5wt%; the lithium oxide content of the spodumene concentrate is 3.0~6.5wt%; the lithium oxide content of the lithium porcelain stone is 0.5~4.0wt%; the lithium oxide content of the lithium-containing glass waste is 1.0~6.0wt%; the lithium oxide content of the aluminum electrolysis overhaul slag is 0.5~3.5wt%.

3. A method for extracting lithium by pre-treating complex lithium ore or lithium-containing waste by low-temperature transformation roasting as claimed in claim 1, characterized in that: The F content of the aluminum electrolysis overhaul slag is 3.0-15.0wt%, and the fluoride includes calcium fluoride.

4. A method for extracting lithium by pre-treating complex lithium ore or lithium-containing waste by low-temperature transformation roasting as claimed in claim 1, characterized in that: The mass ratio of the lithium-containing composite material, the aluminum electrolysis overhaul slag and the fluoride in step S1 is 5-40:1-2:0.5-1; the molar amount of concentrated sulfuric acid added in step S1 is 1 to 4 times the molar amount of F in the aluminum electrolysis overhaul slag and the fluoride.

5. A method for extracting lithium by pre-treating complex lithium ore or lithium-containing waste by low-temperature transformation roasting as claimed in claim 1, characterized in that: The low-temperature calcination pretreatment temperature in step S1 is 200-320° C.; the low-temperature activation calcination time in step S1 is 30-60 min.

6. A method for extracting lithium by pre-treating complex lithium ore or lithium-containing waste by low-temperature transformation roasting as claimed in claim 1, characterized in that: The intermediate transformation roasting temperature in step S2 is 800-950° C.; the intermediate transformation roasting time in step S2 is 20-100 min.

7. A method for extracting lithium by pre-treating complex lithium ore or lithium-containing waste by low-temperature transformation roasting as claimed in claim 1, characterized in that: In step S3, the amount of sulfuric acid mixed with the conversion material is 1.2 to 1.4 times the theoretical acid consumption when all the lithium elements in the lithium-containing composite material are converted into lithium sulfate, and the sulfuric acid is concentrated sulfuric acid.

8. A method for extracting lithium by pre-treating complex lithium ore or lithium-containing waste by low-temperature transformation roasting as claimed in claim 1, characterized in that: The temperature of the low-temperature acidification calcination in step S3 is 150-300° C.; the time of the low-temperature acidification calcination in step S3 is 10-60 min.

9. A method for extracting lithium by pre-treating complex lithium ore or lithium-containing waste by low-temperature transformation roasting as claimed in claim 1, characterized in that: The water immersion reaction conditions in step S4 are: liquid-to-solid ratio of 1-5 mL:1 g, temperature of 25-90° C., and time of 5-60 min.

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