A method for selectively extracting and recovering thallium from lithium mica smelting slag

Abstract

The present application relates to the technical field of lithium resource recovery and extraction, and particularly relates to a method for selectively extracting and recovering thallium from lepidolite smelting slag. The present application aims to solve the problem that coexisting metals are difficult to be efficiently removed in the traditional thallium recovery process. The technical scheme is as follows: the lepidolite smelting slag is pretreated to release thallium in the form of Tl + , and a pretreated leaching solution is obtained; the pretreated leaching solution is treated by introducing a pulse electric field to convert Tl + into Tl 3+ , and a pulse product containing Tl 3+ is obtained; the pulse product containing Tl 3+ is treated by combining an electrochemical gradient field to promote the directional migration of Tl 3+ , a suspension containing Tl(OH)3 is formed, solid-liquid separation is performed, a clear liquid and a Tl(OH)3 filter cake are obtained; and the electrolyte containing the Tl(OH)3 filter cake is subjected to electrodeposition by adopting a cyclone electrodeposition technology, and metallic thallium is obtained.

Description

Technical Field

[0001] This invention relates to the field of lithium resource recovery and extraction technology, specifically to a method for selectively extracting and recovering thallium from lepidolite smelting slag. Background Technology

[0002] Thallium (Tl) and its compounds possess both significant industrial value and extreme biological toxicity. As a rare dispersed metal, thallium plays an irreplaceable role in high-tech fields such as semiconductors, superconductors, specialty glasses, and medical contrast agents.

[0003] Currently, domestic and international technologies for treating thallium in lepidolite smelting slag mainly focus on traditional processes such as chemical leaching, solidification, and physical adsorption. These technologies have significant limitations in practical applications. Thallium in lepidolite smelting slag typically exists as easily soluble sulfates, often accompanied by large amounts of alkali metals such as potassium and sodium. Due to thallium's extreme toxicity and high permeability, traditional treatment processes are difficult to completely remove it, easily leading to secondary pollution of the waste residue. Although some new processes claim to bring thallium up to the "general Class I industrial solid waste standard," risks remain regarding its long-term stability, and large quantities of chemical reagents are still required. Furthermore, current main technologies (such as simple acid leaching or roasting-leaching methods) generally result in low thallium recovery rates, making high-value utilization difficult. Traditional processes typically involve high-temperature roasting and strong acid leaching, resulting in extremely high energy consumption and making precise control of thallium's specific valence state impossible.

[0004] Currently, some literature mentions the use of differential pulsed anodic stripping voltammetry (DPASV) or multi-pulse electric field technology for the detection of trace thallium. However, these methods are mainly used for analytical detection, and when dealing with complex slag leachates, they require complex pre-separation steps. There is still a lack of methods specifically targeting thallium in specific valence states (such as Tl) in lepidolite smelting slag. + Convert to Tl 3+ The highly selective separation technology makes it difficult to efficiently remove coexisting metals.

[0005] In summary, existing technologies cannot achieve highly selective extraction and high-purity recovery of thallium while reducing energy consumption, and also pose potential environmental safety hazards. Summary of the Invention

[0006] To address the problems existing in the prior art, this invention provides a method for selectively extracting and recovering thallium from lepidolite smelting slag. This method solves the problem that traditional thallium recovery processes cannot precisely control the specific valence state of thallium and lack highly selective separation technology for thallium in specific valence states in lepidolite smelting slag, resulting in the difficulty of efficiently removing coexisting metals. It also solves the problem that traditional thallium recovery processes are difficult to achieve high-purity thallium recovery while reducing energy consumption, and pose potential environmental safety hazards.

[0007] The technical solution of the present invention is as follows: This invention provides a method for selectively extracting and recovering thallium from lepidolite smelting slag, comprising the following steps: S1. Pretreatment of lepidolite smelting slag to obtain thallium using Tl + The pretreated leachate and pretreated leachate residue are released in the form of a solution. S2. Introduce a pulsed electric field to treat the pretreated leachate, so that Tl + Convert to Tl 3+ , to obtain Tl 3+ The pulse products; S3, combined with pulsed electric field, Tl 3+ The concentration difference and the electrochemical gradient field formed by the pH automatic control system affect the concentration of Tl. 3+ The pulse products are treated to promote Tl 3+ Directional migration forms a suspension containing Tl(OH)3. Solid-liquid separation yields a clear liquid and a Tl(OH)3 filter cake. S4. Add an acid solution to the Tl(OH)3 filter cake to obtain an electrolyte. Electrodeposit the electrolyte using a vortex electrodeposition technique to obtain metallic thallium.

[0008] This invention pretreats lepidolite smelting slag to reduce the amount of thallium in the slag to Tl. + The release of thallium in this form effectively activated and initially separated thallium, laying the foundation for subsequent precise and selective thallium extraction. By introducing a pulsed electric field to treat the pretreatment leachate, the potential difference between thallium and coexisting metals was differentiated, allowing for precise thallium extraction. + Oxidized to Tl 3+ This technology solves the problem in traditional thallium recovery processes that lack precise control over the specific valence state of thallium and the highly selective separation technology for thallium in specific valence states in lepidolite smelting slag, resulting in the difficulty of efficiently removing coexisting metals. By establishing an electrochemical gradient field and combining it with an ion-selective membrane, the process promotes the efficient removal of thallium in thallium. 3+Directed migration forms a suspension containing Tl(OH)3, achieving efficient separation of thallium from other components in lepidolite smelting slag. This effectively prevents secondary leaching of thallium during landfilling and reuse of the slag, completely solving the problem of thallium remediation in lepidolite lithium slag. The Tl(OH)3 filter cake obtained in the electrochemical control and directed migration unit is formulated into an electrolyte. Through cyclone electrodeposition technology in the continuous recovery unit, the electrolyte is electrodeposited and recovered, forming a continuous process that achieves a thallium recovery rate of over 90% and a purity of over 99% in lepidolite slag, meeting the industrial requirements for metal purification. The high selectivity of this invention not only extracts thallium but also retains other valuable rare metals (such as rubidium and cesium) in the lithium slag during separation, allowing for extraction in subsequent processes and achieving comprehensive resource utilization benefits. In terms of cost, this invention, through closed-loop process and precise electrochemical control, can significantly reduce production costs, offering advantages such as low energy consumption and low process temperature, meeting the requirements of green processes. This invention pretreats lepidolite smelting slag, so that thallium is added using Tl. + The process involves releasing thallium in a pretreated leachate, which is then extracted through an electrochemical gradient field and electrodeposition. This process achieves the resource recycling of thallium, eliminates the risk of secondary emissions, and solves the problem that traditional thallium recovery processes struggle to achieve high-purity thallium recovery while reducing energy consumption, and also pose potential environmental safety hazards.

[0009] Furthermore, in step S1, the pretreatment includes mechanical activation and acid leaching; Mechanical activation includes: crushing the lepidolite smelting slag to a particle size of less than 10 mm to obtain lepidolite smelting slag with qualified particle size; Anhydrous ethanol was added to the lepidolite smelting slag with qualified particle size, and the mixture was ball-milled to obtain lepidolite smelting slag powder. Acid leaching includes: heating a dilute sulfuric acid aqueous solution, adding the lithium mica smelting slag powder, acid leaching, solid-liquid separation, and obtaining filtrate and filter cake; The filter cake is washed with water or a circulating lean thallium anolyte to obtain a washing solution and pretreated leaching residue. The filtrate and washing liquid are combined to obtain the pretreated leachate.

[0010] By cleverly combining mechanical activation and acid leaching, lepidolite smelting slag is pretreated to enhance its properties. + The form of efficient release significantly improves the leaching rate of thallium.

[0011] Furthermore, the amount of anhydrous ethanol added is 0.5% of the total mass of the lepidolite smelting slag; The ball milling is carried out using a planetary ball mill with a ball milling cylinder speed of 400~500 rpm and a ball milling time of 4~6 hours. At the same time, an intermittent operation mode is activated, which runs for 30~40 minutes and stops for 10 minutes. The solid-liquid ratio of the lithium mica smelting slag powder and the dilute sulfuric acid aqueous solution is 1 kg: 5~8 L; The mass percentage of sulfuric acid in the first dilute sulfuric acid aqueous solution is 2% (w / w); the first dilute sulfuric acid aqueous solution is heated to 60~65℃; The acid leaching time is 2 hours; The washing is performed 1 to 2 times; The pH of the pretreated leachate is 1.

[0012] Furthermore, S2 specifically includes: A dual-chamber electrolyzer is designed, comprising an anode chamber and a cathode chamber, with a cation exchange membrane provided between the anode chamber and the cathode chamber. The anode chamber is connected to an anode chamber storage tank, and the cathode chamber is connected to a cathode chamber storage tank. The pretreated leachate is pumped into the anode chamber storage tank, causing it to circulate between the anode chamber storage tank and the first anode surface of the anode chamber, with a flow rate controlled at 0.2~0.5 m / s. The catholyte is pumped into the cathode chamber storage tank, causing it to circulate between the cathode chamber storage tank and the first cathode surface of the cathode chamber, with a flow rate controlled at 0.2~0.5 m / s. Electrolysis employs a constant potential pulse mode to create a pulsed electric field, which is used to control Tl. + Oxidation to produce Tl 3+ , to obtain Tl 3+ The pulse products.

[0013] By using a dual-chamber electrolytic cell and a pulsed electric field to precisely control the redox potential, the potential difference between thallium ions and other coexisting metals was distinguished.

[0014] Furthermore, the cathode chamber uses materials with a specific surface area >1200m². 2 / g porous graphene / MnO2 composite electrode is used as the first cathode, and a titanium-based DSA electrode is used as the first anode in the anode chamber. The cation exchange membrane is selected from perfluorosulfonic acid cation exchange membrane. The cathode solution is selected from a dilute sulfuric acid aqueous solution with a sulfuric acid concentration of 0.5~1.0 mol / L. Using an Ag / AgCl electrode as the reference electrode, the peak value of the anode potential pulse was set to 1.3~1.5V; the pulse frequency of the anode potential was 10~20Hz, and the duty cycle was 40~60%. The electrolysis temperature is 40~50℃.

[0015] Furthermore, S3 specifically includes: An initial solution containing a buffer is added to the cathode chamber, and the automatic pH control system is activated to control the pH of the mixture of the initial solution containing the buffer and the catholyte in the cathode chamber to be 3.5~4.5. Through pulsed electric field, Tl 3+ The concentration difference and pH automatic control system establishes an electrochemical gradient field between the anode and cathode chambers; Electrochemical gradient field for Tl 3+ The pulse products are treated to promote Tl 3+ Tl migrates directionally from the anode chamber through the cation exchange membrane to the cathode chamber, entering the cathode chamber... 3+ Hydrolysis forms a suspension containing Tl(OH)3; The suspension containing Tl(OH)3 was subjected to sedimentation for 6-8 hours, followed by solid-liquid separation to obtain a clear liquid and a Tl(OH)3 filter cake.

[0016] By introducing pulsed electric field modulation technology into the electrochemical regulation and directional migration unit, a portion of the high-temperature calcination process is replaced, which greatly reduces energy consumption and achieves precise electronic-level control.

[0017] Furthermore, S4 specifically includes: A cyclone electrolytic cell is provided, comprising an outer cylinder and an inner cylinder arranged concentrically, the inner cylinder being sleeved inside the outer cylinder, the outer cylinder being the cell body, and a second anode in the form of a cage or plate being provided on the cell wall of the cell body, the inner cylinder being a rotatable second cathode; Add sulfuric acid aqueous solution to Tl(OH)3 filter cake, stir to dissolve, and obtain electrolyte; The electrolyte is pumped into the vortex electrolytic cell, where it circulates continuously and is heated to a set temperature. The rotation speed of the second cathode is adjusted, the DC power supply is turned on, a constant current density is set, and electrodeposition begins. When Tl in the electrolyte 3+ When the concentration drops to 10~20g / L, electrodeposition is stopped, the second cathode is lifted out, and the spongy thallium metal on the second cathode is peeled off. The spongy thallium metal is washed with water, dried, smelted, and cast into ingots to obtain the product thallium metal.

[0018] In the continuous recovery unit, a cyclone electrolyzer with an inner and outer cylinder is installed, allowing the electrolyte to continuously circulate within the cyclone electrolyzer. Combined with a rotatable cathode, this effectively renews the electrolyte layer at the electrode interface, reduces concentration polarization, and improves Tl. 3+ Electrodeposition efficiency. By precisely controlling the electrodeposition endpoint, at Tl 3+Electrodeposition is terminated when the concentration drops to 10-20 g / L, which effectively avoids the back dissolution of thallium metal and the waste of electrical energy, improving process stability and economy. Combined with post-processing steps such as washing, drying, smelting, and ingot casting, it can effectively remove impurities and purify thallium metal, ensuring high thallium recovery rate and product purity, and is suitable for industrial continuous production.

[0019] Furthermore, the tank body is made of polyvinylidene fluoride, the second anode is selected from titanium-based iridium-tantalum oxide coated DSA anode, and the second cathode is selected from high-purity stainless steel or titanium rod. The sulfuric acid in the aqueous solution has a concentration of 80-120 g / L. The set temperature is 40~50℃; The rotational speed of the second cathode is 800~1500 rpm; The constant current density is 300 A / m 2 .

[0020] Furthermore, the chemical composition of the lepidolite smelting slag, by mass percentage, includes: 53.3% SiO2, 25.3% Al2O3, 0.55% Fe2O3, 3.25% CaO, 0.32% MgO, 0.07% SO3, 7.45% Na2O, 5.12% K2O, 2.24% TiO2, and other impurities. The loss on ignition rate of the lepidolite smelting slag is 0.15%, and the thallium content in the lepidolite smelting slag is 20.42 mg / kg.

[0021] Furthermore, the leaching rate of the metal thallium is 95.3-97.2%, the recovery rate is 91.6-93.2%, and the purity is >99%.

[0022] The beneficial effects of this invention: This invention pretreatment of lepidolite smelting slag reduces the amount of thallium in the slag to Tl. + The release of thallium in this form effectively achieved the activation and initial separation of thallium, laying the foundation for subsequent precise and selective thallium extraction. By designing a dual-chamber electrolytic cell in the electrochemical regulation and directional migration unit and introducing a pulsed electric field, the potential difference between thallium and coexisting metals was distinguished, enabling precise extraction of thallium. + Oxidized to Tl 3+ By establishing an electrochemical gradient field between the anode and cathode chambers of the electrochemical regulation and directional migration unit, and in conjunction with an ion-selective membrane, Tl is promoted. 3+Directed migration forms a suspension containing Tl(OH)3, achieving efficient separation of thallium from other components in lepidolite smelting slag and effectively preventing secondary leaching of thallium during landfilling and reuse. The Tl(OH)3 filter cake obtained in the electrochemical control and directed migration unit is formulated into an electrolyte. Electrodeposition technology in the continuous recovery unit is used to recover the electrolyte, forming a continuous process that achieves a thallium recovery rate of over 90% and a purity of over 99% in lepidolite slag, meeting the industrial requirements for metal purification. The high selectivity of this invention not only extracts thallium but also retains other valuable rare metals (such as rubidium and cesium) in the lithium slag during separation, allowing for subsequent extraction and comprehensive resource utilization. In terms of cost, this invention significantly reduces production costs through a closed-loop process and precise electrochemical control, offering advantages such as low energy consumption and low process temperature, meeting the requirements of green processes. This invention pre-treats lepidolite smelting slag, allowing thallium to be extracted in a Tl... + Thallium is released in a pretreated leachate, which is then extracted through an electrochemical gradient field and electrodeposition. This process achieves the resource recycling of thallium and eliminates the risk of secondary emissions. Attached Figure Description

[0023] Figure 1 A process flow diagram for the selective extraction and high-purity recovery of thallium from lepidolite smelting slag; Figure 2 This is a photograph of thallium metal extracted from lepidolite smelting slag in Example 1. Detailed Implementation

[0024] The present invention will be further described in detail below through embodiments, but in no way is the invention limited.

[0025] The technical principles of the present invention are as follows: 1. Pretreatment Unit: A combination of mechanical activation and acid leaching with a 2% (w / w) dilute sulfuric acid aqueous solution is used to treat thallium with Tl. + Released in the form of leaching, with a leaching rate >95%.

[0026] In most thallium-containing solid wastes (such as flue dust and water-quenched slag), thallium exists in the form of inclusions, isomorphic compounds, or low-solubility compounds. Mechanical activation involves subjecting the material to intense impact, shearing, and grinding using a high-energy ball mill (planetary ball mill). This process not only reduces particle size and increases specific surface area, but more importantly, it introduces numerous lattice defects, dislocations, and amorphous regions within the particles. This injection of mechanical energy significantly enhances the surface activity and chemical reactivity of the particles, disrupts the inclusion structure, and greatly improves the kinetics of the subsequent acid leaching process, resulting in a substantial increase in leaching rate and leaching percentage.

[0027] 2. Electrochemical Regulation and Directed Migration Unit (1) Electrochemical control: A dual-chamber electrolytic cell was designed, and the cathode adopted a porous graphene / MnO2 composite electrode with a special structure (specific surface area >1200m²). 2 / g), the anode is a titanium-based DSA electrode, and Tl is precisely controlled by a pulsed electric field (frequency 5~20Hz, duty cycle 40~60%). + →Tl 3+ The oxidation process enables the differentiation of thallium from coexisting metals by potential difference.

[0028] The basis of electrochemical separation is the difference in standard electrode potentials between different redox couples. In acidic media, the standard electrode potentials of the relevant redox couples are shown in equations (I), (II), (III), (IV), and (V): (I) (II) (III) (IV) (V) The potential value shows that Tl + Oxidized to Tl 3+ The required potential (above +1.25V) is much higher than that of other common metal ions (such as Cu). 2+ Pb 2+ Zn 2+ The oxidation potential of the anode constitutes a large "potential window". By precisely controlling the working potential of the anode above +1.25V using an electrochemical workstation, and under pulsed electric field and enhanced mass transfer conditions, Tl is preferentially achieved. + Selective oxidation.

[0029] The pulsed electric field, through periodic "on-off" or "forward-reverse" voltage / current cycles, allows the ion concentration near the electrode surface to recover during the "off" period (or low potential period) of the pulse, effectively thinning the diffusion layer and enhancing mass transfer. During the high-potential pulse period, the reaction proceeds with a high instantaneous current density, while the system relaxes during the low-potential period. This allows for more efficient use of electrical energy for the target reaction, reducing side reactions (such as oxygen evolution) caused by mass transfer lag. Therefore, pulsed current ensures a more uniform and controllable oxidation process.

[0030] (2) Directional migration: utilizing the introduced pulsed electric field, Tl 3+ An automatic concentration difference and pH control system establishes an electrochemical gradient field between the anode and cathode chambers, promoting the directional migration of thallium ions and achieving a separation efficiency >99%. The electrochemical gradient field consists of a potential gradient constructed from a pulsed electric field and a Tl...3+ The concentration gradient formed by the large concentration difference across the membrane and the transmembrane pH gradient regulated by the automatic pH control system are superimposed to achieve Tl. This is achieved through the transmembrane structure of the dual-chamber electrolyzer. 3+ Highly efficient directional separation.

[0031] This migration process is essentially an electrodialysis process. In a two-chamber electrolyzer, the anode and cathode chambers are separated by a perfluorosulfonic acid (PFSA) cation exchange membrane. When an electric field is applied, positively charged cations migrate towards the cathode, while negatively charged anions migrate towards the anode. By inserting a cation exchange membrane at this point, only cations (such as Tl) are allowed to migrate. 3+ H + Cu 2+ Zn 2+ (etc.) pass through, while blocking anions (such as SO4) 2- ).

[0032] The proposed solution involves controlling the pH of the cathode chamber between 3.5 and 4.5. This is based on the principle of utilizing Tl... 3+ Unique hydrolysis properties. Tl 3+ It exists stably as a hydrated ion at low pH values ​​(pH<2), but when the pH increases, it undergoes strong hydrolysis and eventually precipitates as thallium hydroxide (Tl(OH)3) or thallium oxide (Tl2O3), as shown in formula (VI): (VI) The initial pH for hydrolysis precipitation is approximately 2.5–3.0. Therefore, when the Tl in the anode chamber… 3+ Driven by an electric field, it migrates through the cation membrane and enters the cathode chamber (which can be the cathode chamber or a separate intermediate chamber) with a pH of 3.5 to 4.5, where it rapidly hydrolyzes and precipitates as a solid.

[0033] This migration-precipitation mechanism will Tl 3+ It is sealed in the cathode chamber, which greatly reduces the Tl in the cathode chamber. 3+ The free ion concentration thus maintains a large Tl on both sides of the membrane. 3+ The concentration gradient, in turn, greatly enhanced Tl. 3+ The migration driving force. Other coexisting ions such as Cu... 2+ Zn 2+ The pH of the hydroxide precipitates of these ions is much higher than 4.5, so even if they migrate across the membrane, they remain in solution in ionic form. Through solid-liquid separation, the Tl(OH)3 precipitate can be completely separated from other ions.

[0034] 3. Continuous recovery unit: Employs cyclone electrodeposition technology at 300 A / m2 High-purity (>99%) thallium electrodeposition is achieved at high current densities. The entire system employs a closed-loop process, reducing energy consumption by 40% compared to traditional methods, achieving a processing capacity of 50 kg / h, and a thallium recovery rate of >90%.

[0035] First, the Tl(OH)3 precipitate needs to be redissolved with sulfuric acid to obtain a Tl-rich solution. 3+ A pure electrolyte. Then, in the electrolytic cell, Tl 3+ It gains electrons on the cathode surface and is reduced to metallic thallium. The cathode reaction is shown in formula (VII), and the anode reaction is shown in formula (VIII).

[0036] Tl 3+ + 3e - → Tl(s) E 0 = +0.72 V (VII) 2H₂O → O₂↑ + 4H + + 4e - (VIII) The use of vortex electrodeposition technology significantly reduces the thickness of the diffusion layer on the cathode surface through forced convection (high-speed rotating cathode or high-speed flowing electrolyte), resulting in Tl 3+ It can rapidly replenish the reaction interface. The high-speed fluid flow has a shearing and leveling effect on the growth of crystal nuclei, which is beneficial for the formation of a dense, smooth, and high-purity metal deposition layer. Furthermore, enhanced mass transfer ensures the Tl content on the cathode surface. 3+ Maintaining a high concentration level suppressed the occurrence of side reactions such as hydrogen evolution caused by the depletion of the main ion.

[0037] The chemical composition of the lepidolite smelting slag used in the following examples and comparative examples, by mass percentage, includes: 53.3% SiO2, 25.3% Al2O3, 0.55% Fe2O3, 3.25% CaO, 0.32% MgO, 0.07% SO3, 7.45% Na2O, 5.12% K2O, 2.24% TiO2, and other impurities. The loss on ignition rate of the lepidolite smelting slag is 0.15%, and the thallium content in the lepidolite smelting slag is 20.42 mg / kg. Other impurities include Cu, Pb, and Zn.

[0038] The raw materials for the titanium-based iridium-tantalum oxide coated DSA anode used in the following examples include a titanium substrate, coating active component raw materials, and a solvent. Among them, the titanium substrate is usually a TA1 or TA2 grade pure titanium plate or titanium mesh, used for conduction and supporting the substrate; the coating active component raw materials include: (1) hexachloroiridium tetrahydrate (H2IrCl6·6H2O), a precursor of iridium, which forms IrO2 (electrochemically active material) after thermal decomposition; (2) tantalum ethoxide, a precursor of tantalum, which forms Ta2O5 (stabilizer) after thermal decomposition; the molar ratio of IrO2 to Ta2O5 is 7:3; the solvent is an organic solvent such as isopropanol or anhydrous ethanol, used to prepare the coating solution.

[0039] In the examples below, the main component (Fe-based alloy) of the high-purity stainless steel used has a mass fraction of not less than 99%.

[0040] Example 1

[0041] like Figure 1 As shown, a method for highly selectively extracting and recovering thallium with high purity from lepidolite smelting slag includes the following steps: S1, Pretreatment Unit: Mechanical activation and acid leaching of lepidolite smelting slag; S11. Mechanical Activation: 1 kg of lepidolite smelting slag is fed into the silo via a conveyor belt. The jaw crusher is started to crush the lepidolite smelting slag. The crushed lepidolite smelting slag is then passed through a vibrating screen. Lepidolite smelting slag with a particle size less than 10 mm is sent to the intermediate silo for ball milling. Lepidolite smelting slag with an oversized particle size is returned to the jaw crusher for further crushing until the particle size is acceptable. The acceptable lepidolite smelting slag is then fed into a planetary ball mill. Based on the total mass of the lepidolite smelting slag, 0.5 wt% anhydrous ethanol is added to the acceptable lepidolite smelting slag, and ball milling is performed. The ball mill cylinder speed is set to 400 rpm, and the ball milling time is 4 hours. An intermittent operation mode is simultaneously activated, with a 30-minute run followed by a 10-minute stop. After ball milling, lepidolite smelting slag powder is obtained. S12, Acid Leaching: A dilute sulfuric acid aqueous solution with a sulfuric acid mass percentage of 2% (w / w) is prepared using industrial pure water and 98% concentrated sulfuric acid as raw materials.

[0042] In the leaching reactor, a 2% (w / w) dilute sulfuric acid aqueous solution was added. The jacket heating system was turned on to heat the dilute sulfuric acid aqueous solution to 60°C. The stirrer was started, and the lepidolite smelting slag powder was slowly and evenly added to the leaching reactor while stirring. Acid leaching was carried out for 2 hours. After acid leaching, the leached slurry was obtained. The solid-liquid ratio (mass / volume) of the lepidolite smelting slag powder and the dilute sulfuric acid aqueous solution was 1 kg: 6 L, and the stirrer speed was 160 rpm.

[0043] The leached slurry is pumped into a chamber filter press for solid-liquid separation to obtain filtrate and filter cake. The filter cake is washed twice with warm water at 35°C and with a mass equal to that of the filter cake to obtain washing liquid and washed filter cake (pretreatment leaching residue). The filtrate and washing liquid are combined to obtain pretreatment leachate, which is sent to the subsequent electrochemical unit. The washed filter cake (pretreatment leaching residue) is unloaded and sent to the hazardous waste treatment area for harmless treatment.

[0044] S2, Electrochemical Regulation and Directed Migration Unit S21. Electrochemical Control: A dual-chamber electrolytic cell was designed, and a pulsed electric field was introduced to distinguish the potential difference between thallium and coexisting metals. The specific steps are as follows: A dual-chamber electrolyzer is designed, comprising an anode chamber and a cathode chamber, with a perfluorosulfonic acid (PFSA) cation exchange membrane between them. The anode chamber is connected to an anode storage tank, and the cathode chamber is connected to a cathode storage tank. A porous graphene / MnO2 composite electrode (specific surface area >1200 m²) is used in the cathode chamber. 2 / g) is used as the cathode, and a titanium-based DSA electrode is used as the anode in the anode chamber. When a pulsed electric field is applied, positively charged cations move towards the cathode, while negatively charged anions move towards the anode. Only cations (such as Tl) are allowed to pass through the perfluorosulfonic acid (PFSA) cation exchange membrane between the anode and cathode chambers. 3+ H + Cu 2+ Zn 2+ (etc.) pass through, while blocking anions (such as SO4) 2- ).

[0045] A dilute sulfuric acid aqueous solution with a sulfuric acid concentration of 0.5 mol / L was prepared using industrial pure water and 98% concentrated sulfuric acid as raw materials.

[0046] The pretreated leachate from S1 is pumped into the anode chamber storage tank. At this point, the pretreated leachate is the anolyte, with a pH of 1. A pre-prepared 0.5 mol / L dilute sulfuric acid solution is pumped into the cathode chamber storage tank. This 0.5 mol / L dilute sulfuric acid solution is the catholyte. The circulation pump is started, allowing the anolyte to circulate between the anode chamber storage tank and the anode surface of the anode chamber at a velocity controlled at 0.3 m / s. The catholyte circulates between the cathode chamber storage tank and the cathode surface of the cathode chamber at a velocity controlled at 0.3 m / s. After the anolyte and catholyte have stabilized in their respective storage tanks and on their respective electrode surfaces, the electrolysis power supply is connected. Electrolysis is performed in a constant potential pulse mode, forming a pulsed electric field. The pulse peak value of the anode potential (relative to the Ag / AgCl reference electrode) is set at 1.3 V, the pulse frequency is 10 Hz, and the duty cycle is 40%. Tl is precisely controlled by the pulsed electric field. + →Tl3+ The oxidation process enables the differentiation of thallium from coexisting metals by potential difference.

[0047] S22, Directional Migration: Utilizing the introduced pulsed electric field, Tl 3+ The concentration difference and pH automatic control system establishes an electrochemical gradient field between the anode and cathode chambers to promote the directional migration of thallium ions. Through solid-liquid separation, the Tl(OH)3 precipitate is separated from other ions. The specific steps are as follows: An initial solution containing acetate-sodium acetate buffer is added to the cathode chamber, and an automatic pH control system is activated to maintain the pH of the mixture of the initial solution containing acetate-sodium acetate buffer and the catholyte in the cathode chamber at 3.5, thus creating a transmembrane pH gradient between the cathode and anode chambers. During electrochemical regulation in S21, the potential gradient constructed by the pulsed electric field and Tl between the cathode and anode chambers... 3+ The concentration gradient formed by the concentration difference across the membrane and the transmembrane pH gradient regulated by the automatic pH control system together form an electrochemical gradient field, resulting in Tl in the anode chamber. 3+ Under the influence of an electrochemical gradient field, Tl migrates through the perfluorosulfonic acid (PFSA) cation exchange membrane to the cathode chamber, and enters the cathode chamber. 3+ Rapid hydrolysis forms a suspension containing Tl(OH)3. The suspension in the cathode chamber is overflowed and sent to a settling tank for 7 hours for solid-liquid separation, yielding a clear liquid and a Tl(OH)3 filter cake. The Tl(OH)3 filter cake is sent to a subsequent recovery unit, while the clear liquid is returned to the cathode chamber.

[0048] As electrolysis and ion migration continue, the thallium component of the anolyte in the anode chamber of S21 is continuously converted into Tl. 3+ The thallium migrates across the membrane, and the anolyte gradually transforms into a thallium-lean anolyte. When the concentration of thallium in the thallium-lean anolyte decreases to 0.05 mol / L, the thallium-lean anolyte is recycled back to the pretreatment unit for washing the filter cake.

[0049] Cathode Chamber: In the cathode chamber, the cathode is a porous graphene / MnO2 composite electrode. Under the action of a pulsed electric field, the main side reaction of hydrogen evolution occurs on the cathode surface: 2H+ + + 2e - → H2↑, the generated hydrogen gas is discharged through the exhaust port above the cathode chamber.

[0050] S3, Continuous Recovery Unit: Employs vortex electrodeposition technology to achieve high-purity electrodeposition of thallium; A cyclone electrolytic cell is provided, comprising an outer cylinder and an inner cylinder arranged concentrically, with the inner cylinder nested inside the outer cylinder. The outer cylinder forms the cell body, and cage-shaped or plate-shaped anodes are provided on the cell walls. The inner cylinder is a rotatable cathode. An annular gap is formed between the outer and inner cylinders. External pipelines, a circulation storage tank, and a delivery pump are provided outside the cyclone electrolytic cell. A cathode rotation motor is installed below the cyclone electrolytic cell and is directly connected to the inner cylinder via a drive shaft to drive its rotation.

[0051] A vertical concentric cylindrical cyclone electrolytic cell is used for cyclone electrodeposition. The cell body is made of polyvinylidene fluoride (PVDF). The inner wall of the cell body is equipped with an anode, which is a titanium-based iridium-tantalum oxide coated DSA anode (titanium-based IrO2-Ta2O5 coated DSA anode). The center of the cyclone electrolytic cell body is equipped with a high-speed rotating inner cylinder that serves as the cathode. The cathode is made of stainless steel.

[0052] Add the Tl(OH)3 filter cake to a dissolving vessel, and slowly add 0.5L of sulfuric acid aqueous solution with a concentration of 80g / L. Stir until the Tl(OH)3 filter cake is completely dissolved to prepare the electrolyte. Pump the electrolyte into a cyclone electrolytic cell, where it continuously circulates. Heat the electrolyte to 40°C, start the cathode rotation motor, adjust the stainless steel cathode speed to 1000rpm, turn on the DC power supply, and set a constant current density of 300A / m. 2 Electrodeposition begins. The continuous circulation of the electrolyte within the cyclone electrolytic cell refers to the following: the electrolyte enters from the bottom of the cyclone electrolytic cell, flows upward along the annular gap between the outer and inner cylinders, exits from the top, and returns to the bottom of the cyclone electrolytic cell via external pipelines, a circulation tank, and a transfer pump, forming a closed loop. The upward flow of the electrolyte within the annular gap, combined with the rotatable cathode, creates a swirling effect, enhancing mass transfer on the electrode surface.

[0053] Periodic analysis (every 30 minutes) of Tl in the electrolyte 3+ Concentration, when Tl 3+ When the concentration drops to 18 g / L, electrodeposition is stopped, the stainless steel cathode is removed, and the spongy thallium deposited on the cathode is peeled off with a scraper. The collected spongy thallium is washed with water, vacuum dried, and then sent to a medium-frequency induction furnace for melting and ingot casting to obtain the product thallium. The discharged lean thallium electrolyte is returned to the electrochemical regulation and directional migration unit. A physical image of the product thallium is shown below. Figure 2 As shown.

[0054] Example 2

[0055] like Figure 1 As shown, a method for highly selectively extracting and recovering thallium with high purity from lepidolite smelting slag includes the following steps: S1, Pretreatment Unit: Mechanical activation and acid leaching of lepidolite smelting slag; S11. Mechanical Activation: 1 kg of lepidolite smelting slag is fed into the silo via a conveyor belt. The jaw crusher is started to crush the lepidolite smelting slag. The crushed lepidolite smelting slag is then passed through a vibrating screen. Lepidolite smelting slag with a particle size less than 10 mm is fed into an intermediate silo for ball milling. Lepidolite smelting slag with an oversized particle size is returned to the jaw crusher for further crushing until the particle size is acceptable. The acceptable lepidolite smelting slag is then fed into a planetary ball mill. Based on the total mass of the lepidolite smelting slag, 0.5 wt% anhydrous ethanol is added to the acceptable lepidolite smelting slag, and ball milling is performed. The ball mill cylinder speed is set to 450 rpm, and the ball milling time is 5 hours. An intermittent operation mode is simultaneously activated, with a 35-minute run followed by a 10-minute stop. After ball milling, lepidolite smelting slag powder is obtained. S12, Acid Leaching: A dilute sulfuric acid aqueous solution with a sulfuric acid mass percentage of 2% (w / w) is prepared using industrial pure water and 98% concentrated sulfuric acid as raw materials.

[0056] In the leaching reactor, a 2% (w / w) dilute sulfuric acid aqueous solution was added. The jacket heating system was turned on, and the dilute sulfuric acid aqueous solution was heated to 62°C. The stirrer was started, and the lepidolite smelting slag powder was slowly and evenly added to the leaching reactor while stirring. Acid leaching was carried out for 2 hours. After acid leaching, the leached slurry was obtained. The solid-liquid ratio (mass / volume) of the lepidolite smelting slag powder and the dilute sulfuric acid aqueous solution was 1 kg: 7 L, and the stirrer speed was 180 rpm.

[0057] The leached slurry is pumped into a chamber filter press for solid-liquid separation, yielding filtrate and filter cake. The filter cake is washed twice with warm water at 40°C and twice its mass to obtain washing liquid and washed filter cake (pretreatment leaching residue). The filtrate and washing liquid are combined to obtain pretreatment leachate, which is then sent to the subsequent electrochemical unit. The washed filter cake (pretreatment leaching residue) is unloaded and sent to the hazardous waste treatment area for harmless treatment.

[0058] S2, Electrochemical Regulation and Directed Migration Unit S21. Electrochemical Control: A dual-chamber electrolytic cell was designed, and a pulsed electric field was introduced to distinguish the potential difference between thallium and coexisting metals. The specific steps are as follows: A dual-chamber electrolyzer is designed, comprising an anode chamber and a cathode chamber, with a perfluorosulfonic acid (PFSA) cation exchange membrane between them. The anode chamber is connected to an anode storage tank, and the cathode chamber is connected to a cathode storage tank. A porous graphene / MnO2 composite electrode (specific surface area >1200 m²) is used in the cathode chamber. 2 / g) is used as the cathode, and a titanium-based DSA electrode is used as the anode in the anode chamber. When a pulsed electric field is applied, positively charged cations move towards the cathode, while negatively charged anions move towards the anode. Only cations (such as Tl) are allowed to pass through the perfluorosulfonic acid (PFSA) cation exchange membrane between the anode and cathode chambers. 3+ H + Cu 2+ Zn 2+ (etc.) pass through, while blocking anions (such as SO4) 2- ).

[0059] A dilute sulfuric acid aqueous solution with a sulfuric acid concentration of 0.8 mol / L was prepared using industrial pure water and 98% concentrated sulfuric acid as raw materials.

[0060] The pretreated leachate from S1 is pumped into the anode chamber storage tank. At this point, the pretreated leachate is the anolyte, with a pH of 1. A pre-prepared 0.8 mol / L dilute sulfuric acid solution is pumped into the cathode chamber storage tank. This 0.8 mol / L dilute sulfuric acid solution is the catholyte. The circulation pump is started, allowing the anolyte to circulate between the anode chamber storage tank and the anode surface of the anode chamber at a velocity controlled at 0.4 m / s. The catholyte circulates between the cathode chamber storage tank and the cathode surface of the cathode chamber at a velocity controlled at 0.4 m / s. After the anolyte and catholyte have stabilized in their respective storage tanks and on their respective electrode surfaces, the electrolysis power supply is connected. Electrolysis uses a constant potential pulse mode to create a pulsed electric field. The pulse peak value of the anode potential (relative to the Ag / AgCl reference electrode) is set at 1.4 V, the pulse frequency is 15 Hz, and the duty cycle is 50%. Tl is precisely controlled by the pulsed electric field. + →Tl 3+ The oxidation process enables the differentiation of thallium from coexisting metals by potential difference.

[0061] S22, Directional Migration: Utilizing the introduced pulsed electric field, Tl 3+ The concentration difference and pH automatic control system establishes an electrochemical gradient field between the anode and cathode chambers to promote the directional migration of thallium ions. Through solid-liquid separation, the Tl(OH)3 precipitate is separated from other ions. The specific steps are as follows: An initial solution containing acetate-sodium acetate buffer is added to the cathode chamber of the directional migration unit, and the automatic pH control system is activated to maintain the pH of the mixture of the initial solution containing acetate-sodium acetate buffer and the catholyte in the cathode chamber at 4.0, thereby creating a transmembrane pH gradient between the cathode and anode chambers. During electrochemical regulation in S21, the potential gradient constructed by the pulsed electric field and Tl between the cathode and anode chambers... 3+The concentration gradient formed by the huge concentration difference across the membrane and the transmembrane pH gradient regulated by the automatic pH control system together form an electrochemical gradient field, with Tl in the anode chamber... 3+ Under the influence of an electrochemical gradient field, Tl migrates through the perfluorosulfonic acid (PFSA) cation exchange membrane to the cathode chamber, and enters the cathode chamber. 3+ Rapid hydrolysis forms a suspension containing Tl(OH)3. The suspension in the cathode chamber is overflowed and sent to a settling tank for 8 hours for solid-liquid separation, yielding a clear liquid and a Tl(OH)3 filter cake. The Tl(OH)3 filter cake is sent to a subsequent recovery unit, while the clear liquid is returned to the cathode chamber.

[0062] As electrolysis and ion migration continue, the thallium component of the anolyte in the anode chamber of S21 is continuously converted into Tl. 3+ The thallium migrates across the membrane, and the anolyte gradually transforms into a thallium-lean anolyte. When the concentration of thallium in the thallium-lean anolyte decreases to 0.05 mol / L, the thallium-lean anolyte is recycled back to the pretreatment unit for washing the filter cake.

[0063] Cathode Chamber: In the cathode chamber, the cathode is a porous graphene / MnO2 composite electrode. Under the action of a pulsed electric field, the main side reaction of hydrogen evolution occurs on the cathode surface: 2H+ + + 2e - → H2↑, the generated hydrogen gas is discharged through the exhaust port above the cathode chamber.

[0064] S3, Continuous Recovery Unit: Employs vortex electrodeposition technology to achieve high-purity electrodeposition of thallium; A cyclone electrolytic cell is provided, comprising an outer cylinder and an inner cylinder arranged concentrically, with the inner cylinder nested inside the outer cylinder. The outer cylinder forms the cell body, and cage-shaped or plate-shaped anodes are provided on the cell walls. The inner cylinder is a rotatable cathode. An annular gap is formed between the outer and inner cylinders. External pipelines, a circulation storage tank, and a delivery pump are provided outside the cyclone electrolytic cell. A cathode rotation motor is installed below the cyclone electrolytic cell and is directly connected to the inner cylinder via a drive shaft to drive its rotation.

[0065] A vertical concentric cylindrical cyclone electrolytic cell is used for cyclone electrodeposition. The cell body is made of polyvinylidene fluoride (PVDF). The inner wall of the cell body is equipped with an anode, which is a titanium-based iridium-tantalum oxide coated DSA anode (titanium-based IrO2-Ta2O5 coated DSA anode). The center of the cyclone electrolytic cell body is equipped with a high-speed rotating inner cylinder that serves as the cathode. The cathode is made of stainless steel.

[0066] Add the Tl(OH)3 filter cake to a dissolving vessel, and slowly add 0.6L of a 100g / L sulfuric acid aqueous solution. Stir until the Tl(OH)3 filter cake is completely dissolved to prepare the electrolyte. Pump the electrolyte into a cyclone electrolytic cell, where it continuously circulates. Heat the electrolyte to 45°C, start the cathode rotation motor, adjust the stainless steel cathode speed to 1200rpm, turn on the DC power supply, and set a constant current density of 300A / m. 2 Electrodeposition begins. The continuous circulation of the electrolyte within the cyclone electrolytic cell refers to the following: the electrolyte enters from the bottom of the cyclone electrolytic cell, flows upward along the annular gap between the outer and inner cylinders, exits from the top, and returns to the bottom of the cyclone electrolytic cell via external pipelines, a circulation tank, and a transfer pump, forming a closed loop. The upward flow of the electrolyte within the annular gap, combined with the rotatable cathode, creates a swirling effect, enhancing mass transfer on the electrode surface.

[0067] Periodic analysis (every 30 minutes) of Tl in the electrolyte 3+ Concentration, when Tl 3+ When the concentration drops to 15 g / L, electrodeposition is stopped, the stainless steel cathode is lifted out, and the spongy thallium metal deposited on the stainless steel cathode is peeled off with a scraper. The collected spongy thallium metal is washed with water and vacuum dried, and then sent to a medium-frequency induction furnace for melting and casting to obtain the product thallium metal. The discharged lean thallium electrolyte is returned to the electrochemical regulation and directional migration unit.

[0068] Example 3

[0069] like Figure 1 As shown, a method for highly selectively extracting and recovering thallium with high purity from lepidolite smelting slag includes the following steps: S1, Pretreatment Unit: Mechanical activation and acid leaching of lepidolite smelting slag; S11. Mechanical Activation: 1 kg of lepidolite smelting slag is fed into the silo via a conveyor belt. The jaw crusher is started to crush the lepidolite smelting slag. The crushed lepidolite smelting slag is then passed through a vibrating screen. Lepidolite smelting slag with a particle size less than 10 mm is sent to the intermediate silo for ball milling. Lepidolite smelting slag with an oversized particle size is returned to the jaw crusher for further crushing until the particle size is acceptable. The acceptable lepidolite smelting slag is then fed into a planetary ball mill. Based on the total mass of the lepidolite smelting slag, 0.5 wt% anhydrous ethanol is added to the acceptable lepidolite smelting slag, and ball milling is performed. The ball mill cylinder speed is set to 500 rpm, and the ball milling time is 6 hours. An intermittent operation mode is simultaneously activated, with a 40-minute run followed by a 10-minute stop. After ball milling, lepidolite smelting slag powder is obtained. S12, Acid Leaching: A dilute sulfuric acid aqueous solution with a sulfuric acid mass percentage of 2% (w / w) is prepared using industrial pure water and 98% concentrated sulfuric acid as raw materials.

[0070] In the leaching reactor, a 2% (w / w) dilute sulfuric acid aqueous solution was added. The jacket heating system was turned on, and the dilute sulfuric acid aqueous solution was heated to 65°C. The stirrer was started, and the lepidolite smelting slag powder was slowly and evenly added to the leaching reactor while stirring. Acid leaching was carried out for 2 hours. After acid leaching, the leached slurry was obtained. The solid-liquid ratio (mass / volume) of the lepidolite smelting slag powder and the dilute sulfuric acid aqueous solution was 1 kg: 8 L, and the stirrer speed was 200 rpm.

[0071] The leached slurry is pumped into a chamber filter press for solid-liquid separation to obtain filtrate and filter cake. The filter cake is washed twice with warm water at 45°C and twice the mass of the filter cake to obtain washing liquid and washed filter cake (pretreatment leaching residue). The filtrate and washing liquid are combined to obtain pretreatment leachate, which is sent to the subsequent electrochemical unit. The washed filter cake (pretreatment leaching residue) is unloaded and sent to the hazardous waste treatment area for harmless treatment.

[0072] S2, Electrochemical Regulation and Directed Migration Unit S21. Electrochemical Control: A dual-chamber electrolytic cell was designed, and a pulsed electric field was introduced to distinguish the potential difference between thallium and coexisting metals. The specific steps are as follows: A dual-chamber electrolyzer is designed, comprising an anode chamber and a cathode chamber, with a perfluorosulfonic acid (PFSA) cation exchange membrane between them. The anode chamber is connected to an anode storage tank, and the cathode chamber is connected to a cathode storage tank. A porous graphene / MnO2 composite electrode (specific surface area >1200 m²) is used in the cathode chamber. 2 / g) is used as the cathode, and a titanium-based DSA electrode is used as the anode in the anode chamber. When a pulsed electric field is applied, positively charged cations move towards the cathode, while negatively charged anions move towards the anode. Only cations (such as Tl) are allowed to pass through the perfluorosulfonic acid (PFSA) cation exchange membrane between the anode and cathode chambers. 3+ H + Cu 2+ Zn 2+ (etc.) pass through, while blocking anions (such as SO4) 2- ).

[0073] A dilute sulfuric acid aqueous solution with a sulfuric acid concentration of 1.0 mol / L was prepared using industrial pure water and 98% concentrated sulfuric acid as raw materials.

[0074] The pretreated leachate from S1 is pumped into the anode chamber storage tank. At this point, the pretreated leachate is the anolyte, with a pH of 1. A pre-prepared 1.0 mol / L dilute sulfuric acid solution is pumped into the cathode chamber storage tank. This 1.0 mol / L dilute sulfuric acid solution is the catholyte. The circulation pump is started, allowing the anolyte to circulate between the anode chamber storage tank and the anode surface of the anode chamber at a velocity controlled at 0.5 m / s. The catholyte circulates between the cathode chamber storage tank and the cathode surface of the cathode chamber at a velocity controlled at 0.5 m / s. After the anolyte and catholyte have stabilized in their respective storage tanks and on their respective electrode surfaces, the electrolysis power supply is connected. Electrolysis is performed in a constant potential pulse mode, forming a pulsed electric field. The pulse peak value of the anode potential (relative to the Ag / AgCl reference electrode) is set at 1.5 V, the pulse frequency is 20 Hz, and the duty cycle is 60%. Tl is precisely controlled through the pulsed electric field. + →Tl 3+ The oxidation process enables the differentiation of thallium from coexisting metals by potential difference.

[0075] S22, Directional Migration: Utilizing the introduced pulsed electric field, Tl 3+ The concentration difference and pH automatic control system establishes an electrochemical gradient field between the anode and cathode chambers to promote the directional migration of thallium ions. Through solid-liquid separation, the Tl(OH)3 precipitate is separated from other ions. The specific steps are as follows: An initial solution containing acetate-sodium acetate buffer is added to the cathode chamber, and the automatic pH control system is activated to maintain the pH of the mixture of the initial solution containing acetate-sodium acetate buffer and the catholyte in the cathode chamber at 4.5, thus creating a transmembrane pH gradient between the cathode and anode chambers. During electrochemical regulation in S21, the potential gradient constructed by the pulsed electric field and Tl between the cathode and anode chambers... 3+ The concentration gradient formed by the huge concentration difference across the membrane and the transmembrane pH gradient regulated by the automatic pH control system together form an electrochemical gradient field, with Tl in the anode chamber... 3+ Under the influence of an electrochemical gradient field, Tl migrates through the perfluorosulfonic acid (PFSA) cation exchange membrane to the cathode chamber, and enters the cathode chamber. 3+ Rapid hydrolysis forms a suspension containing Tl(OH)3. The suspension in the cathode chamber is then overflowed into a settling tank for 8 hours to separate the solid and liquid, yielding a clear liquid and a Tl(OH)3 filter cake. The Tl(OH)3 filter cake is sent to a subsequent recovery unit, while the clear liquid is returned to the cathode chamber.

[0076] As electrolysis and ion migration continue, the thallium component of the anolyte in the anode chamber of S21 is continuously converted into Tl. 3+The thallium migrates across the membrane, and the anolyte gradually transforms into a thallium-lean anolyte. When the concentration of thallium in the thallium-lean anolyte decreases to 0.05 mol / L, the thallium-lean anolyte is recycled back to the pretreatment unit for washing the filter cake.

[0077] Cathode Chamber: In the cathode chamber, the cathode is a porous graphene / MnO2 composite electrode. Under the action of a pulsed electric field, the main side reaction of hydrogen evolution occurs on the cathode surface: 2H+ + + 2e - → H2↑, the generated hydrogen gas is discharged through the exhaust port above the cathode chamber.

[0078] S3, Continuous Recovery Unit: Employs vortex electrodeposition technology to achieve high-purity electrodeposition of thallium; A cyclone electrolytic cell is provided, comprising an outer cylinder and an inner cylinder arranged concentrically, with the inner cylinder nested inside the outer cylinder. The outer cylinder forms the cell body, and cage-shaped or plate-shaped anodes are provided on the cell walls. The inner cylinder is a rotatable cathode. An annular gap is formed between the outer and inner cylinders. External pipelines, a circulation storage tank, and a delivery pump are provided outside the cyclone electrolytic cell. A cathode rotation motor is installed below the cyclone electrolytic cell and is directly connected to the inner cylinder via a drive shaft to drive its rotation.

[0079] A vertical concentric cylindrical cyclone electrolytic cell is used for cyclone electrodeposition. The cell body is made of polyvinylidene fluoride (PVDF). The inner wall of the cell body is equipped with an anode, which is a titanium-based iridium-tantalum oxide coated DSA anode (titanium-based IrO2-Ta2O5 coated DSA anode). The center of the cyclone electrolytic cell body is equipped with a high-speed rotating inner cylinder that serves as the cathode. The cathode is made of stainless steel.

[0080] Add the Tl(OH)3 filter cake to a dissolving vessel, and slowly add 0.7L of a sulfuric acid aqueous solution with a concentration of 120g / L. Stir until the Tl(OH)3 filter cake is completely dissolved to prepare the electrolyte. Pump the electrolyte into a cyclone electrolytic cell, where it continuously circulates. Heat the electrolyte to 50°C, start the cathode rotation motor, adjust the stainless steel cathode speed to 1500rpm, turn on the DC power supply, and set a constant current density of 300A / m. 2 Electrodeposition begins. The continuous circulation of the electrolyte within the cyclone electrolytic cell refers to the following: the electrolyte enters from the bottom of the cyclone electrolytic cell, flows upward along the annular gap between the outer and inner cylinders, exits from the top, and returns to the bottom of the cyclone electrolytic cell via external pipelines, a circulation tank, and a transfer pump, forming a closed loop. The upward flow of the electrolyte within the annular gap, combined with the rotatable cathode, creates a swirling effect, enhancing mass transfer on the electrode surface.

[0081] Periodic analysis (every 30 minutes) of Tl in the electrolyte 3+ Concentration, when Tl3+ When the concentration drops to 10 g / L, electrodeposition is stopped, the stainless steel cathode is lifted out, and the spongy thallium metal deposited on the stainless steel cathode is peeled off with a scraper. The collected spongy thallium metal is washed with water and vacuum dried, and then sent to a medium-frequency induction furnace for melting and casting to obtain the product thallium metal. The discharged lean thallium electrolyte is returned to the electrochemical regulation and directional migration unit.

[0082] Comparative Example 1 A method for recovering thallium from lepidolite smelting slag includes the following steps: S1, Pretreatment Unit: Mechanical activation and acid leaching of lepidolite smelting slag; S11. Mechanical Activation: 1 kg of lepidolite smelting slag is fed into the silo via a conveyor belt. The jaw crusher is started to crush the lepidolite smelting slag. The crushed lepidolite smelting slag is then passed through a vibrating screen. Lepidolite smelting slag with a particle size less than 10 mm is sent to the intermediate silo for ball milling. Lepidolite smelting slag with an oversized particle size is returned to the jaw crusher for further crushing until the particle size is acceptable. The acceptable lepidolite smelting slag is then fed into a planetary ball mill. Based on the total mass of the lepidolite smelting slag, 0.5 wt% anhydrous ethanol is added to the acceptable lepidolite smelting slag, and ball milling is performed. The ball mill cylinder speed is set to 300 rpm, and the ball milling time is 4 hours. After ball milling, lepidolite smelting slag powder is obtained. S12, Acid Leaching: A dilute sulfuric acid aqueous solution with a sulfuric acid mass percentage of 5% (w / w) is prepared by using industrial pure water and 98% concentrated sulfuric acid as raw materials.

[0083] In the leaching reactor, a 5% (w / w) dilute sulfuric acid aqueous solution was added. The stirrer was started, and the lepidolite smelting slag powder was slowly and evenly added to the leaching reactor while stirring. Acid leaching was carried out for 12 hours. After acid leaching, the leached slurry was obtained. The solid-liquid ratio of the lepidolite smelting slag powder to the dilute sulfuric acid aqueous solution was 1 kg: 10 L, and the stirrer speed was 300 rpm.

[0084] The leached slurry is pumped into a chamber filter press for solid-liquid separation to obtain filtrate and filter cake. The filter cake is washed three times with warm water at 35°C and with a mass equal to that of the filter cake to obtain washing liquid and washed filter cake (pre-treated leaching residue). The filtrate and washing liquid are combined to obtain pre-treated leachate, which is sent to the sedimentation unit. The washed filter cake is then unloaded and sent to the hazardous waste treatment area for harmless treatment.

[0085] S2, Precipitation Unit Sodium hydroxide was added to the pretreated leachate to adjust the pH to 11.0. Chlorine gas was then introduced to carry out an oxidation reaction for 30 minutes. Solid-liquid separation was performed to obtain a primary impurity-removed solution. The pH of the primary impurity-removed solution was adjusted to 8.0, and sodium thallium sulfide and PAM flocculant were added. The mixture was stirred and reacted for 40 minutes. Filtration was then performed to obtain a secondary impurity-removed solution and thallium sulfide precipitate.

[0086] S3, Recycling Unit Electrolysis is performed in a static cell, where the anode is a graphite anode and the cathode is a stainless steel plate.

[0087] Thallium sulfide precipitate was dissolved in sulfuric acid to obtain a Tl₂SO₄ solution; a 20% (w / w) sulfuric acid aqueous solution was added to the Tl₂SO₄ solution to adjust the Tl₂SO₄ concentration. 3+ An electrolyte solution with a concentration of 8.0 g / L was prepared. The electrolyte solution was pumped into a settling tank and heated to 40°C. The DC power supply was turned on, and the electrolysis current density was set to 200 A / m³. 2 Electrolyzed into the electrolyte Tl 3+ Electrolysis was stopped when the concentration dropped to 5.0 g / L. The metallic thallium deposited at the cathode was recovered, washed with water, vacuum dried, and then smelted and cast into ingots to obtain the product metallic thallium.

[0088] Comparative Example 2 A method for recovering thallium from lepidolite smelting slag includes the following steps: S1, Pretreatment Unit: Mechanical activation and acid leaching of lepidolite smelting slag; S11. Mechanical Activation: 1 kg of lepidolite smelting slag is fed into the silo via a conveyor belt. The jaw crusher is started to crush the lepidolite smelting slag. The crushed lepidolite smelting slag is then passed through a vibrating screen. Lepidolite smelting slag with a particle size less than 10 mm is sent to the intermediate silo for ball milling. Lepidolite smelting slag with an oversized particle size is returned to the jaw crusher for further crushing until the particle size is acceptable. The acceptable lepidolite smelting slag is then fed into a planetary ball mill. Based on the total mass of the lepidolite smelting slag, 0.5 wt% anhydrous ethanol is added to the acceptable lepidolite smelting slag, and ball milling is performed. The ball mill cylinder speed is set to 450 rpm, and the grinding time is 5 hours. After ball milling, lepidolite smelting slag powder is obtained. S12, Acid Leaching: A dilute sulfuric acid aqueous solution with a sulfuric acid mass percentage of 2% (w / w) is prepared using industrial pure water and 98% concentrated sulfuric acid as raw materials.

[0089] In the leaching reactor, a 2% (w / w) dilute sulfuric acid aqueous solution was added. The jacket heating system was turned on, and the dilute sulfuric acid aqueous solution was heated to 62°C. The stirrer was started, and the lepidolite smelting slag powder was slowly and evenly added to the leaching reactor while stirring. Acid leaching was carried out for 2 hours. After acid leaching, the leached slurry was obtained. The solid-liquid ratio of the lepidolite smelting slag powder to the dilute sulfuric acid aqueous solution was 1 kg: 7 L, and the stirrer speed was 180 rpm.

[0090] The leached slurry is pumped into a chamber filter press for solid-liquid separation to obtain filtrate and filter cake. The filter cake is washed twice with warm water at 45°C and twice the mass of the filter cake to obtain washing liquid and washed filter cake (pretreatment leaching residue). The filtrate and washing liquid are combined to obtain pretreatment leachate, which is sent to the subsequent electrochemical unit. The washed filter cake (pretreatment leaching residue) is unloaded and sent to the hazardous waste treatment area for harmless treatment.

[0091] S2, Electrochemical Regulation and Directed Migration Unit S21, Electrochemical Regulation: A dual-chamber electrolyzer is designed, comprising an anode chamber and a cathode chamber, with a perfluorosulfonic acid (PFSA) cation exchange membrane between them. The anode chamber is connected to an anode storage tank, and the cathode chamber is connected to a cathode storage tank. A porous graphene / MnO2 composite electrode (specific surface area >1200 m²) is used in the cathode chamber. 2 / g) is used as the cathode, and a titanium-based DSA electrode is used as the anode in the anode chamber. When a DC electric field is applied, positively charged cations move towards the cathode, while negatively charged anions move towards the anode. Only cations (such as Tl) are allowed to pass through the perfluorosulfonic acid (PFSA) cation exchange membrane between the anode and cathode chambers. 3+ H + Cu 2+ Zn 2+ (etc.) pass through, while blocking anions (such as SO4) 2- ).

[0092] A dilute sulfuric acid aqueous solution with a sulfuric acid concentration of 1.0 mol / L was prepared using industrial pure water and 98% concentrated sulfuric acid as raw materials.

[0093] The pretreated leachate from S1 is pumped into the anode chamber storage tank. At this point, the pretreated leachate is the anolyte, with a pH of 1. A pre-prepared 1.0 mol / L dilute sulfuric acid solution is pumped into the cathode chamber storage tank. This 1.0 mol / L dilute sulfuric acid solution is the catholyte. The circulation pump is started, causing the anolyte to circulate in the anode chamber storage tank and on the anode surface of the anode chamber at a flow rate controlled at 0.5 m / s. The catholyte circulates in the cathode chamber storage tank and on the cathode surface of the cathode chamber at a flow rate controlled at 0.5 m / s. After the anolyte and catholyte have stabilized in their respective storage tanks and on the electrode surfaces, the electrolysis power supply is connected. Electrolysis is performed in constant potential DC mode to form a DC electric field. The anode potential (relative to the Ag / AgCl reference electrode) is set at 1.5V (continuous voltage output, no pulse).

[0094] S22, Directional Migration: Utilizing the introduced pulsed electric field, Tl 3+ An automatic concentration difference and gradient pH control system is used to establish an electrochemical gradient field between the anode and cathode chambers, promoting the directional migration of thallium ions. Through solid-liquid separation, the Tl(OH)3 precipitate is separated from other ions. The specific steps are as follows: An initial solution containing acetate-sodium acetate buffer is added to the cathode chamber, and the automatic pH control system is activated to maintain the pH of the mixture of the initial solution containing acetate-sodium acetate buffer and the catholyte in the cathode chamber at 4.5, thus creating a transmembrane pH gradient between the cathode and anode chambers. During electrochemical regulation in S21, the potential gradient constructed by the DC electric field and Tl between the cathode and anode chambers... 3+ The concentration gradient formed by the huge concentration difference across the membrane and the transmembrane pH gradient regulated by the automatic pH control system together form an electrochemical gradient field, with Tl in the anode chamber... 3+ Under the influence of an electrochemical gradient field, Tl migrates through the perfluorosulfonic acid (PFSA) cation exchange membrane to the cathode chamber, and enters the cathode chamber. 3+ Rapid hydrolysis forms a suspension containing Tl(OH)3. The suspension in the cathode chamber is then overflowed into a settling tank for 8 hours to separate the solid and liquid, yielding a clear liquid and a Tl(OH)3 filter cake. The Tl(OH)3 filter cake is sent to a subsequent recovery unit, while the clear liquid is returned to the cathode chamber.

[0095] As electrolysis and ion migration continue, the thallium component of the anolyte in the anode chamber of S2 is continuously converted into Tl. 3+ The thallium migrates across the membrane, and the anolyte gradually transforms into a thallium-lean anolyte. When the concentration of thallium in the thallium-lean anolyte decreases to 0.05 mol / L, the thallium-lean anolyte is recycled back to the pretreatment unit for washing the filter cake.

[0096] Cathode Chamber: In the cathode chamber, the cathode is a porous graphene / MnO2 composite electrode. Under the action of a pulsed electric field, the main side reaction of hydrogen evolution occurs on the cathode surface: 2H+ + + 2e - → H2↑, the generated hydrogen gas is discharged through the exhaust port above the cathode chamber.

[0097] S3, Continuous Recovery Unit: Employs electrodeposition technology to achieve thallium electrodeposition; Electrodeposition is performed using an electrolytic cell with a traditional square tank structure. The cell body is made of polyvinylidene fluoride (PVDF), and the inner wall of the cell is equipped with a titanium-based iridium tantalum oxide-coated DSA anode. The cathode is located in the center of the cell and is arranged opposite to the titanium-based iridium tantalum oxide-coated DSA anode. The cathode is a flat stainless steel cathode.

[0098] The Tl(OH)3 filter cake was added to a dissolving vessel, and 0.7 L of a 120 g / L sulfuric acid aqueous solution was slowly added. The mixture was stirred until the Tl(OH)3 filter cake was completely dissolved, thus preparing the electrolyte. The electrolyte was pumped into an electrolytic cell, where it continuously circulated. The electrolyte was heated to 50°C, and the DC power supply was turned on, setting a constant current density of 300 A / m. 2 Electrodeposition begins. The continuous circulation of the electrolyte within the electrolytic cell refers to the following: the electrolyte enters from the bottom of the electrolytic cell, flows upwards in the electrode region between the anode and cathode of the titanium-based iridium-tantalum oxide coated DSA, exits from the top of the electrolytic cell, and returns to the bottom of the electrolytic cell via external pipelines, a circulation tank, and a transfer pump, forming a closed-loop circulation circuit.

[0099] Periodic analysis (every 30 minutes) of Tl in the electrolyte 3+ Concentration, when Tl 3+ When the concentration drops to 10 g / L, electrodeposition is stopped, the cathode is lifted out, and the spongy thallium metal deposited on the cathode is peeled off with a scraper. The collected spongy thallium metal is washed with water and vacuum dried before being sent to a medium-frequency induction furnace for melting and casting to obtain the product thallium metal. The discharged lean thallium electrolyte is returned to the electrochemical regulation and directional migration unit.

[0100] Comparative Example 3 A traditional method for extracting thallium from lepidolite smelting slag includes the following steps: S1. Pretreatment Unit: Mechanical activation and acid leaching of lepidolite smelting slag. S11. Mechanical Activation: 1 kg of lepidolite smelting slag is fed into the silo via a conveyor belt. A jaw crusher is started to crush the slag. The crushed slag is then passed through a vibrating screen. Slag with a particle size less than 10 mm is fed into an intermediate silo for ball milling. Slag with oversized particles is returned to the jaw crusher for further crushing until the particle size is within acceptable limits. The sized slag is then fed into a planetary ball mill. Based on the total mass of the slag, 0.5 wt% anhydrous ethanol is added. A traditional drum ball mill is used for grinding, with the mill speed set at 170 rpm and the milling time at 10 hours. After milling, lepidolite smelting slag powder is obtained. S12, Acid Leaching: A dilute sulfuric acid aqueous solution with a sulfuric acid mass percentage of 2% (w / w) is prepared using industrial pure water and 98% concentrated sulfuric acid as raw materials.

[0101] In the leaching reactor, a 2% (w / w) dilute sulfuric acid aqueous solution was added. The stirrer was started, and the lepidolite smelting slag powder was slowly and evenly added to the leaching reactor while stirring. Acid leaching was carried out for 2 hours. After acid leaching, the leached slurry was obtained. The solid-liquid ratio (mass / volume) of the lepidolite smelting slag powder and the dilute sulfuric acid aqueous solution was 1 kg: 8 L, and the stirring speed was 200 rpm.

[0102] The leached slurry is pumped into a chamber filter press for solid-liquid separation to obtain filtrate and filter cake. The filter cake is washed twice with warm water at 35°C with a mass equal to that of the filter cake to obtain washing liquid and washed filter cake (pre-treatment leaching residue). The filtrate and washing liquid are combined to obtain pre-treatment leachate, which is sent to the subsequent chemical precipitation unit. The washed filter cake (pre-treatment leaching residue) is unloaded and sent to the hazardous waste treatment area for harmless treatment.

[0103] S2, Chemical Precipitation Unit The pretreated leachate obtained from the pretreatment unit is directly fed into the oxidation precipitation reactor, where H2O2 is added to remove Tl. + Oxidized to Tl 3+ The solution was obtained, and then the pH of the solution was adjusted to 11 to form Tl(OH)3 precipitate. The precipitate was collected by filtration.

[0104] S3, Recycling Unit The Tl(OH)3 filter cake (precipitate) collected in step S2 was washed 2-3 times with a small amount of deionized water and then vacuum dried at 60-80℃ to constant weight.

[0105] The dried Tl(OH)3 solid was placed in a muffle furnace and first dehydrated at 200℃ to convert it into Tl2O3 (thallium trioxide). The temperature was then raised to 350℃ to thermally decompose it into Tl2O (thallium suboxide).

[0106] The decomposition product Tl₂O was transferred to a reduction furnace, and hydrogen gas was introduced to carry out a reduction reaction at 400℃: Tl₂O + H₂ → 2Tl + H₂O (heating). The spongy metallic thallium obtained by reduction was cooled and then sent to a medium-frequency induction furnace (temperature controlled at 350℃) for melting and casting under inert gas protection to obtain the product metallic thallium.

[0107] Detection and Analysis 1. Tl + Determination of leaching rate, product purity and recovery rate (1) Determination of the mass fraction of Tl in lepidolite smelting slag The mass fraction of Tl in the lepidolite smelting slag of Examples 1, 2, 3, Comparative Examples 1, 2, and 3 was determined, and the steps are as follows: Sampling: Representative analytical samples were obtained from 1 kg of lepidolite smelting slag using standard sampling methods such as the quartering method.

[0108] Sample preparation and digestion: The analytical sample is dried and ground to an analytical fineness (below 200 mesh) to obtain analytical sample powder. Then, 0.1 g of the analytical sample powder is accurately weighed and placed in a polytetrafluoroethylene digestion vessel. A mixed strong acid (such as a nitric acid-hydrofluoric acid-perchloric acid system) is added, and microwave digestion is performed using a programmed temperature ramp to completely dissolve all components (including thallium) in the analytical sample powder into the solution. After digestion, the acid is usually removed, and the solution is brought to a final volume of 50 mL to obtain the digestion solution.

[0109] Sample detection: The Tl content in the digestion solution can be determined by measuring it using a standard inductively coupled plasma mass spectrometer (ICP-MS). + The concentration of the elements is as follows. Relevant national standard methods can be found in "HJ 700-2014 Determination of 65 Elements in Water by Inductively Coupled Plasma Mass Spectrometry".

[0110] (2) Tl in the pretreated leachate + Concentration analysis and detection The Tl content in the pretreated leachates of Examples 1, 2, 3, Comparative Examples 1, 2, and 3 was analyzed respectively. + The concentration was analyzed and detected using the following steps: Pretreatment of the leachate: The pretreatment leachate has a complex matrix and may contain high concentrations of acids and other metal ions. Before direct analysis, the pretreatment leachate needs to be diluted 1000 times with a 2% (w / w) dilute nitric acid aqueous solution to ensure that the thallium concentration falls within the linear operating range of the instrument and to reduce matrix effects.

[0111] Detection of diluted pretreated leachate: Referring to HJ 700-2014 "Determination of 65 Elements in Water by Inductively Coupled Plasma Mass Spectrometry", the concentration of Tl in the diluted pretreated leachate was analyzed. + The concentration was detected.

[0112] Leaching rate calculation: Thallium ion leaching rate in pretreated leachate ( The calculation formula for ) is shown in equation (IX): (IX) In equation (IX), This indicates the mass concentration of thallium ions in the pretreated leachate, expressed in mg / L. This indicates the volume of the pretreated leachate, expressed in liters (L). The mass of the lepidolite smelting slag is expressed in kg, and is 1 kg in Examples 1, 2, 3, Comparative Examples 1, 2, and 3. This indicates the mass fraction of thallium in lepidolite smelting slag, expressed in mg / kg.

[0113] (3) Recovery rate of thallium metal in the product The spongy thallium metal after washing and vacuum drying was precisely weighed using an analytical balance, and the total mass of the thallium metal (M_product) was finally obtained. Determination of thallium purity: The purity (P_product) of thallium was determined by a standard inductively coupled plasma mass spectrometer (ICP-MS) according to the standard "HJ 700-2014 Determination of 65 Elements in Water Quality by Inductively Coupled Plasma Mass Spectrometry".

[0114] The formula for calculating the recovery rate of product metal Tl is shown in equation (X): (X) In formula (X), M_product represents the total mass of thallium metal in the product, in mg; P_product represents the purity of thallium metal in the product, in %; and M_initial_Tl represents the total mass of thallium contained in the lepidolite smelting slag, in mg / kg.

[0115] Tl in pretreatment leachate + The results of the determination of leaching rate, metal thallium purity and metal Tl recovery rate are shown in Table 1: Table 1. Tl in pretreated leachate + Results of leaching rate, thallium purity and Tl recovery rate determination

[0116] Table 1 shows that the Tl content in the pretreatment leachate of Examples 1, 2, and 3 is... + The leaching rates were 95.3%, 96.4%, and 97.2%, respectively, with an optimal leaching rate of 97.2% under the best conditions (Example 3). In contrast, the Tl content in the pretreated leachate of Comparative Example 1 (without intermittent ball milling and without heated acid leaching) was significantly lower. + The leaching rate was only 86.5%, while the Tl content in the pretreated leachate of Comparative Example 2 (without intermittent ball milling) was significantly lower. + The leaching rate was 96.5%, and the Tl content in the pretreated leachate of Comparative Example 3 (conventional drum ball milling without heated acid leaching) was significantly lower. + The leaching rate was the lowest, at only 79.4%. This indicates that the combined treatment method of mechanical activation by planetary ball milling and heated acid leaching significantly improved the leaching efficiency of the pretreated leachate.

[0117] In Examples 1, 2, and 3, the purity of metallic thallium reached 99.0%~99.5%, while the purity of metallic thallium in Comparative Example 1 was 95.0%, in Comparative Example 2 it was 92.2%, and in Comparative Example 3 it was only 90.5%. This demonstrates that the electrochemical pulsed field synergistic process has a significant advantage in purity control, thanks to its precise potential regulation and directional migration mechanism.

[0118] The recovery rates of metal Tl in Examples 1, 2, and 3 ranged from 91.6% to 93.2%, with Example 3 showing the best recovery rate (93.2%). The recovery rates of metal Tl in Comparative Examples 1, 2, and 3 were 85.4%, 82.1%, and 79.8%, respectively, all significantly lower than those in the Examples. Notably, Comparative Example 2, which used DC electrolysis instead of pulsed electrolysis, achieved a lower recovery rate than Comparative Example 1, indicating that the synergistic effect of the pulsed mode was more crucial.

[0119] In summary, the method for highly selective extraction and high-purity recovery of thallium from lepidolite smelting slag used in Example 3, compared with the traditional process in Comparative Example 3, increases the leaching rate by approximately 17.8 percentage points, the purity by approximately 9.0 percentage points, and the recovery rate by approximately 13.4 percentage points. This fully demonstrates the significant technical advantages of this method for highly selective extraction and high-purity recovery of thallium from lepidolite smelting slag in the recovery of thallium resources from lepidolite slag.

[0120] 2. Determination of toxicity of leaching residue from pretreated lepidolite. The heavy metal leaching performance indicators of the pretreated leaching residues obtained in Examples 1, 2, 3, Comparative Examples 1, 2, and 3 were determined respectively, and the steps are as follows: 100g of dry basis sample of pretreated 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 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 clarified pretreated leachate was then obtained by filtration through a 0.45μm filter membrane. Quantitative analysis of heavy metals in the pretreated leachate was performed. The results are shown in Table 2. Table 2 Heavy metal leaching performance indicators of lepidolite pretreated leaching residue

[0121] Table 2 shows that the heavy metal leaching concentrations in Examples 1, 2, and 3 were low, all below the limits set by DB36 / T1968-2024. Comparative Example 3, using the traditional sulfuric acid leaching method, showed that lead, cadmium, manganese, and thallium in the lithium slag exceeded the limits. Thallium levels were measured using a standard inductively coupled plasma mass spectrometer (ICP-MS). Analysis of Comparative Example 3 shows that, according to the DB36 / T1968-2024 standard, thallium levels exceeded the limit by 7 times. Comparative Examples 1 and 2 also showed slight excesses of thallium in their lithium slag, all failing to meet the emission standards for general solid waste. Accumulated thallium levels will pose a significant environmental hazard. These results indicate that the electrochemical pulsed field synergistic redox control technology for extracting thallium from lepidolite slag effectively achieves comprehensive recovery of thallium from the pretreated lepidolite leaching solution, and the leached lithium slag possesses the chemical safety profile suitable for general solid waste.

Claims

1. A method for selectively extracting and recovering thallium from lepidolite smelting slag, characterized in that, Includes the following steps: S1. Pretreatment of lepidolite smelting slag to obtain thallium using Tl + The pretreated leachate and pretreated leachate residue are released in the form of a solution. S2. Introduce a pulsed electric field to treat the pretreated leachate, so that Tl + Convert to Tl 3+ , to obtain Tl 3+ The pulse products; S3, combined with pulsed electric field, Tl 3+ The concentration difference and the electrochemical gradient field formed by the pH automatic control system affect the concentration of Tl. 3+ The pulse products are treated to promote Tl 3+ Directional migration forms a suspension containing Tl(OH)3. Solid-liquid separation yields a clear liquid and a Tl(OH)3 filter cake. S4. Add an acid solution to the Tl(OH)3 filter cake to obtain an electrolyte. Electrodeposit the electrolyte using a vortex electrodeposition technique to obtain metallic thallium.

2. The method according to claim 1, characterized in that, The pretreatment in S1 includes mechanical activation and acid leaching; Mechanical activation includes: crushing the lepidolite smelting slag to a particle size of less than 10 mm to obtain lepidolite smelting slag with qualified particle size; Anhydrous ethanol was added to the lepidolite smelting slag with qualified particle size, and the mixture was ball-milled to obtain lepidolite smelting slag powder. Acid leaching includes: heating a dilute sulfuric acid aqueous solution, adding the lithium mica smelting slag powder, acid leaching, solid-liquid separation, and obtaining filtrate and filter cake; The filter cake is washed with water or a circulating lean thallium anolyte to obtain a washing solution and pretreated leaching residue. The filtrate and washing liquid are combined to obtain the pretreated leachate.

3. The method according to claim 2, characterized in that, The amount of anhydrous ethanol added is 0.5% of the total mass of the lepidolite smelting slag; The ball milling is carried out using a planetary ball mill with a ball milling cylinder speed of 400~500 rpm and a ball milling time of 4~6 hours. At the same time, an intermittent operation mode is activated, which runs for 30~40 minutes and stops for 10 minutes. The solid-liquid ratio of the lithium mica smelting slag powder and the dilute sulfuric acid aqueous solution is 1 kg: 5~8 L; The mass percentage of sulfuric acid in the first dilute sulfuric acid aqueous solution is 2% (w / w); the first dilute sulfuric acid aqueous solution is heated to 60~65℃; The acid leaching time is 2 hours; The washing is performed 1 to 2 times; The pH of the pretreated leachate is 1.

4. The method according to claim 1, characterized in that, S2 specifically includes: A dual-chamber electrolyzer is designed, comprising an anode chamber and a cathode chamber, with a cation exchange membrane provided between the anode chamber and the cathode chamber. The anode chamber is connected to an anode chamber storage tank, and the cathode chamber is connected to a cathode chamber storage tank. The pretreated leachate is pumped into the anode chamber storage tank, causing it to circulate between the anode chamber storage tank and the first anode surface of the anode chamber, with a flow rate controlled at 0.2~0.5 m / s. The catholyte is pumped into the cathode chamber storage tank, causing it to circulate between the cathode chamber storage tank and the first cathode surface of the cathode chamber, with a flow rate controlled at 0.2~0.5 m / s. Electrolysis employs a constant potential pulse mode to create a pulsed electric field, which is used to control Tl. + Oxidation to produce Tl 3+ , to obtain Tl 3+ The pulse products.

5. The method according to claim 4, characterized in that, The cathode chamber uses materials with a specific surface area >1200 m². 2 / g porous graphene / MnO2 composite electrode is used as the first cathode, and a titanium-based DSA electrode is used as the first anode in the anode chamber. The cation exchange membrane is selected from perfluorosulfonic acid cation exchange membrane. The cathode solution is selected from a dilute sulfuric acid aqueous solution with a sulfuric acid concentration of 0.5~1.0 mol / L. Using an Ag / AgCl electrode as the reference electrode, the peak value of the anode potential pulse was set to 1.3~1.5V; the pulse frequency of the anode potential was 10~20Hz, and the duty cycle was 40~60%. The electrolysis temperature is 40~50℃.

6. The method according to claim 4, characterized in that, S3 specifically includes: An initial solution containing a buffer is added to the cathode chamber, and the automatic pH control system is activated to control the pH of the mixture of the initial solution containing the buffer and the catholyte in the cathode chamber to be 3.5~4.

5. Through pulsed electric field, Tl 3+ The concentration difference and pH automatic control system establishes an electrochemical gradient field between the anode and cathode chambers; Electrochemical gradient field for Tl 3+ The pulse products are treated to promote Tl 3+ Tl migrates directionally from the anode chamber through the cation exchange membrane to the cathode chamber, entering the cathode chamber... 3+ Hydrolysis forms a suspension containing Tl(OH)3; The suspension containing Tl(OH)3 was subjected to sedimentation for 6-8 hours, followed by solid-liquid separation to obtain a clear liquid and a Tl(OH)3 filter cake.

7. The method according to claim 1, characterized in that, S4 specifically includes: A cyclone electrolytic cell is provided, comprising an outer cylinder and an inner cylinder arranged concentrically, the inner cylinder being sleeved inside the outer cylinder, the outer cylinder being the cell body, and a second anode in the form of a cage or plate being provided on the cell wall of the cell body, the inner cylinder being a rotatable second cathode; Add sulfuric acid aqueous solution to Tl(OH)3 filter cake, stir to dissolve, and obtain electrolyte; The electrolyte is pumped into the vortex electrolytic cell, where it circulates continuously and is heated to a set temperature. The rotation speed of the second cathode is adjusted, the DC power supply is turned on, a constant current density is set, and electrodeposition begins. When Tl in the electrolyte 3+ When the concentration drops to 10~20g / L, electrodeposition is stopped, the second cathode is lifted out, and the spongy thallium metal on the second cathode is peeled off. The spongy thallium metal is washed with water, dried, smelted, and cast into ingots to obtain the product thallium metal.

8. The method according to claim 7, characterized in that, The tank is made of polyvinylidene fluoride, the second anode is selected from titanium-based iridium tantalum oxide coated DSA anode, and the second cathode is selected from high-purity stainless steel or titanium rod. The sulfuric acid in the aqueous solution has a concentration of 80-120 g / L. The set temperature is 40~50℃; The rotational speed of the second cathode is 800~1500 rpm; The constant current density is 300 A / m 2 .

9. The method according to claim 1, characterized in that, The chemical composition of the lepidolite smelting slag, by mass percentage, includes: 53.3% SiO2, 25.3% Al2O3, 0.55% Fe2O3, 3.25% CaO, 0.32% MgO, 0.07% SO3, 7.45% Na2O, 5.12% K2O, 2.24% TiO2, and other impurities. The loss on ignition rate of the lepidolite smelting slag is 0.15%, and the thallium content in the lepidolite smelting slag is 20.42 mg / kg.

10. The method according to claim 1, characterized in that, The leaching rate of the metal thallium is 95.3-97.2%, the recovery rate is 91.6-93.2%, and the purity is >99%.

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