Method for extracting lithium and thallium from lepidolite concentrate

Abstract

The invention relates to the technical field of concentrate lithium extraction, in particular to a method for extracting lithium and thallium from lepidolite concentrate. According to the technical scheme, the method comprises the following steps: adding a composite acid mixed complexing system and a surfactant into lepidolite concentrate, mixing and grinding, carrying out ultrasonic treatment, heating and extracting, and filtering to obtain a lithium-containing leaching solution; mixing a complexing agent, a diluent and an extraction agent of thallium to construct a first polar extraction system, and adding the lithium-containing leachate into the first polar extraction system for mixing and extraction to obtain a thallium-rich organic phase and raffinate; desorbing the thallium-rich organic phase to obtain a thallium-rich solution and a thallium-poor organic phase; and mixing a lithium complexing agent, a diluent and an extraction agent to construct a second polar extraction system, then adding the raffinate into the second polar extraction system, mixing and extracting to obtain a lithium complex, purifying, and eluting to obtain a Li < + >-rich solution. The method is mainly used for providing a new choice for the method for extracting the lithium and the thallium from the lepidolite concentrate.

Description

Technical Field

[0001] This invention relates to the field of lithium extraction technology from concentrates, and specifically to a method for extracting lithium and thallium from lepidolite concentrate. Background Technology

[0002] Traditional lithium extraction processes from lepidolite mainly rely on high-temperature roasting (above 800℃) or strong acid leaching.

[0003] Among them, the high-temperature roasting method requires a high temperature of 850-1000℃, which has significant drawbacks: (1) huge energy consumption and serious equipment wear; (2) the associated thallium element will be discharged with the flue gas. In order to capture the thallium in the flue gas, a huge spray purification system needs to be configured. However, the use of spray purification method produces high thallium wastewater with complex composition and high treatment difficulty, resulting in huge wastewater treatment costs; (3) some of the unvolatile thallium will enter the subsequent lithium smelting slag (lithium slag), causing the lithium slag to become thallium-rich hazardous waste. For every ton of lithium carbonate produced, 30 tons of lithium slag can be generated. Due to the potential environmental toxicity of thallium, it will lead to thallium pollution problems. Therefore, it is currently difficult to safely dispose of and utilize such waste residue. Most of them can only be stored in rented warehouses for storage management, which has created serious solid waste disposal pressure and environmental hazards.

[0004] Traditional acid leaching processes typically use HF or concentrated sulfuric acid. While this method avoids high temperatures, it also has significant drawbacks: (1) it is highly corrosive and consumes a lot of acid; (2) in acidic environments, associated metals such as thallium are easily dissolved along with lithium, resulting in poor product separation selectivity and high wastewater treatment costs. In recent years, some studies have attempted to extract lithium from lepidolite using mechanical activation + acid leaching. Although mechanical activation + acid leaching can achieve a lithium recovery rate of over 90%, it still requires high temperatures or long reaction times, making the process complex and energy consumption high. In addition, existing processes are mostly one-time extractions, making it difficult to achieve synergistic and efficient recovery of lithium and thallium, and the water recycling rate is low, resulting in a large environmental burden.

[0005] Therefore, developing a new technology that can efficiently activate lepidolite under mild conditions, selectively extract lithium, and achieve targeted separation and recovery of associated thallium, thereby reducing pollution at the source and lowering energy consumption and solid waste environmental risks, has become a key issue that urgently needs to be addressed in this field. Summary of the Invention

[0006] To address the problems existing in the prior art, the present invention aims to provide a method for extracting lithium and thallium from lepidolite concentrate, thereby solving the problem that existing lepidolite lithium extraction processes cannot simultaneously achieve efficient activation of lepidolite, selective extraction of lithium, and directional separation and recovery of associated thallium under mild conditions.

[0007] The technical solution of the present invention is as follows: This invention provides a method for extracting lithium and thallium from lepidolite concentrate, comprising the following steps: A composite acid complexing system and a surfactant are added to lepidolite concentrate, and the mixture is ground to form a lepidolite concentrate solids slurry. The lepidolite concentrate solids slurry is then subjected to ultrasonic treatment to obtain an ultrasonically broken lepidolite concentrate solids slurry. The solid content slurry of lepidolite concentrate after ultrasonic cracking is heated and leached, then filtered to obtain lithium-containing leachate and lepidolite leaching residue. A thallium complexing agent, a diluent, and an extractant are mixed to construct a first polar extraction system. The lithium-containing leachate is then added to the first polar extraction system for mixing and extraction to obtain a thallium-rich organic phase and a raffinate. The thallium-rich organic phase was desorbed to obtain a thallium-rich solution and a thallium-poor organic phase; A lithium complexing agent, a diluent, and an extractant are mixed to construct a second polar extraction system. The raffinate is then added to the second polar extraction system for mixing and extraction to obtain a lithium complex. The lithium complex is then purified and eluted to obtain a Li-rich compound. + Solution.

[0008] First, by grinding the lepidolite concentrate, the specific surface area of ​​the particles is significantly increased, and the interlayer spacing is broken by mechanical compression, allowing metal ions (Li) to escape. + 、Tl +It is easier to contact with external reagents. Then, by combining the chemical chelation activation of the complex acid mixed complex system, the dispersion and anti-agglomeration effect of the surfactant, and the cavitation impact effect of ultrasound, deep activation of lepidolite under mild conditions below 100℃ is achieved. Then, by leaching and filtering the lepidolite concentrate slurry after ultrasonic cracking, a lithium-containing leachate is obtained. Then, extraction is carried out in two steps. First, the complexing agent, diluent and extractant of thallium are mixed to construct a first polar extraction system. Then, the lithium-containing leachate is added to the first polar extraction system for mixing and extraction to obtain a thallium-rich organic phase and raffinate. Thallium is extracted separately to form a thallium-rich organic phase. Then, the thallium-rich organic phase is desorbed to obtain a thallium-rich solution. Secondly, a lithium complexing agent, diluent, and extractant are mixed to construct a second polar extraction system. The raffinate is then added to this system for mixing and specific lithium extraction, forming a lithium complex. Finally, the lithium complex is purified and eluted to obtain a lithium-rich solution. This entire process, through a clever design of stepwise activation and extraction, achieves effective separation of lithium and thallium. It recovers both valuable elements while avoiding interference between the two metal ions during extraction, thus improving the purity of the final product. Ultimately, this process achieves highly efficient activation of lepidolite, enabling efficient extraction and targeted recovery of lithium and thallium. It effectively solves the technical problem of existing lepidolite lithium extraction processes being unable to simultaneously achieve highly efficient activation of lepidolite, selective lithium extraction, and targeted separation and recovery of associated thallium under mild conditions.

[0009] Furthermore, the degree of the first grinding is to grind the lepidolite concentrate to D90 < 45µm.

[0010] Further purification of the lithium complex includes the following steps: transferring the lithium complex into an ion exchange system and purifying it with resin to obtain a lithium-adsorbed resin; eluting the lithium-adsorbed resin to obtain a Li-rich resin. + Solution.

[0011] Furthermore, the resin filler content, calculated by bed volume, is 5-10 L of resin per cubic meter of lithium complex; the flow rate of the lithium complex in the ion exchange system is 15-50 L / h, and the circulation time is 6-8 hours / cycle; the eluent is 0.1-0.5 mol / L dilute hydrochloric acid, and the elution flow rate is 0.5-0.8 bed volume / hour; the Li-rich... + The pH of the solution is 1.0 to 1.5.

[0012] Furthermore, the resin is D201 strong acid sulfonic acid resin, and the resin after lithium adsorption is D201 strong acid sulfonic acid resin after lithium adsorption.

[0013] Furthermore, the mass ratio of the lithium mica concentrate, the complex acid mixed complex system, and the surfactant is 1000:1000~1500:5.

[0014] Furthermore, the complex acid mixed complex system includes an acid and nitric acid (NTA), wherein the acid is selected from one or both of citric acid and succinic acid, the concentration of the acid in the complex acid mixed complex system is 0.15-0.25 mol / L, the concentration of nitric acid (NTA) is 0.05-0.08 mol / L, and the molar ratio of acid to nitric acid is 3:1.

[0015] One or both of citric acid and succinic acid, under weakly acidic conditions (pH≈3~4), form metal-citric acid / succinic acid complexes using their carboxyl groups, which can efficiently chelate Li between the layers of lepidolite. + 、Tl + The isocations achieve and partially dissolve, thereby weakening the charge balance of layered aluminosilicates. NTA is a tricarboxylic acid amine ligand capable of forming stable five-membered ring chelates with almost all metal ions (stability constants around 10). 6 ~10 8 (Scope). The strong complexing effect of NTA extracts more interlayer metal ions, significantly reducing the electrostatic attraction of interlayer bonds. In addition, the synergistic complexing effect formed by the combination of one or both citric acid and succinic acid with nitrilotriacetic acid results in higher activation efficiency compared to single organic acid systems. This achieves deep activation of lepidolite under mild conditions, eliminating the need for harsh processes such as high temperature and high pressure, significantly reducing energy consumption and equipment costs in the activation process, while simultaneously achieving Li… + with Tl + The directional complexation lays a key foundation for the selective extraction of lithium and the directional separation and recovery of thallium in subsequent processes.

[0016] Furthermore, the mass ratio of the lithium mica concentrate to the composite acid complex system is 1:1.0 to 1.5.

[0017] Furthermore, the surfactant is one or both of sodium dodecyl sulfate and sodium dodecyl sulfonate.

[0018] Sodium dodecyl sulfate / sodium dodecyl sulfonate is an anionic surfactant that adsorbs onto the surface of mineral particles in the slurry, forming a negatively charged layer. This significantly reduces the solid-liquid interfacial tension and improves the wettability and dispersibility of the slurry. Simultaneously, the mutual repulsion of the negative charges prevents fine particles from re-aggregating during subsequent high-energy impacts, ensuring that the shock waves generated by cavitation act uniformly on each layered wafer.

[0019] Furthermore, based on lepidolite concentrate, the amount of the surfactant used is 0.5 wt%.

[0020] Furthermore, the complexing agent for thallium is dibenzo-18-crown-6-ether, and the concentration of the complexing agent for thallium in the first polar extraction system is 0.8~1.2 g / L.

[0021] Furthermore, the lithium complexing agent is ethyl acetoacetate, with a concentration of 1.5~2.0 g / L.

[0022] Furthermore, the extractant is a 15-20 vol% solution of tributyl phosphate (TBP).

[0023] Furthermore, the solvent in the tributyl phosphate (TBP) solution is aviation kerosene.

[0024] Using dibenzo-18-crown-6-ether (DB-18-crown-6) to target Tl + Its high selective complexing ability, combined with the synergistic extraction effect of the extractant tributyl phosphate (TBP) solvent, enables the specific extraction of thallium from lithium-containing leachate, effectively avoiding the simultaneous extraction of lithium ions. This significantly improves the purity of raw materials in subsequent lithium extraction processes and enables the targeted enrichment and recovery of associated thallium, providing convenient conditions for the resource utilization or harmless disposal of thallium.

[0025] Under slightly higher pH conditions of 5.5–6.0, lithium ions are selectively complexed using ethyl acetoacetate (EAA) at a concentration of 1.5–2.0 g / L, combined with synergistic extraction using tributyl phosphate at a volume percentage concentration of 17–20%, and further purified by ion exchange resin. This method enables the targeted enrichment and separation of lithium ions in the raffinate, effectively avoiding co-extraction of other metal impurities, significantly improving the extraction selectivity and recovery rate of lithium, and further enhancing the purity of the lithium product through subsequent ion exchange resin purification.

[0026] Furthermore, the diluent is sulfonated kerosene.

[0027] Preferably, the sulfonated kerosene is industrial pure sulfonated kerosene.

[0028] Furthermore, the pH value of the first polar extraction system is 4.0~4.5, and the saponification rate is 40~65%; the total ratio of aqueous phase to organic phase in the first polar extraction system is 3~8:1, that is, the volume ratio of aqueous phase to organic phase in the first polar extraction system is 3~8:1.

[0029] Furthermore, the desorbent is an aqueous solution of H2SO4, the concentration of H2SO4 in the aqueous solution is 1.0~1.5 mol / L, and the ratio of thallium-rich organic phase to desorbent is 3~5:1.

[0030] Furthermore, the pH value of the second polar extraction system is 5.5~6.0, and the saponification rate is 40~65%; the total ratio of aqueous phase to organic phase in the second polar extraction system is 3~8:1, that is, the volume ratio of aqueous phase to organic phase in the second polar extraction system is 3~8:1.

[0031] Furthermore, the grinding is carried out in a sand mill, the main shaft speed of the sand mill is 150~200 r / min, the stirring axis speed of the sand mill is 0.5~1.5 m / s, the material-to-ball ratio of the ball mill is 10:1, the grinding concentration of the ball mill is 65~70%, and the grinding time is 30~60 min.

[0032] Furthermore, the conditions for ultrasonic treatment include: ultrasonic frequency of 20~25kHz, power density of 320~350W / L, treatment temperature controlled between 80~95℃, treatment time of 45~60min, continuous flow treatment, slurry flow rate of 8~12L / min, and cavitation number controlled between 0.8~1.2.

[0033] Under ultrasonic treatment conditions of 20–25 kHz frequency, 320–350 W / L power density, and 85–95 °C, a microenvironment with instantaneous pressures exceeding 5000 K and 100 MPa can be generated. The collapse process is accompanied by high-speed microjet and shock waves, which can generate strong shear and impact stresses at the nanoscale. These impact stresses directly tear the Si-O-Al bonds of layered aluminosilicates, forming microcracks and interlayer delamination, which can significantly reduce overall energy consumption and avoid large-scale high-temperature calcination.

[0034] Furthermore, the heating extraction adopts a three-stage countercurrent water leaching process, the liquid-to-solid ratio of the leaching process is 4~5:1, the heating extraction temperature is 90~95℃, and the heating extraction time is 180~240min; the heating extraction is also accompanied by stirring, and the stirring speed is 300~400rpm; the filtration is vacuum filtration.

[0035] Furthermore, the process of adding the lithium-containing leachate to the first polar extraction system for mixing and extraction includes: using a mechanical stirrer to stir and mix at a speed of 250~300 rpm for 2~3 min; the extraction is a 3~4 stage countercurrent extraction.

[0036] Furthermore, the addition of the raffinate to the second polar extraction system for mixing and extraction includes: using a mechanical stirrer to stir and mix at a speed of 250~300 rpm for 2~3 min; the extraction is a 3~4 stage countercurrent extraction.

[0037] Furthermore, the desorption temperature is 25~35℃, the desorption time is 10~15min, and the desorption stage is 2.

[0038] Furthermore, the chemical composition and mass percentage of the lepidolite concentrate are as follows: 1.16% Li, 0.003% Tl, 1.39% F, 0.67% Na, 0.06% Mg, 12.11% Al, 20.56% Si, 0.13% P, 0.02% S, 0.02% Cl, 4.97% K, 0.07% Ca, 0.0 0.04% Nb, 0.009% W, 39.47% O, 0.02% Ti, 0.005% Cr, 0.24% Mn, 0.49% Fe, 0.003% Cu, 0.032% Zn, 0.006% Ga, 0.01% As, 0.61% Rb, 0.001% Zr, 0.06% Cs, and 0.009% Pb.

[0039] The present invention also provides lithium or thallium recovered by the above method, wherein the extraction rate of lithium is greater than 97% and the extraction rate of thallium is greater than 99%.

[0040] This invention also provides the application of the above method in the extraction of lithium and / or thallium from lepidolite concentrate.

[0041] The beneficial effects of this invention are: This invention provides a method for extracting lithium and thallium from lepidolite concentrate. First, the lepidolite concentrate is ground to significantly increase the specific surface area of ​​the particles, mechanically breaking down the interlayer spacing, thus releasing the lithium and thallium ions (Li). + 、Tl +It is easier to contact with external reagents. Then, by combining the chemical chelation activation of the complex acid mixed complex system, the dispersion and anti-agglomeration effect of the surfactant, and the cavitation impact effect of ultrasound, deep activation of lepidolite under mild conditions below 100℃ is achieved. Then, by leaching and filtering the lepidolite concentrate slurry after ultrasonic cracking, a lithium-containing leachate is obtained. Then, extraction is carried out in two steps. First, the complexing agent, diluent and extractant of thallium are mixed to construct a first polar extraction system. Then, the lithium-containing leachate is added to the first polar extraction system for mixing and extraction to obtain a thallium-rich organic phase and raffinate. Thallium is extracted separately to form a thallium-rich organic phase. Then, the thallium-rich organic phase is desorbed to obtain a thallium-rich solution. Secondly, a lithium complexing agent, diluent, and extractant are mixed to construct a second polar extraction system. The raffinate is then added to this system for mixing and dedicated lithium extraction, forming a lithium complex. Finally, the lithium complex is purified and eluted to obtain a lithium-rich solution. This entire process, through a clever design of stepwise activation and extraction, achieves effective separation of lithium and thallium. It recovers both valuable elements while avoiding interference between the two metal ions during extraction, thus improving the purity of the final product. Ultimately, this achieves efficient activation of lepidolite, resulting in efficient extraction and targeted recovery of lithium and thallium. Attached Figure Description

[0042] Figure 1 The image shows the elemental analysis results of lepidolite concentrate. Figure 2 The figure shows the detection results of thallium toxicity leaching in the lepidolite leaching residue obtained in Comparative Example 3. Figure 3 The image shows the results of third-party testing for the toxicity of thallium and beryllium in the leaching residue of lepidolite in Example 3. Detailed Implementation

[0043] The present invention will be further described in detail below through embodiments, but in no way is the invention limited.

[0044] In the following examples, the element Li in the lepidolite concentrate was quantitatively analyzed by atomic absorption spectrometry (AAS), and the Li content was found to be approximately 1.16 wt%. Elements after boron (B) in the lepidolite concentrate were determined by XRF (X-ray fluorescence spectrometry), and the results are as follows. Figure 1 As shown. By Figure 1It is known that the chemical composition and mass percentage of other elements in lepidolite concentrate are approximately: Tl content approximately 0.003 wt%, F 1.39%, Na 0.67%, Mg 0.06%, Al 12.11%, Si 20.56%, P 0.13%, S 0.02%, Cl 0.02%, K 4.97%, Ca 0.07%, and so on. % Nb, 0.009% W, 39.47% O, 0.02% Ti, 0.005% Cr, 0.24% Mn, 0.49% Fe, 0.003% Cu, 0.032% Zn, 0.006% Ga, 0.01% As, 0.61% Rb, 0.001% Zr, 0.06% Cs, 0.009% Pb.

[0045] The technical principles of the present invention are as follows: In the technical solution of this invention: firstly, lepidolite is ground to a D90 < 45µm, which significantly increases the particle specific surface area. The interlayer spacing is mechanically broken down, allowing the metal ions (Li) present in the interlayer to be released. + 、Tl + It is easier to come into contact with external reagents.

[0046] Citric acid / succinic acid forms metal-citric acid / succinic acid complexes under weakly acidic conditions (pH≈3~4) due to their carboxyl groups, which can efficiently chelate Li between the layers of lepidolite. + 、Tl + The isocations achieve and partially dissolve, thereby weakening the charge balance of layered aluminosilicates. Ammonitriletriacetic acid (NTA) is a tricarboxylic acid amine ligand capable of forming stable five-membered ring chelates (stability constants around 10) with most metal ions. 6 ~10 8 (Scope). The strong complexing effect of NTA extracts more interlayer metal ions, significantly reducing the electrostatic attraction of interlayer bonds. The combined effect of these organic acids and complexing agents is to reduce interlayer adhesion and increase interlayer spacing, providing a chemical prerequisite for subsequent mechanical impact. Sodium dodecyl sulfate / sodium dodecyl sulfonate is an anionic surfactant that adsorbs onto the surface of mineral particles in the slurry, forming a negatively charged layer that significantly reduces solid-liquid interfacial tension and improves the wettability and dispersibility of the slurry. At the same time, the mutual repulsion of negative charges prevents fine particles from re-aggregating during subsequent high-energy impacts, ensuring that the shock wave generated by cavitation can act uniformly on each layered wafer.

[0047] Under ultrasonic treatment conditions of 20–25 kHz frequency, 320–350 W / L power density, and 85–95 °C, a microenvironment with instantaneous pressures exceeding 5000 K and 100 MPa can be generated. The collapse process is accompanied by high-speed microjet and shock waves, which can generate strong shear and impact stresses at the nanoscale. These impact stresses directly tear the Si-O-Al bonds of layered aluminosilicates, forming microcracks and interlayer delamination, which can significantly reduce overall energy consumption and avoid large-scale high-temperature calcination.

[0048] The leachate was then separated by selective complexation, and the pH-controlled dibenzo-18-crown-6-ether (DB-18-crown-6) was used to target Tl. + It exhibits highly selective complexing ability, enabling specific extraction of thallium via tributyl phosphate (TBP) solvent extraction; subsequently, ethyl acetoacetate (EAA) is added at a slightly higher pH to target Li. + Selective complexation and purification via ion exchange resin were performed.

[0049] Crown ethers exhibit high selectivity for heavy metal ions (especially soft acid metals). The ring cavity diameter of DB-18-crown-6 is approximately 3.2 Å, which perfectly matches that of Tl. + (Ionic radius ≈ 1.5 Å) This forms the most stable 1:1 coordination complex. This coordination process is primarily driven by the O→Tl provided by the six ether oxygen atoms in the crown ether molecule. + σ - Coordination can form a relatively rigid 18-member ring-6-oxygen coordination cage, Tl + It is completely encapsulated within the crown ether cavity. Within the pH range of 4–4.5, the Tl content in the solution... + It exists as a free ion and is not easily competitive for coordination by hydrated hydrogen ions; while Li + At this pH, the strongly hydrated ion (Li(H2O)6) remains dominant. + It appears in the form of ) and its coordination ability is inhibited by water molecules.

[0050] The phosphate ester oxygen (P=O) of tributyl phosphate (TBP) can react with DB-18-crown-6•Tl + The complex forms a secondary coordination (TBP→Tl) + This reduces the polarity of the complex, making it more soluble in the organic phase. The extraction process can be considered as: DB-18-crown-6+Tl + [DB-18-crown-6•Tl] + (Coordination) [DB-18-crown-6•Tl] + +2TBP [DB-18-crown-6•Tl•(TBP)2] + (weak ion) This weakly ionic complex is stable in organic phases (such as n-butanol / n-hexane), thus achieving Tl. + Highly efficient separation. Due to Tl + The coordination constant with DB-18-crown-6 is much higher than that of other common metals (such as Li). + Na + K + Under the above pH conditions, almost all Tl + All of them are captured and transferred to the organic phase, achieving "selective extraction".

[0051] Ethyl acetoacetate (EAA) belongs to the β-keto ester class and contains an active α-hydrogen in its molecule. EAA is effective against Li... + The coordination constant is much higher than that of Na. + K + (Because Li) + The radius best matches the geometry of a six-membered ring, thus preferentially forming a complex under the same conditions. At pH 5.5–6.0, a small number of hydroxyl groups are deprotonated to form a ketool enol anion, which provides two oxygen groups (carbonyl O and enol O) as coordination sites. + It forms a six-membered ring coordination with these two oxygen atoms (Li–O–C–C–O), resulting in a relatively neutral Li-EAA complex. Similar examples of Li-EAA six-membered ring coordination have been reported in the literature. Furthermore, when the pH is increased to 5.5–6.0, the H+ in the solution… + As the concentration decreases, the ketone-alcohol tautomerism equilibrium of EAA tilts towards the enol anion, enhancing its coordination ability; simultaneously, Li + Its hydrated shell is slightly loose and easily replaced by ligands.

[0052] D201 resin in an acidic aqueous phase as H + The Li-EAA complex exists in a relatively neutral form in the aqueous phase. Upon entering the resin bed, the complex dissociates, releasing Li... + With resin –SO3 2- Ion exchange occurs (Li + +–SO3 2- H + Li + -SO3 2- +H + ), thereby realizing Li + The resin was then eluted with dilute acid to obtain Li-rich... + Solution.

[0053] The above selective complexation-separation process utilizes the matching between the "hard acid and soft base" properties of metal ions and the ring cavity size of the ligand: firstly, under weakly acidic conditions, the soft acid metal Tl is captured using dibenzo-18-crown-6-ether (DB-18-crown-6). + Extracted to the organic phase by tributyl phosphate (TBP); subsequently, ethyl acetoacetate (EAA) was oxidized to form an enol anion at pH 5.5–6.0, specifically coordinating with the hard acid metal Li. + The process utilizes D201 strong acid sulfonic acid resin for ion exchange purification. The entire process achieves efficient and selective separation of thallium and lithium through the synergistic effect of pH, ligand structure, and extraction / exchange resin.

[0054] The entire process is completed below 100℃, featuring a gentle process that enables recycling. Energy consumption is significantly reduced compared to traditional roasting methods, and the amount of acid used is drastically decreased, resulting in a significant reduction in waste liquid generation. Therefore, this technology overcomes the energy consumption and equipment limitations of high-temperature roasting, as well as the corrosion and low selectivity issues of strong acid leaching. Through selective complexing agents and solid-phase extraction technology, valuable metals such as thallium are efficiently recovered, reducing environmental pressure on subsequent lithium smelting slag treatment and improving resource utilization. Ultimately, a lithium extraction rate of ≥97% and a thallium extraction rate of ≥99% can be achieved, providing a feasible path for the green and efficient utilization of lithium mica resources.

[0055] Example 1 A method for extracting lithium thallium from lepidolite using an ultrasonic cavitation-selective complexation synergistic system includes the following steps: S1. Complex acid mixed complex system: composed of citric acid, nitric acid triacetic acid (NTA), and water. In the complex acid mixed complex system, the concentration of citric acid is 0.2 mol / L, the concentration of nitric acid triacetic acid (NTA) is 0.06 mol / L, and the molar ratio of citric acid to nitric acid triacetic acid (NTA) is 3:1. Citric acid, nitric acid triacetic acid, and water are mixed to form a transparent complex mother liquor. The transparent complex mother liquor is then placed in a constant temperature environment of 45℃ and stirred until homogeneous to obtain the complex acid mixed complex system, which is ready for use. S2. Add 1000g of lepidolite concentrate to a sand mill with water for ball milling. Set the main shaft speed of the sand mill to 150r / min, the stirring axis speed to 0.5m / s, the material-to-ball ratio to 10:1, the grinding concentration to 65%, and the ball milling time to 30min. Five minutes after the start of ball milling (when the lepidolite concentrate is ground to D90 < 45µm), inject the composite acid mixed complex system prepared in S1 into the mill inlet using a metering pump, controlling the mass ratio of lepidolite concentrate to the composite acid mixed complex system to 1:1.2. Simultaneously add the surfactant sodium dodecyl sulfate, with a dosage of 0.5wt% based on the lepidolite concentrate (i.e., the amount of sodium dodecyl sulfate added is 0.5% of the mass of the lepidolite concentrate). Continue ball milling until the set time is reached to form a uniform lepidolite concentrate solids slurry.

[0056] S3. In a titanium alloy reactor, the lepidolite concentrate slurry obtained in S2 is ultrasonically treated to achieve deep disintegration of layered silicates. The specific steps are as follows: The lepidolite concentrate slurry obtained in S2 is fed into a Ti-6Al-4V titanium alloy reactor for ultrasonic treatment to obtain an ultrasonically disintegrated lepidolite concentrate slurry. The Ti-6Al-4V titanium alloy reactor is 5mm thick, the ultrasonic frequency is controlled at 20kHz, the power density is 320W / L, the treatment temperature is controlled at 85℃, the treatment time is 50 minutes, continuous flow treatment is used, the slurry flow rate is 10L / min, and the cavitation number is controlled at 1.0.

[0057] S4. The slurry of lepidolite concentrate with high solid content obtained from S3 by ultrasonic cracking is transferred into a water bath. Lithium is extracted using a three-stage countercurrent water leaching process. The leaching liquid-to-solid ratio is set at 4:1, the leaching temperature is 90℃, the leaching time is 200min, and the stirring speed is 300rpm. After leaching, the mixture is filtered to obtain lithium-containing leachate and lepidolite leaching residue. S5. The complexing agent, diluent, and extractant of thallium are mixed to construct a first polar extraction system, wherein dibenzo-18-crown-6-ether (DB-18-crown-6) is used as the complexing agent of thallium, the concentration of dibenzo-18-crown-6-ether in the first polar extraction system is 1.0 g / L, industrial pure sulfonated kerosene is used as the diluent, and 18 vol% tributyl phosphate (TBP) solution is used as the extractant; the pH value of the first polar extraction system is controlled to be 4.2 by adding hydrochloric acid and sodium hydroxide; the saponification rate of the first polar extraction system is controlled to be 50%, and the total phase (i.e., volume) ratio of aqueous phase to organic phase is 3.2:1; wherein, 18 vol% tributyl phosphate (TBP) solution refers to the volume percentage concentration of tributyl phosphate in the organic phase being 18%, and the solvent in the tributyl phosphate (TBP) solution is aviation kerosene.

[0058] Then, the lithium-containing leachate obtained in S4 was added to the first polar extraction system and mixed using a mechanical stirrer at a speed of 250 rpm for 2.5 minutes. Three-stage countercurrent extraction was performed, followed by a standing separation time of 18 minutes to achieve selective complexation of thallium, resulting in a thallium-rich organic phase and raffinate.

[0059] The core of three-stage countercurrent extraction is to achieve deep separation of target components (or impurities) through multi-stage reverse contact between the organic phase and the aqueous phase. Compared with single-stage extraction, it can significantly improve separation efficiency and reduce the amount of extractant used.

[0060] S6. Desorb the thallium-rich organic phase obtained in S5. After desorption, let it stand for 18 minutes, and collect the lower thallium-rich solution and the upper thallium-depleted organic phase using a separatory funnel to obtain a thallium-rich solution and a thallium-depleted organic phase. The desorbent is a 1.0 mol / L H2SO4 solution with water as the solvent. The desorption temperature is 25℃, the desorption time is 10 min, the ratio of the volume of the thallium-rich organic phase to the volume of the desorbent is 3:1, and the desorption stage is 2.

[0061] S7. Regenerate the thallium-depleted organic phase obtained after desorption in S6 (the organic phase system in which thallium ions have been removed after desorption). The specific steps are as follows: Wash and neutralize the thallium-depleted organic phase to pH 6.5 with a Na2CO3 solution with a concentration of 0.1 mol / L. Then filter it through a 100-mesh silica gel column to remove degradation products and solid particles. Finally, perform vacuum degassing treatment at -0.08 MPa to remove dissolved gases. The vacuum degassing time is 30 min to obtain the regenerated organic phase, which can be recycled for thallium extraction.

[0062] The solvent in the Na2CO3 solution is water.

[0063] S8. A lithium complexing agent, diluent, and extractant are mixed to construct a second polar extraction system. The lithium complexing agent is 1.5 g / L ethyl acetoacetate (EAA), industrial pure sulfonated kerosene is used as the diluent, and 18 vol% tributyl phosphate (TBP) solution is used as the extractant. The pH of the second polar extraction system is controlled to be 5.5 by adding hydrochloric acid and sodium hydroxide. The saponification rate of the second polar extraction system is controlled to be 50%, and the total ratio of aqueous phase to organic phase (A / O) is 3.2:1. Then, the raffinate obtained in S5 was added to the second polar extraction system for lithium complexation and extraction. The mixture was stirred at 250 rpm using a mechanical stirrer. The mixing time of the two phases was 2 min, and the settling time after contact was 15 min. The entire process adopted a three-stage countercurrent extraction process to ensure efficient enrichment of lithium and obtain lithium complexes.

[0064] S9. The lithium complex obtained in S8 is transferred into an ion exchange system and purified using D201 strong acid sulfonic acid resin to obtain D201 strong acid sulfonic acid resin after lithium adsorption. The specific parameters for purification are as follows: D201 strong acid sulfonic acid resin uses the chlorinated form (Cl... - (form) The ion exchange column in the ion exchange system has a column diameter of 150 mm and a height of 1200 mm. The amount of D201 strong acid sulfonic acid resin packing is calculated based on the bed volume. Each cubic meter of mother liquor (lithium complex) corresponds to 8 L of D201 strong acid sulfonic acid resin. The flow rate is 30 L / h, and the operation cycle time is 7 hours / cycle.

[0065] S10. Using a dilute hydrochloric acid solution with a concentration of 0.1 mol / L as the eluent, the D201 strongly acidic sulfonic acid resin after lithium adsorption is eluted to obtain Li-rich... + The solution has a pH of 1.5.

[0066] The solvent for the dilute hydrochloric acid solution is water.

[0067] During elution, the elution flow rate is controlled at 0.5 BV·h. -1 (0.5 bed volume / hour); collect 2 BV of eluent, monitoring Li during the process. + The concentration was increased until it dropped below the baseline, at which point collection was stopped, and the resulting 2 BV eluent was Li-rich. + Solution.

[0068] S11 and D201 strong acid sulfonic acid resins undergo chemical regeneration every 25 operating cycles. The specific steps are as follows: First, soak the D201 strong acid sulfonic acid resin in a mixed solution of 5wt% NaCl and 2wt% NaOH for 4 hours to remove organic contaminants adsorbed on the surface of the resin. Then, treat the D201 strong acid sulfonic acid resin with a 0.5mol / L HCl solution to remove inorganic precipitates, obtaining regenerated D201 strong acid sulfonic acid resin. After regeneration, the D201 strong acid sulfonic acid resin can be reused in the ion exchange purification process. The solvent in the 5wt% NaCl and 2wt% NaOH mixed solution is water.

[0069] Example 2 A method for extracting lithium thallium from lepidolite using an ultrasonic cavitation-selective complexation synergistic system includes the following steps: S1. Complex acid mixed complex system: composed of succinic acid, nitric acid triacetic acid (NTA), and water. In the complex acid mixed complex system, the concentration of succinic acid is 0.18 mol / L, the concentration of nitric acid triacetic acid (NTA) is 0.07 mol / L, and the molar ratio of succinic acid to nitric acid triacetic acid (NTA) is 3:1. Succinic acid, nitric acid triacetic acid, and water are mixed to form a transparent complex mother liquor. The transparent complex mother liquor is then placed in a constant temperature environment of 48℃ and stirred until uniformly mixed to obtain the complex acid mixed complex system for later use. S2. 1000g of lepidolite concentrate is added to a sand mill with water for ball milling. The main shaft speed of the sand mill is set to 170r / min, the stirring axis speed is 0.8m / s, the material-to-ball ratio is 10:1, the grinding concentration is 66%, and the ball milling time is 40min. Five minutes after the start of ball milling (when the lepidolite concentrate is ground to D90 < 45µm), the composite acid mixed complex system prepared in S1 is injected into the mill inlet through a metering pump. The mass ratio of lepidolite concentrate to the composite acid mixed complex system is controlled to be 1:1.3. Sodium dodecyl sulfate surfactant is added simultaneously. Based on the lepidolite concentrate, the amount of sodium dodecyl sulfate is 0.5wt% (i.e., the amount of sodium dodecyl sulfate added is 0.5% of the mass of lepidolite concentrate). Ball milling continues until the set time is reached to form a uniform lepidolite concentrate solids slurry.

[0070] S3. In a titanium alloy reactor, the lepidolite concentrate slurry obtained in S2 is treated with ultrasound to achieve deep disintegration of layered silicates. The specific steps are as follows: The lepidolite concentrate slurry obtained in S2 is fed into a Ti-6Al-4V titanium alloy reactor for ultrasonic treatment to obtain an ultrasonically disintegrated lepidolite concentrate slurry. The Ti-6Al-4V titanium alloy reactor is 5mm thick, the ultrasonic frequency is controlled at 23kHz, the power density is 330W / L, the treatment temperature is controlled at 90℃, the treatment time is 55 minutes, continuous flow treatment is used, the slurry flow rate is 9L / min, and the cavitation number is controlled at 1.1.

[0071] S4. The slurry of lepidolite concentrate with high solid content after ultrasonic cracking in S3 is transferred into a water bath. Lithium is extracted using a three-stage countercurrent water leaching process. The leaching liquid-to-solid ratio is set to 4:1, the leaching temperature is 90℃, the leaching time is 240 min, and the stirring speed is 350 rpm. After leaching, the mixture is filtered to obtain lithium-containing leachate and lepidolite leaching residue. S5. The thallium complexing agent, diluent, and extractant are mixed to construct a first polar extraction system, wherein dibenzo-18-crown-6-ether (DB-18-crown-6) is used as the thallium complexing agent, the concentration of dibenzo-18-crown-6-ether in the first polar extraction system is 0.9 g / L, industrial pure sulfonated kerosene is used as the diluent, and 17 vol% tributyl phosphate (TBP) solution is used as the extractant; the pH value of the first polar extraction system is controlled to be 4.3 by adding hydrochloric acid and sodium hydroxide; the saponification rate of the first polar extraction system is controlled to be 65%, and the total phase (i.e., volume) ratio of aqueous phase to organic phase is 2.5:1; wherein, 17 vol% tributyl phosphate (TBP) solution refers to a volume percentage concentration of tributyl phosphate in the organic phase of 17%, and the solvent in the tributyl phosphate (TBP) solution is aviation kerosene.

[0072] Then, the lithium-containing leachate obtained in S4 was added to the first polar extraction system and mixed using a mechanical stirrer at a speed of 280 rpm for 2 minutes. Four-stage countercurrent extraction was performed, followed by a standing separation time of 17 minutes to achieve selective complexation of thallium, resulting in a thallium-rich organic phase and raffinate.

[0073] Among them, the core of the four-stage countercurrent extraction is to achieve deep separation of the target component (or impurities) through multi-stage reverse contact between the organic phase and the aqueous phase. Compared with single-stage extraction, it can significantly improve separation efficiency and reduce the amount of extractant used.

[0074] S6. Desorb the thallium-rich organic phase obtained in S5. After desorption, let it stand for 18 minutes, and collect the lower thallium-rich solution and the upper thallium-depleted organic phase using a separatory funnel to obtain a thallium-rich solution and a thallium-depleted organic phase. The desorbent is a 1.2 mol / L H2SO4 solution with water as the solvent. The desorption temperature is 30℃, the desorption time is 12 min, the ratio of the volume of the thallium-rich organic phase to the volume of the desorbent is 4:1, and the desorption stage is 2.

[0075] S7. Regenerate the thallium-depleted organic phase obtained after desorption in S6 (the organic phase system in which thallium ions have been removed after desorption). The specific steps are as follows: Wash and neutralize the thallium-depleted organic phase to pH 7.0 with a Na2CO3 solution with a concentration of 0.1 mol / L. Then filter it through a 120-mesh silica gel column to remove degradation products and solid particles. Finally, perform vacuum degassing treatment at -0.08 MPa to remove dissolved gases. The vacuum degassing time is 35 min to obtain the regenerated organic phase, which can be recycled for thallium extraction.

[0076] The solvent in the Na2CO3 solution is water.

[0077] S8. A lithium complexing agent, diluent, and extractant are mixed to construct a second polar extraction system. The lithium complexing agent is 1.7 g / L ethyl acetoacetate (EAA), industrial pure sulfonated kerosene is used as the diluent, and a 17 vol% tributyl phosphate (TBP) solution is used as the extractant. The pH of the second polar extraction system is controlled to be 5.5 by adding hydrochloric acid and sodium hydroxide. The saponification rate of the second polar extraction system is controlled to be 65%, and the total ratio of aqueous phase to organic phase (A / O) is 2.5:1. Then, the raffinate obtained in S5 was added to the second polar extraction system for lithium complexation and extraction. The mixture was stirred at 280 rpm using a mechanical stirrer. The mixing time of the two phases was 2 minutes, and the settling time after contact was 17 minutes. The entire process adopted a 4-stage countercurrent extraction process to ensure efficient enrichment of lithium and obtain lithium complexes.

[0078] S9. The lithium complex obtained in S8 is transferred into an ion exchange system and purified using D201 strong acid sulfonic acid resin to obtain D201 strong acid sulfonic acid resin after lithium adsorption. The specific parameters for purification are as follows: D201 strong acid sulfonic acid resin uses the chlorinated form (Cl... - (form) The ion exchange column in the ion exchange system has a column diameter of 180 mm and a height of 1000 mm. The amount of D201 strong acid sulfonic acid resin packing is calculated by bed volume. Each cubic meter of mother liquor (lithium complex) corresponds to 6 L of D201 strong acid sulfonic acid resin. The flow rate is 25 L / h, and the operation cycle time is 6.5 hours / cycle.

[0079] S10. Using a dilute hydrochloric acid solution with a concentration of 0.3 mol / L as the eluent, the D201 strongly acidic sulfonic acid resin after lithium adsorption is eluted to obtain Li-rich... + The solution has a pH of 1.2.

[0080] The solvent for the dilute hydrochloric acid solution is water.

[0081] During elution, the elution flow rate is controlled at 0.7 BV·h. -1 (0.7 bed volume / hour); collect 2.5 BV of eluent, monitoring Li during the process. + The concentration was increased until it dropped below the baseline, at which point collection was stopped, and the resulting 2 BV eluent was Li-rich. + Solution.

[0082] S11 and D201 strong acid sulfonic acid resins undergo chemical regeneration every 28 operating cycles. The specific steps are as follows: First, soak the D201 strong acid sulfonic acid resin in a mixed solution of 5wt% NaCl and 2wt% NaOH for 4 hours to remove organic contaminants adsorbed on the surface of the resin. Then, treat the D201 strong acid sulfonic acid resin with a 0.5mol / L HCl solution to remove inorganic precipitates, obtaining regenerated D201 strong acid sulfonic acid resin. After regeneration, the D201 strong acid sulfonic acid resin can be reused in the ion exchange purification process. The solvent in the 5wt% NaCl and 2wt% NaOH mixed solution is water.

[0083] Example 3 A method for extracting lithium thallium from lepidolite using an ultrasonic cavitation-selective complexation synergistic system includes the following steps: S1. Complex acid mixed complex system: Composed of citric acid, succinic acid, nitric acid (NTA), and water. In the complex acid mixed complex system, the total concentration of citric acid and succinic acid is 0.2 mol / L, the concentration of nitric acid (NTA) is 0.08 mol / L, and the molar ratio of citric acid, succinic acid, and nitric acid (NTA) is 2:1:1. Citric acid, succinic acid, nitric acid, and water are mixed to form a transparent complex mother liquor. The transparent complex mother liquor is then placed in a constant temperature environment of 50℃ and stirred until homogeneous to obtain the complex acid mixed complex system, which is ready for use. S2. 1000g of lepidolite concentrate is added to a sand mill with water for ball milling. The main shaft speed of the sand mill is set to 200r / min, the stirring axis speed is 1.5m / s, the material-to-ball ratio is 10:1, the grinding concentration is 70%, and the ball milling time is 50min. Five minutes after the start of ball milling (when the lepidolite concentrate is ground to D90 < 45µm), the composite acid mixed complex system prepared in S1 is injected into the mill inlet using a metering pump, controlling the mass ratio of lepidolite concentrate to the composite acid mixed complex system to be 1:1.5. Simultaneously, sodium dodecyl sulfate, a surfactant, is added at a rate of 0.5wt% (i.e., the amount of sodium dodecyl sulfate added is 0.5% of the mass of the lepidolite concentrate) based on the lepidolite concentrate. Ball milling continues until the set time is reached, forming a uniform slurry with a high solids content.

[0084] S3. In a titanium alloy reactor, the lepidolite concentrate slurry obtained in S2 is ultrasonically treated to achieve deep disintegration of layered silicates. The specific steps are as follows: The lepidolite concentrate slurry obtained in S2 is fed into a Ti-6Al-4V titanium alloy reactor for ultrasonic treatment to obtain an ultrasonically disintegrated lepidolite concentrate slurry. The Ti-6Al-4V titanium alloy reactor is 5mm thick, the ultrasonic frequency is controlled at 25kHz, the power density is 350W / L, the treatment temperature is controlled at 95℃, the treatment time is 60 minutes, continuous flow treatment is used, the slurry flow rate is 11L / min, and the cavitation number is controlled at 1.2.

[0085] S4. The slurry of lepidolite concentrate with high solid content obtained from S3 by ultrasonic cracking is transferred into a water bath. Lithium is extracted using a four-stage countercurrent water leaching process. The leaching liquid-to-solid ratio is set to 5:1, the leaching temperature is 95℃, the leaching time is 180 min, and the stirring speed is 400 rpm. After leaching, the mixture is filtered to obtain lithium-containing leachate and lepidolite leaching residue. S5. The complexing agent, diluent, and extractant of thallium are mixed to construct a first polar extraction system, wherein dibenzo-18-crown-6-ether (DB-18-crown-6) is used as the complexing agent of thallium, the concentration of dibenzo-18-crown-6-ether in the first polar extraction system is 1.2 g / L, industrial pure sulfonated kerosene is used as the diluent, and 20 vol% tributyl phosphate (TBP) solution is used as the extractant; the pH value of the first polar extraction system is controlled to be 4.5 by adding hydrochloric acid and sodium hydroxide; the saponification rate of the first polar extraction system is controlled to be 40%, and the total phase (i.e., volume) ratio of aqueous phase to organic phase is 4:1; wherein, 20 vol% tributyl phosphate (TBP) solution means that the volume percentage concentration of tributyl phosphate in the organic phase is 20%, and the solvent in the tributyl phosphate (TBP) solution is aviation kerosene.

[0086] Then, the lithium-containing leachate obtained in S4 was added to the first polar extraction system and mixed using a mechanical stirrer at a speed of 300 rpm for 3 minutes. Four-stage countercurrent extraction was performed, followed by a standing separation time of 20 minutes to achieve selective complexation of thallium, resulting in a thallium-rich organic phase and raffinate.

[0087] Among them, the core of the four-stage countercurrent extraction is to achieve deep separation of the target component (or impurities) through multi-stage reverse contact between the organic phase and the aqueous phase. Compared with single-stage extraction, it can significantly improve separation efficiency and reduce the amount of extractant used.

[0088] S6. Desorb the thallium-rich organic phase obtained in S5. After desorption, let it stand for 18 minutes, and collect the lower thallium-rich solution and the upper thallium-depleted organic phase using a separatory funnel to obtain a thallium-rich solution and a thallium-depleted organic phase. The desorbent is a 1.5 mol / L H2SO4 solution with water as the solvent. The desorption temperature is 35℃, the desorption time is 15 min, the ratio of the volume of the thallium-rich organic phase to the volume of the desorbent is 5:1, and the desorption stage is 2.

[0089] S7. Regenerate the thallium-depleted organic phase obtained after desorption in S6 (the organic phase system in which thallium ions have been removed after desorption). The specific steps are as follows: Wash and neutralize the thallium-depleted organic phase to pH 7.0 with a Na2CO3 solution with a concentration of 0.1 mol / L. Then filter it through a 150-mesh silica gel column to remove degradation products and solid particles. Finally, perform vacuum degassing treatment at -0.08 MPa to remove dissolved gases. The vacuum degassing time is 40 min to obtain the regenerated organic phase, which can be recycled for thallium extraction.

[0090] The solvent in the Na2CO3 solution is water.

[0091] S8. A lithium complexing agent, diluent, and extractant are mixed to construct a second polar extraction system. The lithium complexing agent is 2.0 g / L ethyl acetoacetate (EAA), industrial pure sulfonated kerosene is used as the diluent, and a 20 vol% tributyl phosphate (TBP) solution is used as the extractant. The pH of the second polar extraction system is controlled to be 6.0 by adding hydrochloric acid and sodium hydroxide. The saponification rate of the second polar extraction system is controlled to be 40%, and the total ratio of aqueous phase to organic phase (A / O) is 4:1. Then, the raffinate obtained in S5 was added to the second polar extraction system for lithium complexation extraction. The mixture was stirred at 300 rpm using a mechanical stirrer. The mixing time of the two phases was 3 minutes, and the settling time after contact was 20 minutes. The entire process adopted a 4-stage countercurrent extraction process to ensure efficient enrichment of lithium and obtain lithium complexes.

[0092] S9. The lithium complex obtained in S8 is transferred into an ion exchange system and purified using D201 strong acid sulfonic acid resin to obtain D201 strong acid sulfonic acid resin after lithium adsorption. The specific parameters for purification are as follows: D201 strong acid sulfonic acid resin uses the chlorinated form (Cl... -(form) The ion exchange column in the ion exchange system has a column diameter of 200 mm and a height of 1500 mm. The amount of D201 strong acid sulfonic acid resin packing is calculated by bed volume. Each cubic meter of mother liquor (lithium complex) corresponds to 8 L of D201 strong acid sulfonic acid resin. The flow rate is 35 L / h, and the operation cycle time is 7 hours / cycle.

[0093] S10. Using a dilute hydrochloric acid solution with a concentration of 0.5 mol / L as the eluent, the D201 strongly acidic sulfonic acid resin after lithium adsorption is eluted to obtain Li-rich... + The solution has a pH of 1.0.

[0094] The solvent for the dilute hydrochloric acid solution is water.

[0095] During elution, the elution flow rate is controlled at 0.8 BV·h. -1 (0.8 bed volume / hour); collect 3 BV eluent, monitoring Li during the process. + The concentration was increased until it dropped below the baseline, at which point collection was stopped, and the resulting 2 BV eluent was Li-rich. + Solution.

[0096] S11 and D201 strong acid sulfonic acid resins undergo chemical regeneration every 30 operating cycles. The specific steps are as follows: First, soak the D201 strong acid sulfonic acid resin in a mixed solution of 5wt% NaCl and 2wt% NaOH for 4 hours to remove organic contaminants adsorbed on the surface of the resin. Then, treat the D201 strong acid sulfonic acid resin with a 0.5mol / L HCl solution to remove inorganic precipitates, obtaining regenerated D201 strong acid sulfonic acid resin. After regeneration, the D201 strong acid sulfonic acid resin can be reused in the ion exchange purification process. The solvent in the 5wt% NaCl and 2wt% NaOH mixed solution is water.

[0097] Comparative Example 1 A method for extracting lithium thallium from lepidolite using an ultrasonic cavitation-selective complexation synergistic system includes the following steps: S1. Prepare a sulfuric acid solution with a concentration of 1.0 mol / L. The solvent in the sulfuric acid solution is water.

[0098] S2. 1000g of lepidolite concentrate is ball-milled in a sand mill with water. The main shaft speed of the sand mill is set to 200 r / min, the stirring axis speed is 1.5m / s, the material-to-ball ratio is 10:1, the grinding concentration is 70%, and the ball-milling time is 50min. Five minutes after the start of ball milling, a 1.0mol / L sulfuric acid solution is injected into the mill inlet through a metering pump, controlling the mass ratio of lepidolite concentrate to sulfuric acid to be 1:1.5. Sodium dodecyl sulfate, a surfactant, is added simultaneously. Based on the lepidolite concentrate, the amount of sodium dodecyl sulfate is 0.5wt% (i.e., the amount of sodium dodecyl sulfate added is 0.5% of the mass of lepidolite concentrate). Ball milling continues until the set time is reached to form a uniform lepidolite concentrate solids slurry.

[0099] S3. In a titanium alloy reactor, the lepidolite concentrate slurry obtained in S2 is treated with ultrasound to achieve deep disintegration of layered silicates. The specific steps are as follows: The lepidolite concentrate slurry obtained in S2 is fed into a Ti-6Al-4V titanium alloy reactor for ultrasonic treatment to obtain an ultrasonically disintegrated lepidolite concentrate slurry. The Ti-6Al-4V titanium alloy reactor is 5mm thick, the ultrasonic frequency is controlled at 25kHz, the power density is 350W / L, the treatment temperature is controlled at 95℃, the treatment time is 60 minutes, continuous flow treatment is used, the slurry flow rate is 11L / min, and the cavitation number is controlled at 1.2.

[0100] S4. The slurry of lepidolite concentrate with high solid content obtained from S3 after ultrasonic cracking is added to a water bath. Lithium is extracted using a four-stage countercurrent water leaching process. The leaching liquid-to-solid ratio is set to 5:1, the leaching temperature is 95℃, the leaching time is 180 min, and the stirring speed is 400 rpm. After leaching, the mixture is filtered to obtain lithium-containing leachate and lepidolite leaching residue. S5. The thallium complexing agent, diluent, and extractant are mixed to construct a first polar extraction system. Dibenzo-18-crown-6-ether (DB-18-crown-6) is used as the thallium complexing agent, with a concentration of 1.2 g / L. Industrial pure sulfonated kerosene is used as the diluent, and a 20 vol% tributyl phosphate (TBP) solution is used as the extractant. The pH of the first polar extraction system is controlled to 4.5 by adding hydrochloric acid and sodium hydroxide. The saponification rate of the first polar extraction system is controlled to be 65%, and the total ratio (A / O) of the aqueous phase to the organic phase is 3.8:1. The 20 vol% tributyl phosphate (TBP) solution refers to a 20% volume percentage concentration of tributyl phosphate in the organic phase, and the solvent in the TBP solution is aviation kerosene.

[0101] The lithium-containing leachate obtained in S4 was then added to the first polar extraction system and mixed at 300 rpm for 3 minutes using a mechanical stirrer. A four-stage countercurrent extraction process was employed, followed by a settling time of 20 minutes to selectively complex thallium, yielding a thallium-rich organic phase and raffinate.

[0102] S6. Desorb the thallium-rich organic phase obtained in S5. After desorption, let it stand for 18 minutes, and collect the lower thallium-rich solution and the upper thallium-depleted organic phase using a separatory funnel to obtain a thallium-rich solution and a thallium-depleted organic phase. The desorbent is a 1.5 mol / L H2SO4 solution with water as the solvent. The desorption temperature is 35℃, the desorption time is 15 min, the ratio of thallium-rich organic phase to desorbent is 5:1, and the desorption order is 2.

[0103] S7. Regenerate the thallium-depleted organic phase obtained after desorption in S6. The specific steps are as follows: wash and neutralize to pH 7.0 with a Na2CO3 solution with a concentration of 0.1 mol / L, filter through a 150-mesh silica gel column to remove degradation products and solid particles, and remove dissolved gases by vacuum degassing at -0.08 MPa for 40 min to obtain the regenerated organic phase, which can be recycled for thallium extraction.

[0104] The solvent in the Na2CO3 solution is water.

[0105] S8. A lithium complexing agent, diluent, and extractant are mixed to construct a second polar extraction system. The lithium complexing agent is 2.0 g / L ethyl acetoacetate (EAA), industrial pure sulfonated kerosene is used as the diluent, and a 20 vol% tributyl phosphate (TBP) solution is used as the extractant. The pH of the second polar extraction system is controlled to be 6.0 by adding hydrochloric acid and sodium hydroxide. The saponification rate of the second polar extraction system is controlled to be 65%, and the total ratio of aqueous phase to organic phase (A / O) is 3.8:1. Then, the raffinate obtained in S5 was added to the second polar extraction system for lithium extraction by complexation. The mixture was stirred at 300 rpm using a mechanical stirrer for 3 minutes and allowed to stand for 20 minutes to separate into layers. The entire process used a 4-stage countercurrent extraction process.

[0106] S9. The lithium complex obtained in S8 is transferred into an ion exchange system and purified using D201 strong acid sulfonic acid resin to obtain D201 strong acid sulfonic acid resin after lithium adsorption. The specific parameters for purification are as follows: D201 strong acid sulfonic acid resin uses the chlorinated form (Cl... -(Form), column diameter 200mm, height 1500mm, D201 strong acid sulfonic acid resin filler is calculated by bed volume, 8L resin per cubic meter of mother liquor (lithium complex), flow rate is 35L / h, and operation cycle time is 7 hours / cycle.

[0107] S10. Using a dilute hydrochloric acid solution with a concentration of 0.5 mol / L as the eluent, the D201 strongly acidic sulfonic acid resin after lithium adsorption is eluted to obtain Li-rich... + The solution has a pH of 1.0.

[0108] In this case, the solvent for the dilute hydrochloric acid solution is water. During elution, the elution flow rate is controlled at 0.8 BV·h. -1 Collect 3BV eluent and monitor Li during the process. + The concentration was increased until it dropped below baseline, at which point collection was stopped, and the resulting 3 BV eluent was Li-rich. + Solution.

[0109] S11 and D201 strong acid sulfonic acid resins undergo chemical regeneration every 30 operating cycles. The specific steps are as follows: First, soak the D201 strong acid sulfonic acid resin in a mixed solution of 5wt% NaCl and 2wt% NaOH for 4 hours to remove organic contaminants. Then, treat the D201 strong acid sulfonic acid resin with a 0.5mol / L HCl solution to remove inorganic precipitates, obtaining regenerated D201 strong acid sulfonic acid resin. After regeneration, the D201 strong acid sulfonic acid resin can be reused in the ion exchange purification process. The solvent in the 5wt% NaCl and 2wt% NaOH mixed solution is water.

[0110] Comparative Example 2 A method for selectively activating lithium mica to extract lithium thallium includes the following steps: S1. Complex acid mixed complex system: composed of succinic acid, nitric acid triacetic acid (NTA), and water. In the complex acid mixed complex system, the concentration of succinic acid is 0.18 mol / L, the concentration of nitric acid triacetic acid (NTA) is 0.07 mol / L, and the molar ratio of succinic acid to nitric acid triacetic acid (NTA) is 3:1. Succinic acid, nitric acid triacetic acid, and water are mixed to form a transparent complex mother liquor. The transparent complex mother liquor is stirred and mixed at 48°C until homogeneous to obtain the complex acid mixed complex system, which is then set aside. S2. 1000g of lepidolite concentrate is ball-milled in a sand mill with water. The main shaft speed of the sand mill is set to 170r / min, the stirring axis speed is 0.8m / s, the material-to-ball ratio is 10:1, the grinding concentration is 66%, and the ball-milling time is 40min. Five minutes after the start of ball milling, the composite acid mixed complex system prepared in S1 is injected into the mill inlet through a metering pump, controlling the mass ratio of lepidolite concentrate to the composite acid mixed complex system to be 1:1.3. Sodium dodecyl sulfate, a surfactant, is added simultaneously. Based on the lepidolite concentrate, the amount of sodium dodecyl sulfate is 0.5wt% (i.e., the amount of sodium dodecyl sulfate added is 0.5% of the mass of lepidolite concentrate). Ball milling continues until the set time is reached to form a uniform lepidolite concentrate solids slurry.

[0111] S3. Add the lepidolite concentrate slurry obtained in S2 to a water bath and extract lithium using a four-stage countercurrent water leaching process. Set the leaching liquid-to-solid ratio to 5:1, the leaching temperature to 95℃, the leaching time to 180 min, and the stirring speed to 400 rpm. After leaching, filter the solution to obtain lithium-containing leachate and lepidolite leaching residue. S4. The thallium complexing agent, diluent, and extractant are mixed to construct a first polar extraction system, wherein dibenzo-18-crown-6-ether (DB-18-crown-6) is used as the thallium complexing agent, the concentration of dibenzo-18-crown-6-ether in the first polar extraction system is 0.9 g / L, industrial pure sulfonated kerosene is used as the diluent, and 17 vol% tributyl phosphate (TBP) solution is used as the extractant; the pH value of the first polar extraction system is controlled to be 4.3 by adding hydrochloric acid and sodium hydroxide; the saponification rate of the first polar extraction system is controlled to be 50%, and the total ratio of aqueous phase to organic phase (A / O) is 7.8:1; wherein, 17 vol% tributyl phosphate (TBP) solution refers to the volume percentage concentration of tributyl phosphate solution in the organic phase being 17%, and the solvent in the tributyl phosphate (TBP) solution is aviation kerosene.

[0112] Then, the lithium-containing leachate obtained in S3 was added to the first polar extraction system and mixed at 280 rpm for 2 minutes using a mechanical stirrer. Four-stage countercurrent extraction was performed, followed by a standing separation time of 17 minutes to achieve selective complexation of thallium, resulting in a thallium-rich organic phase and raffinate.

[0113] S5. Desorb the thallium-rich organic phase obtained in S4. After desorption, let it stand for 18 minutes, and collect the lower thallium-rich solution and the upper thallium-depleted organic phase using a separatory funnel to obtain a thallium-rich solution and a thallium-depleted organic phase. The desorbent is a 1.2 mol / L H2SO4 solution with water as the solvent. The desorption temperature is 30℃, the desorption time is 12 min, the ratio of thallium-rich organic phase to desorbent is 4:1, and the desorption order is 2.

[0114] S6. Regenerate the thallium-depleted organic phase obtained after desorption in S5. The specific steps are as follows: wash and neutralize to pH 7.0 with a Na2CO3 solution with a concentration of 0.1 mol / L. Then filter through a 120-mesh silica gel column to remove degradation products and solid particles. Finally, remove dissolved gases by vacuum degassing at -0.08 MPa for 35 min to obtain the regenerated organic phase, which can be recycled for thallium extraction.

[0115] The solvent in the Na2CO3 solution is water.

[0116] S7. A lithium complexing agent, diluent, and extractant are mixed to construct a second polar extraction system. The lithium complexing agent is 1.7 g / L ethyl acetoacetate (EAA), industrial pure sulfonated kerosene is used as the diluent, and a 17 vol% tributyl phosphate (TBP) solution is used as the extractant. The pH of the second polar extraction system is controlled to be 5.5 by adding hydrochloric acid and sodium hydroxide. The saponification rate of the second polar extraction system is controlled to be 50%, and the total ratio of aqueous phase to organic phase (A / O) is 7.8:1. Then, the raffinate obtained in S4 was added to the second polar extraction system for lithium complexation extraction. The mixture was stirred at 280 rpm for 2 minutes using a mechanical stirrer, and the settling time was 17 minutes. The entire process was carried out using four-stage countercurrent extraction.

[0117] S8. The lithium complex obtained in S8 is transferred into an ion exchange system and purified using D201 strong acid sulfonic acid resin to obtain D201 strong acid sulfonic acid resin after lithium adsorption. The specific parameters for purification are as follows: D201 strong acid sulfonic acid resin uses the chlorinated form (Cl... - (form) The ion exchange column in the ion exchange system has a column diameter of 180 mm and a height of 1000 mm. The amount of D201 strong acid sulfonic acid resin packing is calculated by bed volume, with 6 L of resin corresponding to each cubic meter of mother liquor (lithium complex). The flow rate is 25 L / h, and the operating cycle time is 6.5 hours / cycle.

[0118] S9. Using a dilute hydrochloric acid solution with a concentration of 0.3 mol / L as the eluent, the D201 strongly acidic sulfonic acid resin after lithium adsorption was eluted to obtain Li-rich...+ The solution has a pH of 1.2.

[0119] The solvent for the dilute hydrochloric acid solution is water.

[0120] During elution, the elution flow rate is controlled at 0.7 BV·h. -1 Collect 2.5 BV of eluent, and monitor Li during the process. + The concentration was increased until it dropped below baseline, at which point collection was stopped, and the resulting 2.5 BV eluent was Li-rich. + Solution.

[0121] S10 and D201 strong acid sulfonic acid resins undergo chemical regeneration every 28 operating cycles. The specific steps are as follows: First, soak the D201 strong acid sulfonic acid resin in a mixed solution of 5wt% NaCl and 2wt% NaOH for 4 hours to remove organic contaminants. Then, treat the D201 strong acid sulfonic acid resin with a 0.5mol / L HCl solution to remove inorganic precipitates, obtaining regenerated D201 strong acid sulfonic acid resin. After regeneration, the D201 strong acid sulfonic acid resin can be reused in the ion exchange purification process. The solvent in the 5wt% NaCl and 2wt% NaOH mixed solution is water.

[0122] Comparative Example 3 A traditional lithium extraction technology using lepidolite sulfate roasting includes the following steps: S1. Mix 1000 g of lepidolite concentrate powder with sodium sulfate, potassium sulfate and calcium oxide, and then place it in a muffle furnace and calcine it at a temperature of 1050℃ for 120 min to obtain the calcined product; wherein, the mass ratio of lepidolite concentrate powder to sodium sulfate, potassium sulfate and calcium oxide is 1:0.5:0.1:0.1. S2. Add the lepidolite concentrate obtained in S1 to a water bath and perform conventional four-stage countercurrent water leaching. The leaching liquid-to-solid ratio is 5:1, the leaching temperature is 95℃, the leaching time is 180 min, the stirring speed is 400 rpm, and the mixture is filtered to obtain lithium-containing leaching liquid and lepidolite leaching residue. S3. Add Ca(OH)2 to the lithium-containing leachate from S2 to adjust the pH value to 12, thereby removing aluminum / fluorine impurities with the precipitant and obtaining the lithium-containing leachate after impurity removal.

[0123] Detection and Analysis 1. Li + Extraction rate and Tl + Extraction rate determination Li + The formula for calculating the extraction rate is shown in equation (I), Tl + The formula for calculating the extraction rate is shown in (II): Tl +Extraction rate = {[Tl in thallium-rich organic phase]} + [Concentration (g / L) × Volume of thallium-rich organic phase (L)] / [Tl in lithium-containing leachate] + [Concentration (g / L) × Volume of lithium-containing leachate (L)] × 100% (I) Li + Extraction rate = {[Li in lithium complex]} + [Concentration (g / L) × Volume of lithium complex (L)] / [(Li in lithium-containing leachate)] + [Concentration (g / L) × Volume of lithium-containing leachate (L)] × 100% (II) Among them, Tl in the thallium-rich organic phase + There are two methods for determining the concentration of Tl: the first is to measure the Tl concentration in the lithium-containing leachate before extraction. + The concentration of Tl was determined after extraction and then measured in the aqueous phase. + The concentration of Tl in the leachate was then determined. + The concentration of Tl in the aqueous phase after extraction + The concentration of Tl in the thallium-rich organic phase was obtained. + The first method involves adjusting the concentration of thallium; the second method involves taking an appropriate amount of thallium-rich organic phase, adding perchloric acid, and oxidizing the organic matter to obtain a thallium-containing product free of organic matter. + The Tl content in the thallium-rich organic phase can be obtained by analyzing the aqueous solution and then measuring it using a standard inductively coupled plasma mass spectrometer (ICP-MS). + The concentration.

[0124] Li in lithium-containing leachate + The concentration was measured using atomic absorption spectrophotometry (AAS).

[0125] Li in lithium complex + The method for determining the concentration of Li is as follows: Take an appropriate amount of lithium complex, add perchloric acid to destroy the organic matter, and obtain Li-containing compounds without organic matter. + The Li content in the lithium complex can be obtained by measuring the aqueous solution using atomic absorption spectrophotometry. + The concentration.

[0126] The measurement results are shown in Table 1.

[0127] Table 1 Li + and Tl + Extraction rate determination results As shown in Table 1, Examples 1 to 3 effectively achieved the selective separation of lithium and thallium in the lepidolite leachate by combining the ultrasonic cavitation-selective complexation synergistic system for lithium mica activation and extraction with a polar extraction system. The final extraction rates of lithium and thallium were greater than 97% and 99%, respectively, which is much higher than the lithium and thallium extraction processes used in Comparative Examples 1 and 2. The purified lithium and thallium can be used as products.

[0128] 2. Determination of toxicity of lepidolite leaching residue 100g of dry lepidolite leaching residue was placed in a 2L capped wide-mouth polyethylene bottle, and 1L of deionized water was added. The bottle was vertically fixed on a horizontal reciprocating shaker, and the frequency was set to (110±10) times / min with an amplitude of 40mm for continuous shaking for 8 hours. After shaking was stopped, the system was allowed to stand for 16 hours to allow solid-liquid separation. The leachate was then filtered through a 0.45μm filter membrane to obtain a clear leachate. Quantitative analysis of heavy metals in the leachate was performed. The results are shown in Table 2.

[0129] Table 2 Heavy Metal Leaching Performance Indicators of Lithium Mica Leaching Residue Analysis of the heavy metal leaching performance data from Examples 1-3 and Comparative Example 3 shows that the heavy metal leaching concentrations in Examples 1-3 are low, all below the limits specified in DB36 / T 1968-2024. Comparative Example 3, using the traditional high-temperature sulfate roasting method, showed that the lead, cadmium, and thallium levels in the lepidolite leaching residue did not meet the general solid waste emission standards. The detection results of thallium content in the toxic leachate from the lepidolite leaching residue obtained in Comparative Example 3 are as follows: Figure 2 As shown, thallium was measured using a standard inductively coupled plasma mass spectrometer (ICP-MS). From... Figure 2 Analysis shows that, compared with the standard (DB36 / T 1968-2024), thallium exceeded the standard by 6 times, and the leaching rate of heavy metal elements such as chromium and manganese was relatively high, which will pose a significant threat to the environment over time.

[0130] The third-party test results of thallium in the toxic leachate of the lithium mica leaching residue obtained in Example 3 are as follows: Figure 3 As shown. From Figure 3 Analysis shows that the thallium content in the lepidolite leaching residue is 0.004 mg / L, which is lower than the standard value (DB36 / T1968-2024) and meets the standard for roadbed filling.

[0131] The above results demonstrate that the combination of ultrasonic cavitation-selective complexation synergistic system for lithium and thallium extraction from lepidolite activated mica, along with a polar extraction system, effectively achieves the selective separation of lithium and thallium in lepidolite leachate. The lithium extraction rate is greater than 97%, and the thallium extraction rate is greater than 99%. Furthermore, the leached lithium residue possesses the chemical safety profile suitable for use as roadbed filler. In summary, the method for extracting lithium and thallium from lepidolite concentrate provided by this invention has the following advantages: First, by grinding the lepidolite concentrate, the specific surface area of ​​the particles is significantly increased, and the interlayer spacing is broken by mechanical compression, allowing metal ions (Li) to escape. + 、Tl + It is easier to contact with external reagents. Then, by combining the chemical chelation activation of the complex acid mixed complex system, the dispersion and anti-agglomeration effect of the surfactant, and the cavitation impact effect of ultrasound, deep activation of lepidolite under mild conditions below 100℃ is achieved. Then, by leaching and filtering the lepidolite concentrate slurry after ultrasonic cracking, a lithium-containing leachate is obtained. Then, extraction is carried out in two steps. First, the complexing agent, diluent and extractant of thallium are mixed to construct a first polar extraction system. Then, the lithium-containing leachate is added to the first polar extraction system for mixing and extraction to obtain a thallium-rich organic phase and raffinate. Thallium is extracted separately to form a thallium-rich organic phase. Then, the thallium-rich organic phase is desorbed to obtain a thallium-rich solution. Secondly, a lithium complexing agent, diluent, and extractant are mixed to construct a second polar extraction system. The raffinate is then added to this system for mixing and dedicated lithium extraction, forming a lithium complex. Finally, the lithium complex is purified and eluted to obtain a lithium-rich solution. This entire process, through a clever design of stepwise activation and extraction, achieves effective separation of lithium and thallium. It recovers both valuable elements while avoiding interference between the two metal ions during extraction, thus improving the purity of the final product. Ultimately, this achieves efficient activation of lepidolite, resulting in efficient extraction and targeted recovery of lithium and thallium.

Claims

1. A method for extracting lithium and thallium from lepidolite concentrate, characterized in that, Includes the following steps: A composite acid complexing system and a surfactant are added to lepidolite concentrate, and the mixture is ground to form a lepidolite concentrate solids slurry. The lepidolite concentrate solids slurry is then subjected to ultrasonic treatment to obtain an ultrasonically broken lepidolite concentrate solids slurry. The solid content slurry of lepidolite concentrate after ultrasonic cracking is heated and leached, then filtered to obtain lithium-containing leachate and lepidolite leaching residue. A thallium complexing agent, a diluent, and an extractant are mixed to construct a first polar extraction system. The lithium-containing leachate is then added to the first polar extraction system for mixing and extraction to obtain a thallium-rich organic phase and a raffinate. The thallium-rich organic phase was desorbed to obtain a thallium-rich solution and a thallium-poor organic phase; A lithium complexing agent, a diluent, and an extractant are mixed to construct a second polar extraction system. The raffinate is then added to the second polar extraction system for mixing and extraction to obtain a lithium complex. The lithium complex was purified and eluted to obtain Li-rich... + Solution.

2. The method according to claim 1, characterized in that, Purification of the lithium complex includes the following steps: transferring the lithium complex into an ion exchange system and purifying it with resin to obtain a lithium-adsorbed resin; eluting the lithium-adsorbed resin to obtain a Li-rich resin. + Solution.

3. The method according to claim 2, characterized in that, The resin filler content is calculated based on bed volume, with 5-10 L of resin corresponding to each cubic meter of lithium complex; the flow rate of the lithium complex in the ion exchange system is 15-50 L / h, and the circulation time is 6-8 hours / cycle; the eluent is 0.1-0.5 mol / L dilute hydrochloric acid, and the elution flow rate is 0.5-0.8 bed volume / hour; the Li-rich... + The pH of the solution is 1.0 to 1.

5.

4. The method according to claim 2, characterized in that, The resin is D201 strong acid sulfonic acid resin, and the resin after lithium adsorption is D201 strong acid sulfonic acid resin after lithium adsorption.

5. The method according to claim 1, characterized in that, The mass ratio of the lithium mica concentrate, the complex acid mixed complex system, and the surfactant is 1000:1000~1500:

5.

6. The method according to claim 1, characterized in that, The complex acid mixture complex system includes an acid and nitric acid (NTA). The acid is selected from one or both of citric acid and succinic acid. The concentration of the acid in the complex acid mixture complex system is 0.15-0.25 mol / L, the concentration of nitric acid (NTA) is 0.05-0.08 mol / L, and the molar ratio of acid to nitric acid is 3:

1. And / or, the mass ratio of the lithium mica concentrate to the composite acid complex system is 1:1.0 to 1.5; And / or, the surfactant is one or both of sodium dodecyl sulfate and sodium dodecyl sulfonate; And / or, based on lepidolite concentrate, the amount of the surfactant is 0.5 wt%; And / or, the complexing agent of the thallium is dibenzo-18-crown-6-ether, and the concentration of the complexing agent of the thallium in the first polar extraction system is 0.8~1.2 g / L; And / or, the lithium complexing agent is ethyl acetoacetate, with a concentration of 1.5~2.0 g / L; And / or, the extractant is a 15-20 vol% solution of tributyl phosphate (TBP); And / or, the diluent is sulfonated kerosene; And / or, the pH value of the first polar extraction system is 4.0~4.5, the saponification rate is 40~65%; the total ratio of aqueous phase to organic phase in the first polar extraction system is 3~8:1, that is, the volume ratio of aqueous phase to organic phase in the first polar extraction system is 3~8:

1. And / or, the desorbent is an aqueous solution of H2SO4, the concentration of H2SO4 in the aqueous solution is 1.0~1.5 mol / L, and the ratio of thallium-rich organic phase to desorbent is 3~5:1; And / or, the pH value of the second polar extraction system is 5.5~6.0, and the saponification rate is 40~65%; the total ratio of aqueous phase to organic phase in the second polar extraction system is 3~8:1, that is, the volume ratio of aqueous phase to organic phase in the second polar extraction system is 3~8:

1.

7. The method according to claim 1, characterized in that, The grinding is carried out in a sand mill, the main shaft speed of the sand mill is 150~200 r / min, the stirring axis speed of the sand mill is 0.5~1.5m / s, the material-to-ball ratio of the ball mill is 10:1, the grinding concentration of the ball mill is 65~70%, and the grinding time is 30~60min. And / or, the conditions for the ultrasonic treatment include: ultrasonic frequency of 20~25kHz, power density of 320~350W / L, treatment temperature controlled between 80~95℃, treatment time of 45~60min, continuous flow treatment, slurry flow rate of 8~12L / min, and cavitation number controlled at 0.8~1.

2. And / or, the heating extraction adopts a three-stage countercurrent water leaching process, the leaching liquid-to-solid ratio during the heating extraction process is 4~5:1, the heating extraction temperature is 90~95℃, and the heating extraction time is 180~240min; the heating extraction is also accompanied by stirring, and the stirring speed is 300~400rpm; the filtration is vacuum filtration. And / or, adding the lithium-containing leachate to the first polar extraction system for mixing and extraction includes: mixing with a mechanical stirrer at a speed of 250~300 rpm for 2~3 min; the extraction is a 3~4 stage countercurrent extraction; And / or, adding the raffinate to the second polar extraction system for mixing and extraction includes: mixing with a mechanical stirrer at a speed of 250~300 rpm for 2~3 min; the extraction is a 3~4 stage countercurrent extraction; And / or, the desorption temperature is 25~35℃, the desorption time is 10~15min, and the desorption stage is 2.

8. The method according to claim 1, characterized in that, The chemical composition and mass percentage of the lepidolite concentrate are as follows: 1.16% Li, 0.003% Tl, 1.39% F, 0.67% Na, 0.06% Mg, 12.11% Al, 20.56% Si, 0.13% P, 0.02% S, 0.02% Cl, 4.97% K, 0.07% Ca, and 0.004% Fe. The composition includes Nb, 0.009% W, 39.47% O, 0.02% Ti, 0.005% Cr, 0.24% Mn, 0.49% Fe, 0.003% Cu, 0.032% Zn, 0.006% Ga, 0.01% As, 0.61% Rb, 0.001% Zr, 0.06% Cs, and 0.009% Pb.

9. The lithium or thallium recovered by the method according to any one of claims 1 to 8, characterized in that, The extraction rate of lithium is greater than 97%, and the extraction rate of thallium is greater than 99%.

10. The application of the method according to any one of claims 1 to 8 in the extraction of lithium and / or thallium from lepidolite concentrate.

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