Method for treating lepidolite by using lithium-extraction waste residue for synergic roasting

Abstract

The present application provides a method for treating lepidolite by using lithium extraction waste residue for synergistic roasting, comprising: mixing lepidolite with sulfuric acid, and obtaining a first clinker after low-temperature roasting; mixing the first clinker with lithium extraction residue of the lithium sulfate process, and performing high-temperature two-stage roasting; and obtaining a leaching solution containing lithium, potassium, rubidium, beryllium and thallium and a water leaching residue by water leaching and solid-liquid separation of the roasted clinker. The present application uses the hazardous waste lithium extraction residue as a two-stage roasting additive, so that the beryllium and thallium contained in the silicate lattice are dissociated and converted into soluble sulfates in a high-temperature sulfuric acid environment, and then removed by water leaching. The leaching toxicity of the treated water leaching residue is lower than the national hazardous waste limit value, and the residue is converted into general solid waste. The present application fundamentally solves the problem of hazardous waste disposal and greatly reduces the cost and environmental risk. At the same time, the residual silicate skeleton of the lithium extraction residue preferentially captures the newly active aluminum compounds decomposed from lepidolite during roasting, inhibits the side reaction of alkali metal aluminosilicates, significantly improves the water solubility transformation rate of lithium, potassium and rubidium, and realizes efficient extraction of valuable metals.

Description

Technical Field

[0001] This invention belongs to the field of mineral lithium extraction technology, and relates to a method for treating lepidolite by co-roasting lithium extraction waste residue. Specifically, it designs a method for disposing of toxic metals such as beryllium and thallium in the lithium extraction residue of lepidolite sulfate method, and synergistically improves the alkali metal conversion rate during the two-stage sulfuric acid roasting process. Background Technology

[0002] Lithium mica is an important resource for lithium, rubidium, and cesium. Currently, the main methods for extracting alkali metals from lepidolite include the sulfate process, the sulfate process, and the pressure leaching process. Among these, the sulfate process (such as the potassium sulfate or sodium sulfate roasting-water leaching process) is widely used due to its mild reaction and low equipment corrosivity. However, this process generates a large amount of water leaching slag. This slag is not only abundant (20-30 tons of slag are generated for producing 1 ton of Li₂CO₃), but also contains toxic metals such as beryllium and thallium that are present in the raw materials. These slags are enriched in the slag during the roasting-water leaching process, leading to its classification as hazardous solid waste. The high cost of treatment and disposal severely restricts the economic and environmental benefits of the sulfate process.

[0003] The sulfuric acid process has attracted much attention due to its high reaction efficiency and wide applicability. However, traditional sulfuric acid leaching solutions contain large amounts of impurities such as aluminum and iron, leading to lengthy subsequent wet processing, high alkali consumption, and high lithium loss rates. Therefore, a two-stage roasting process can be adopted. The first stage involves low-temperature sulfation roasting to convert lithium into soluble sulfates; the second stage involves high-temperature roasting to further decompose iron, aluminum, and other impurity sulfates, solidifying them in the slag as oxides or silicates, thereby achieving selective sulfation of alkali metals in lepidolite. However, during the high-temperature roasting process, alkali metals such as lithium, potassium, and rubidium readily react with aluminum and silicon to form insoluble complex aluminosilicates, resulting in a significant decrease in the conversion rate of the target metal and causing resource waste.

[0004] Therefore, existing technologies lack effective methods for treating toxic metals in lithium extraction slag produced by the current sulfate process. There is an urgent need for a new method that can reduce the toxicity of lithium extraction slag while simultaneously controlling the reaction behavior of impurities such as iron and aluminum during the sulfuric acid roasting of lepidolite, and inhibiting secondary solidification of alkali metals, thereby overcoming existing bottlenecks. Summary of the Invention

[0005] The present invention aims to solve the problems of low conversion rate in the two-stage sulfuric acid roasting method due to the formation of insoluble complexes between alkali metals and aluminum and silicon at high temperature, and the difficulty in treating toxic metals such as beryllium and thallium in the lithium extraction slag of the sulfate method. The purpose is to provide a method for co-roasting lithium mica using lithium extraction waste slag to improve the lithium / potassium / rubidium conversion rate and co-treat toxic metals.

[0006] To achieve the above objectives, the present invention adopts the following technical solution: The applicant, through in-depth research, discovered that: (1) the lithium extraction slag from the lepidolite sulfate method has a complete aluminum-silicon framework structure, which exhibits a strong affinity and solidification ability for aluminum ions. During the high-temperature roasting process, when newly formed aluminum sulfate or aluminum oxide appears in the system, the water-leached slag can preferentially capture and stabilize it in its aluminum-silicon framework, thereby effectively inhibiting the combination of aluminum with alkali metals (lithium, potassium, rubidium), avoiding the loss of alkali metals in the form of composite aluminosilicates, and significantly improving the conversion rate of the target metals. (2) In the water leaching stage, the toxic metals such as beryllium and thallium that were originally enriched in the lithium extraction slag from the sulfate method are activated by phase recombination during the second-stage roasting process, and enter the water leaching liquid together with the alkali metal sulfates, thereby realizing the migration of toxic metals from the solid phase to the liquid phase, creating conditions for subsequent centralized treatment or recycling, and at the same time significantly reducing the toxicity of the final water-leached slag (mainly siliceous aluminum residue), achieving volume reduction and harmlessness. Based on the research results, a method for co-roasting treatment of lepidolite from lithium extraction waste slag is provided.

[0007] Specifically, a method for co-roasting lithium mica from lithium extraction waste residue is provided, including: Lithium mica is mixed with sulfuric acid and then calcined at low temperature to obtain a first-stage calcined clinker. The first-stage roasted clinker is mixed with lithium slag extracted by the lithium mica sulfate method and roasted at high temperature to obtain the second-stage roasted clinker. The two-stage roasted clinker is subjected to water leaching and solid-liquid separation to obtain a leachate containing lithium, potassium, rubidium, beryllium, and thallium, as well as a water-leached residue.

[0008] Compared with the prior art, one or more of the above technical solutions can achieve at least one of the following beneficial effects: 1) In the traditional sulfate process for lithium extraction, producing 1 ton of Li₂CO₃ generates 20-30 tons of lithium extraction slag. This slag is enriched with toxic metals such as beryllium and thallium, which are contained in a stable aluminosilicate phase. According to GB 5085.3, it is classified as hazardous waste, with disposal costs reaching hundreds to thousands of yuan per ton, posing a long-term environmental risk. This invention uses this hazardous waste slag as an auxiliary roasting agent in the two-stage sulfuric acid roasting process. The high-temperature sulfuric acid roasting environment causes most of the beryllium and thallium to dissociate from the aluminosilicate lattice, converting them into soluble sulfates. These sulfates are then efficiently removed by water leaching, transforming the hazardous waste slag generated by the traditional sulfate process into general solid waste, fundamentally solving its disposal problem. The leaching toxicity concentration of the treated slag is lower than the national standard limit for hazardous waste, and the slag is converted into general industrial solid waste, which can be subsequently used for building materials and other applications. The removed toxic elements are enriched in the water leaching solution, allowing for centralized disposal or recycling, fundamentally solving the hazardous waste slag disposal problem and significantly reducing disposal costs and environmental risks.

[0009] 2) This invention uses lithium extraction slag from hazardous waste as an auxiliary roasting agent to achieve efficient extraction of valuable metals from lepidolite. Specifically, the residual aluminosilicate framework structure in the lithium extraction slag can preferentially capture newly generated active aluminum sulfate / alumina produced by the decomposition of lepidolite under two-stage sulfuric acid roasting conditions, causing it to react and combine in situ on the surface of the slag framework. This effectively inhibits the side reaction of aluminum with alkali metal ions to form insoluble alkali metal aluminum silicates, significantly improving the water-soluble conversion rate of lithium, potassium, and rubidium, which can even reach over 90%. Attached Figure Description

[0010] To more clearly illustrate the technical solutions in the embodiments of the present invention or the prior art, the drawings used in the description of the embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are some embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.

[0011] Figure 1 This is a schematic flowchart of the disposal method for lithium extraction slag using the lepidolite sulfate process in this invention.

[0012] Figure 2 The images show the XRD patterns of the water-leached residues from Example 1 and Comparative Example 1. Detailed Implementation

[0013] Some implementations provide a method for extracting resources from lepidolite, including: Lithium mica is mixed with sulfuric acid and then roasted at low temperature to initially destroy the structure of the lethium mica, and lithium, potassium and rubidium are converted into soluble sulfates to obtain a first-stage roasted clinker. The first-stage roasted clinker is mixed with lithium extraction slag from the lepidolite sulfate method and roasted at high temperature to obtain a second-stage roasted clinker. The complete aluminum-silicon framework structure in the lithium extraction water leaching slag from the lepidolite sulfate method is utilized to preferentially solidify the newly formed aluminum sulfate or aluminum oxide in the roasting system, inhibiting the formation of insoluble complexes between alkali metals and aluminum-silicon. At the same time, the toxic metals such as beryllium and thallium enriched in the water leaching slag are activated at high temperature, creating conditions for subsequent water leaching into the solution. The two-stage roasted clinker is subjected to water leaching and solid-liquid separation to obtain a leachate containing lithium, potassium, rubidium, beryllium, and thallium, as well as a water-leached residue.

[0014] In this invention, the intact aluminum-silicon framework in the lithium extraction slag from the sulfate process is utilized to strongly solidify aluminum. During the second-stage high-temperature roasting, newly formed aluminum sulfate or alumina is preferentially captured, effectively blocking the side reaction that causes aluminum to combine with alkali metals (lithium, potassium, rubidium) to form insoluble complex aluminosilicates. Meanwhile, the beryllium and thallium toxic metal phases in the sulfate process lithium extraction slag are activated and enter the solution along with the alkali metal sulfates during the water leaching stage, realizing the migration of toxic metals from the solid phase to the liquid phase. The final water-leached slag is mainly aluminosilicate residue with a significantly reduced toxicity leaching concentration, transforming it from hazardous waste into general solid waste. This significantly reduces the amount of hazardous solid waste that can be disposed of, achieving a dual improvement in waste-to-waste treatment and environmental benefits.

[0015] In some preferred embodiments, the sulfuric acid is concentrated sulfuric acid with a concentration of 70–99 wt%; the mass ratio of the lepidolite to the concentrated sulfuric acid and water is 1:0.8–1.5:0.1–0.5, for example 1:0.8–1.5:0.1–0.3, 1:0.8–1.5:0.2–0.4, 1:0.8–1.5:0.3–0.5, 1:0.8–1.2:0.1–0.5, 1:0.8–1.2:0.1–0.3, 1:0. 8~1.2:0.3~0.5, 1:0.8~1.2:0.2~0.4, 1:1~1.5:0.1~0.5, 1:1~1.5:0.1~0.3, 1:1~1.5:0.3~0.5, 1:1~1.5:0.2~0.4, 1:1.2~1.5:0.1~0.5, 1:1.2~1.5:0.1~0.3, 1:1.2~1.5:0.2~0.4, 1:1.2~1.5:0.3~0.5, etc.

[0016] In some preferred embodiments, the lithium extraction slag from lepidolite using the sulfate method is the water-leached slag obtained by treating lepidolite using the sulfate method; the sulfate treatment of lepidolite refers to the process of mixing and roasting lepidolite with sulfate and then subjecting it to water leaching, resulting in a mineral with a complete aluminum-silicon framework structure and containing enriched toxic metals such as beryllium and thallium; the sulfate is one or more of potassium sulfate, sodium sulfate, and calcium sulfate, but sulfuric acid cannot be used as a substitute; the treatment temperature is a conventional treatment temperature in the art, usually between 750 and 1000°C.

[0017] In some preferred embodiments, the temperature of the low-temperature calcination is 100-300°C, preferably 150-220°C; and the time of the low-temperature calcination is 0.5-3 hours, preferably 1-2 hours.

[0018] In some preferred embodiments, the mass ratio of the first-stage roasted clinker to the lithium extraction slag from the lepidolite sulfate method is 1:0.1 to 0.5.

[0019] In some preferred embodiments, the high-temperature roasting temperature is 550-850℃, such as 550℃, 600℃, 650℃, 700℃, 750℃, 800℃, 850℃, etc.; the high-temperature roasting time is 0.5-3h, such as 0.5h, 1h, 2h, 3h, etc.

[0020] In some preferred embodiments, the water immersion temperature is 20–90°C, such as 20°C, 30°C, 40°C, 50°C, 60°C, 70°C, 80°C, 90°C, etc.; the water immersion time is 0.5–3 hours, such as 0.5 hours, 1 hour, 1.5 hours, 2 hours, 2.5 hours, 3 hours, etc.; the liquid-to-solid ratio of the water immersion is 2–10:1, such as 2:1, 3:1, 4:1, 5:1, 6:1, 7:1, 8:1, 9:1, 10:1, etc.

[0021] In some preferred embodiments, the water-soaked residue, after being washed and dried, is partially or entirely returned to step two as an auxiliary material for recycling.

[0022] In some preferred embodiments, the method further includes adding an oxidant to the leachate to reduce Tl + All oxidized to Tl 3 + Then, sulfides are added to precipitate beryllium thallium.

[0023] In some preferred embodiments, the amount of oxidant added is Tl. + Oxidized to Tl 3+ The dosage is 1.1 to 1.5 times that of the required oxidant, for example, 1.1 to 1.3 times, 1.2 to 1.5 times, 1.2 to 1.4 times, etc.; the oxidant is one or more of hydrogen peroxide, persulfate, potassium permanganate, manganese dioxide, and ozone, and is more preferably hydrogen peroxide.

[0024] In some preferred embodiments, the amount of the sulfide is 1.1 to 1.5 times the amount of the sulfide precipitant required for beryllium and thallium ions, for example, 1.1 to 1.3 times, 1.1 to 1.2 times, 1.2 to 1.5 times, 1.2 to 1.4 times, etc.; the sulfide is sodium sulfide, potassium sulfide, ammonium sulfide, etc.

[0025] In some preferred embodiments, the pH of the precipitate obtained after solid-liquid separation of precipitated beryllium and thallium is adjusted to 10-12. After the causticization reaction, a deeply purified aqueous solution is obtained through solid-liquid separation. For example, calcium oxide slurry, calcium hydroxide slurry, ammonia, sodium hydroxide solution, potassium hydroxide solution, etc., can be used for causticization treatment.

[0026] In some preferred embodiments, the method further includes separating lithium ions, potassium ions, and rubidium ions from the deeply purified aqueous solution (sulfate aqueous solution), specifically, this can be achieved through extraction.

[0027] To facilitate understanding of the present invention, the present invention will be described more fully and in detail below with reference to the accompanying drawings and preferred embodiments, but the scope of protection of the present invention is not limited to the following specific embodiments.

[0028] Unless otherwise defined, all technical terms used herein have the same meaning as commonly understood by those skilled in the art. The technical terms used herein are for the purpose of describing particular embodiments only and are not intended to limit the scope of the invention.

[0029] Unless otherwise specified, all raw materials, reagents, instruments and equipment used in this invention can be purchased from the market or prepared by existing methods.

[0030] Example 1 The process flow diagram for disposing of lithium residue from the lithium mica sulfate extraction method is shown below. Figure 1 As shown, it includes: Step 1: Take 10g of lepidolite concentrate (Li content 1.1%, Fe content 8.4%, Al content 11.6%, K content 7.9%, Rb content 0.8%, Be content 0.003%, Tl content 0.003%) and mix it evenly with concentrated sulfuric acid (98%) and water at a mass ratio of 1:1.1:0.2. Place it in a tube furnace and roast at 180℃ for 1 hour to obtain roasted clinker. Step 2: Take 20g of the first-stage roasted clinker obtained in Step 1 and mix it evenly with 6g of lithium extraction slag from the lepidolite sulfate method (i.e., the leaching residue obtained after roasting and water leaching of lepidolite with sulfates such as potassium sulfate, sodium sulfate, or calcium sulfate). Roast at 750℃ for 1 hour to obtain the second-stage roasted clinker.

[0031] Step 3: The two-stage roasted clinker obtained in Step 2 is added to pure water at a liquid-to-solid ratio of 3:1 for water immersion treatment. The immersion temperature is 60℃, and the immersion time is 2 hours. After filtration, the leachate is obtained (Li concentration 0.80 g / L, K concentration 9.61 g / L, Rb concentration 0.40 g / L, Be concentration 0.006 g / L, Tl...). + Concentration 0.002 g / L) and water-leached residue. The XRD pattern of the water-leached residue is shown below. Figure 2 As shown.

[0032] According to testing and calculation, the leaching rates of lithium, potassium, and rubidium were 92.5%, 93.2%, and 94.3%, respectively; the leaching rates of beryllium and thallium were 85.6% and 95.3%, respectively.

[0033] Leaching rates of major elements in lepidolite (α)i The calculation method for (%) is as follows: In the formula, i represents Li, K, Rb, Fe, Al, Be, and Tl; C i The concentration of the above ions in the leachate is determined by ICP, in g / L. V i The volume of the leachate is in liters (L). ω i The values ​​represent the mass fractions of the above elements in the roasted clinker, expressed as % (%). m represents the mass of the roasted clinker, in grams.

[0034] Example 2 The leachate obtained in Example 1 was subjected to Be and Tl precipitation treatment, and hydrogen peroxide with an excess coefficient of 1.2 was added as an oxidant to precipitate Tl. + Oxidation to Tl 3+ Then, potassium sulfide with an excess coefficient of 1.4 was added as a precipitant (i.e., the amount of potassium sulfide used was 1.4 times the amount of potassium sulfide required for beryllium and thallium ions). Simultaneously, the pH of the solution was adjusted to 7.0 using potassium hydroxide. After reacting for 1 hour, liquid-solid separation was performed to obtain a purified beryllium-thallium solution. Testing showed that the precipitation rates of beryllium and thallium were 98.1% and 97.7%, respectively, while the precipitation rates of lithium, potassium, and rubidium were all below 0.5%.

[0035] Example 3 The process flow diagram for disposing of lithium residue from the lithium mica sulfate extraction method is shown below. Figure 1 As shown, it includes: Step 1: Take 10g of lepidolite concentrate (Li content 1.1%, Fe content 8.4%, Al content 11.6%, K content 7.9%, Rb content 0.8%) and mix it evenly with concentrated sulfuric acid (98%) and water at a mass ratio of 1:1.1:0.2. Place it in a tube furnace and roast at 180℃ for 1 hour to obtain roasted clinker. Step 2: Take 20 g of the first-stage roasted clinker obtained in Step 1 and mix it evenly with 4 g of lithium extraction slag from the lepidolite sulfate method (i.e., the leaching residue obtained after roasting and water leaching of lepidolite with sulfates such as potassium sulfate, sodium sulfate, and calcium sulfate). Roast at 750℃ for 1 hour to obtain the second-stage roasted clinker.

[0036] Step 3: The two-stage roasted clinker obtained in Step 2 is added to pure water at a liquid-to-solid ratio of 3:1 for water immersion treatment. The immersion temperature is 90℃, and the immersion time is 2 hours. After filtration, the leachate is obtained (Li concentration 0.81 g / L, K concentration 9.78 g / L, Rb concentration 0.41 g / L, Be concentration 0.007 g / L, Tl...). +Concentration 0.002 g / L) and water-leached residue. The XRD pattern of the water-leached residue is shown below. Figure 2 As shown in the figure. Testing revealed that the leaching rates for lithium, potassium, and rubidium were 93.8%, 95.5%, and 95.6%, respectively; while the leaching rates for beryllium and thallium were 90.1% and 94.7%, respectively.

[0037] Example 4 The leachate obtained in Example 3 was subjected to Be and Tl precipitation treatment, and hydrogen peroxide with an excess coefficient of 1.1 was added as an oxidant to precipitate Tl. + Oxidation to Tl 3+ Then, potassium sulfide with an excess coefficient of 1.4 was added as a precipitant, and the pH of the solution was adjusted to 7.0 using potassium hydroxide. After reacting for 1 hour, liquid-solid separation was performed to obtain a beryllium-thallium purified solution. Testing showed that the precipitation rates of beryllium and thallium were 98.5% and 97.1%, respectively, while the precipitation rates of lithium, potassium, and rubidium were all below 0.5%.

[0038] Example 5 The process flow diagram for disposing of lithium residue from the lithium mica sulfate extraction method is shown below. Figure 1 As shown, it includes: Step 1: Take 10g of lepidolite concentrate (Li content 1.1%, Fe content 8.4%, Al content 11.6%, K content 7.9%, Rb content 0.8%) and mix it evenly with concentrated sulfuric acid (98%) and water at a mass ratio of 1:1.2:0.2. Place it in a tube furnace and roast at 180℃ for 1 hour to obtain roasted clinker. Step 2: Take 20 g of the first-stage roasted clinker obtained in Step 1 and mix it evenly with 6 g of lithium extraction slag from the lepidolite sulfate method (i.e., the leaching residue obtained after roasting and water leaching of lepidolite with sulfates such as potassium sulfate, sodium sulfate, and calcium sulfate). Roast at 800℃ for 1 hour to obtain the second-stage roasted clinker.

[0039] Step 3: The two-stage roasted clinker obtained in Step 2 is added to pure water at a liquid-to-solid ratio of 3:1 for water immersion treatment. The immersion temperature is 60℃, and the immersion time is 2 hours. After filtration, the leachate is obtained (Li concentration 0.79 g / L, K concentration 9.22 g / L, Rb concentration 0.35 g / L, Be concentration 0.005 g / L, Tl...). + Concentration 0.002 g / L) and water-leached residue. The XRD pattern of the water-leached residue is shown below. Figure 2 As shown in the figure. Testing revealed that the leaching rates for lithium, potassium, and rubidium were 90.2%, 89.3%, and 88.6%, respectively; while the leaching rates for beryllium and thallium were 82.2% and 91.1%, respectively.

[0040] Example 6 The leachate obtained in Example 5 was subjected to Be and Tl precipitation treatment, and hydrogen peroxide with an excess coefficient of 1.2 was added as an oxidant to precipitate Tl. + Oxidation to Tl3+ Then, potassium sulfide with an excess coefficient of 1.4 was added as a precipitant, and the pH of the solution was adjusted to 7.0 using potassium hydroxide. After reacting for 1 hour, liquid-solid separation was performed to obtain a beryllium-thallium purified solution. Testing showed that the precipitation rates of beryllium and thallium were 98.4% and 98.0%, respectively, while the precipitation rates of lithium, potassium, and rubidium were all below 0.5%.

[0041] Comparative Example 1 (Traditional Two-Stage Sulfuric Acid Roasting Method) Take 10g of the same lepidolite concentrate as in Example 1, and mix it evenly with concentrated sulfuric acid (98%) and water at a mass ratio of 1:1.1:0.2. Place it in a tube furnace and roast at 180°C for 1 hour to obtain a first-stage roasted clinker.

[0042] A section of roasted clinker was directly roasted at 750°C for 1 hour (without adding water leaching residue), and then water leached under the same conditions as in Example 1. The leaching rates of beryllium and thallium were tested to be 86.3% and 94.4%, respectively, while the leaching rates of lithium, potassium, and rubidium were 81.9%, 70.8%, and 58.4%, respectively.

[0043] XRD pattern of water-leached residue is shown below Figure 2 As shown. By Figure 2 It can be seen that the water leaching residue of the traditional two-stage sulfuric acid roasting method contains a large amount of aluminum-silicon oxide, i.e., composite silicates, while the residue after assisted roasting of the sulfuric acid brine leaching residue shows a significant increase in silica phase and a smaller proportion of composite silicates. This indicates that the aluminum-silicon structure in the sulfuric acid brine leaching residue can significantly inhibit the secondary solidification loss of alkali metals caused by the formation of silicates. Furthermore, considering that the leaching rates of lithium, potassium, and rubidium in Example 1 are significantly higher than those in Comparative Example 1, analysis shows that this is because the complete aluminum-silicon framework structure in the lithium extraction residue has a strong affinity for active aluminum compounds, preferentially capturing newly formed aluminum sulfate or aluminum oxide produced by the decomposition of lepidolite at high temperatures, thereby inhibiting the side reaction of aluminum combining with alkali metals to form insoluble aluminum silicates, thus increasing the leaching rates of alkali metals such as lithium, potassium, and rubidium.

[0044] Comparative Example 2 10g of sulfate roasted water-leached residue was roasted at 750℃ for 1 hour, and then water-leached under the same conditions as in Example 1. The leaching rates of beryllium and thallium were 4.4% and 8.2%, respectively, while the leaching rates of lithium, potassium, and rubidium were 3.2%, 2.2%, and 1.1%, respectively.

[0045] Comparing Example 1 and Comparative Example 2, it can be seen that the first-stage roasted clinker obtained from the treatment of lepidolite concentrate can significantly promote the conversion of beryllium and thallium into soluble sulfates in the second-stage roasting process of sulfate roasted water leaching residue, thus significantly improving the leaching rate of beryllium and thallium.

[0046] The comparison between the above embodiments and comparative examples shows that the present invention, by adding sulfate to lithium slag in the two-stage high-temperature roasting stage, not only significantly improves the conversion rate of lithium, potassium and rubidium, but also achieves efficient migration and removal of toxic metals such as beryllium and thallium in the water-leached slag, thus greatly reducing the toxicity of the final water-leached slag.

[0047] Toxicity leaching experiments were conducted on the lithium extraction slag obtained by the lepidolite sulfate method before treatment and the water leaching slag obtained after treatment of the lepidolite sulfate roasting lithium extraction slag in Example 1. The results are shown in Tables 1 and 2. Comparing Tables 1 and 2, it can be seen that the leaching toxicity of the lithium extraction slag obtained after treatment by lepidolite sulfate roasting lithium extraction is significantly reduced. Further analysis, combined with the comparison between Example 1 and Comparative Examples 1 and 2, shows that this is because the lithium extraction slag obtained by lepidolite sulfate roasting lithium extraction serves as an auxiliary roasting agent in the two-stage sulfuric acid roasting process. The first-stage roasting clinker allows toxic metals such as beryllium and thallium to dissociate from the stable aluminosilicate phase in lepidolite sulfate at high temperatures, transforming them into soluble sulfates, which are then efficiently removed from the slag through water leaching.

[0048] Table 1. Toxicity leaching results of lithium extraction slag after lithium mica sulfate roasting. Table 2. Toxicity leaching results of lithium extraction slag from lepidolite sulfate process before treatment. 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 treating lepidolite using lithium extraction waste residue through co-roasting, characterized in that, include: Lithium mica is mixed with sulfuric acid and then calcined at low temperature to obtain a first-stage calcined clinker. The first-stage roasted clinker is mixed with lithium slag extracted by the lithium mica sulfate method and roasted at high temperature to obtain the second-stage roasted clinker. The two-stage roasted clinker is subjected to water leaching and solid-liquid separation to obtain a leachate containing lithium, potassium, rubidium, beryllium, and thallium, as well as a water-leached residue.

2. The method for co-roasting lithium mica using lithium extraction waste residue as described in claim 1, characterized in that, The sulfuric acid is concentrated sulfuric acid; the mass ratio of the lepidolite to the concentrated sulfuric acid and water is 1:0.8-1.5:0.1-0.

5. The lithium extraction slag obtained by the lepidolite sulfate method is the water-leached slag obtained by treating lepidolite using the sulfate method; the lepidolite sulfate method refers to the process of mixing and roasting lepidolite with sulfate and then subjecting it to water leaching, resulting in a mineral with a complete aluminum-silicon framework structure and containing enriched beryllium and thallium toxic metals; the sulfate is one or more of potassium sulfate, sodium sulfate, and calcium sulfate.

3. The method for co-roasting lithium mica using lithium extraction waste residue as described in claim 1, characterized in that, The low-temperature roasting temperature is 100–300°C; the low-temperature roasting time is 0.5–3 hours.

4. The method for treating lepidolite using lithium extraction waste residue through co-roasting as described in claim 1, characterized in that, The mass ratio of the roasted clinker to the lithium extraction slag from the lepidolite sulfate method is 1:0.1 to 0.

5.

5. The method for co-roasting lithium mica using lithium extraction waste residue as described in claim 1, characterized in that, The high-temperature roasting temperature is 550–850°C; the high-temperature roasting time is 0.5–3 hours.

6. The method for co-roasting lithium mica using lithium extraction waste residue as described in claim 1, characterized in that, The water immersion temperature is 20–90°C; the water immersion time is 0.5–3 hours; and the liquid-to-solid ratio of the water immersion is 2–10:

1.

7. The method for co-roasting lithium mica using lithium extraction waste residue as described in claim 1, characterized in that, After being washed and dried, the water-soaked residue is partially or entirely returned to step two as an auxiliary material for recycling.

8. The method for co-roasting lithium mica using lithium extraction waste residue as described in claim 1, characterized in that, It also includes adding an oxidant to the leachate, followed by the addition of sulfides to precipitate beryllium thallium.

9. The method for co-roasting lithium mica using lithium extraction waste residue as described in claim 8, characterized in that, The amount of oxidant added is Tl + Oxidized to Tl 3+ The dosage is 1.1 to 1.5 times the required amount of oxidant; the oxide is one or more of hydrogen peroxide, persulfate, potassium permanganate, manganese dioxide, and ozone; The amount of sulfide used is 1.1 to 1.5 times the amount of sulfide precipitant required for beryllium and thallium ions; the sulfide is one or more of sodium sulfide, potassium sulfide, and ammonium sulfide; The precipitation of beryllium thallium was carried out at a pH of 6–8.

10. The method for treating lepidolite using lithium extraction waste residue through co-roasting as described in claim 8, characterized in that, It also includes adjusting the pH of the precipitate obtained after solid-liquid separation of precipitated beryllium and thallium to 10-12, and after causticization reaction, obtaining a deeply purified aqueous solution through solid-liquid separation; It also includes the separation of lithium ions, potassium ions, and rubidium ions from deeply purified aqueous solutions.

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