Method for selectively extracting and preparing metal thallium from thallium-containing material

Abstract

The invention belongs to the field of solid waste treatment, and particularly relates to a method for selectively extracting and preparing metal thallium from a thallium-containing material, which comprises the following steps: carrying out reduction alkaline leaching on the thallium-containing material to obtain a Tl < + > alkaline leaching solution; the Tl < + > alkaline leaching solution is subjected to vulcanization precipitation, and Tl2S slag is obtained; the Tl2S slag is subjected to potential-controlled oxidation acid leaching, negative-valence sulfur in the Tl2S slag is selectively oxidized, and Tl < + > acid leaching liquid is obtained; mixing the Tl < + > pickle liquor and a compound additive, and then carrying out rotational flow electrodeposition treatment to obtain metal thallium and a liquid after rotational flow electrodeposition; the cathode material in the rotational flow electrodeposition is an aluminum plate, a copper plate or a titanium plate; the compound additive comprises a component A and a component B, and the component A comprises a Tl liquefying agent; the component B comprises at least one of gallic acid, benzalacetone, benzophenone, bone glue, polyether, FES-993 and quaternary ammonium salt; and the current density of the rotational flow electrodeposition is 50-500A / m < 2 >. According to the synergistic process, high-quality metal thallium can be efficiently recycled.

Description

Technical Field

[0001] The present invention belongs to the field of metallurgical chemistry, and particularly relates to a method for selectively extracting and preparing metallic thallium from thallium-containing materials. Background Art

[0002] Thallium is a rare and dispersed metallic element, and its Clark value is only 7×10 -5 %. Thallium rarely forms independent ore deposits in nature. It often occurs in natural minerals such as sphalerite, zinc blende, chalcopyrite and lepidolite, and disperses and migrates into the environment during subsequent industrial activities. Currently, the steel industry, lead-zinc smelting industry, cement industry and lithium product processing industry are the main sources of thallium emissions.

[0003] Thallium is a typical toxic heavy metal element and is one of the 13 priority-controlled metal pollutants recognized in the world. Its toxicity is greater than that of Cd, Pb, and Hg. Thallium mainly enters the human body through drinking water, food intake or skin contact, and accumulates in the human body, with latent and strong cumulative toxicity. The toxic effects of thallium on the human body are mainly manifested as neurotoxicity, which can cause damage to the functions of multiple organs such as the kidneys and liver.

[0004] Thallium not only causes serious harm to organisms, but also has a very great impact on the production process. Taking the lead-zinc smelting industry as an example: Relevant data show that only the zinc smelting industry will produce nearly 500 tons of thallium-rich hazardous waste, 60,000 tons of thallium-containing solid waste, 10,000 tons of high-thallium soot and 3 million cubic meters of thallium-containing contaminated acid every year. To alleviate the shortage of raw materials, and in addition, in recent years, the environmental protection department has imposed stricter restrictions on the long-term stacking of zinc oxide soot by enterprises. Therefore, smelters have strengthened the utilization of secondary resources such as zinc oxide soot in the production process. Zinc oxide soot has a high zinc content and relatively simple composition, which can well make up for the shortage of primary zinc ore resources. However, it contains a high content of thallium. If it is directly returned to leaching, thallium will enter the leaching solution, which will have an adverse impact on subsequent smelting. Especially in the subsequent zinc electrodeposition process, since the electrode potential of thallium is more positive than that of zinc, "burning of plates" is likely to occur during the electrowinning process, and it will also induce "burning of plates" of other impurities, which will have an adverse impact on the product output and technical and economic indicators, and even lead to the inability to carry out normal production. In addition, at present, the high-thallium soot produced by many lead-zinc smelting enterprises is mainly sold externally. However, with the improvement of the requirements of the environmental protection department for the transportation of thallium-containing materials, the external sales cost and risk are increasing day by day, and the corresponding income is also decreasing. In addition, a large amount of thallium-containing wastewater and waste residues will also be generated during the process of extracting lithium from lithium ore. These wastes will not only affect the purity of the product, but also cause the upgrade of the waste hazard level, resulting in the current situation that a large amount of waste is difficult to handle and accumulates in large quantities. In short, thallium is extremely harmful in industries such as chemical engineering and smelting. Therefore, the removal of thallium in the above processes is crucial.

[0005] Although thallium is significantly toxic, thallium and related alloys have certain uses in high-temperature superconductivity, medical detection, precision optical instruments, etc. Therefore, based on the requirements of environmental protection, high-value utilization, and resource reserve, etc., the clean extraction and metallization recovery of thallium from thallium-containing materials are of great significance. Summary of the Invention

[0006] The object of the present invention is to provide a method for selectively extracting and preparing metallic thallium from thallium-containing materials, aiming to solve the problems of great thallium hazards, difficult open-circuit, and poor resource reserve in the current thallium-related industries.

[0007] The content of thallium in natural minerals or industrial products is extremely low. Therefore, the single leaching concentration is extremely low, and it is difficult to extract and prepare metallic thallium therefrom. In view of this problem, the present invention proposes a process route for leaching-enriching-electrowinning to extract and prepare metallic thallium. Among them, the electrodeposition of thallium is a major problem in the industry. At present, the industry mainly uses the replacement of active metals to extract and prepare metallic thallium. This method has the problem of low purity of metallic thallium and can only obtain sponge thallium. Although a small number of researchers have tried to use plate electrowinning to prepare metallic thallium, this method can also only produce sponge thallium. In addition, the earlier research results of the present invention show that the electrodeposition of thallium not only has the problem of difficult plate formation, but also thallium has a problem of valence change during the electrowinning process, that is, Tl + will be oxidized to Tl 3+ , and cause Tl + and Tl 3+ to reciprocate between the anode and cathode, and at the same time, Tl 3+ will also react with cathode Tl, resulting in the back dissolution of thallium at the cathode. In view of the above problems, the following improvement solutions are proposed in this study:

[0008] A method for selectively extracting and preparing metallic thallium from thallium-containing materials, the thallium-containing materials are subjected to reduction alkali leaching to obtain a Tl + alkali leaching solution; the Tl + alkali leaching solution is subjected to sulfide precipitation to obtain a Tl2S slag;

[0009] The Tl2S slag is subjected to controlled-potential oxidation acid leaching to selectively oxidize the sulfur therein to obtain a Tl + acid leaching solution;

[0010] The Tl + acid leaching solution and a composite additive are mixed and then subjected to hydrocyclone electrodeposition treatment to obtain metallic thallium and a hydrocyclone electrodeposition post-liquid;

[0011] The cathode material in the cyclone electrodeposition is aluminum plate, copper plate or titanium plate; the composite auxiliary agent includes component A and component B, wherein component A includes a Tl liquefier; the component B includes at least one of gallic acid, benzylideneacetone, benzophenone, bone glue, polyether, FES-993, quaternary ammonium salt; the cyclone electrodeposition current density is 50-500 A / m 2 .

[0012] In the present invention, innovatively, the thallium-containing material is subjected to reduction alkali leaching treatment to obtain a Tl + alkali leaching solution, and then subjected to sulfide precipitation to obtain a Tl2S slag, and then the Tl2S slag is subjected to controlled potential oxidation acid leaching to selectively oxidize the sulfur in the Tl2S slag without oxidizing the Tl therein + , so as to obtain a Tl + acid leaching solution; and then innovatively, the Tl + acid leaching solution is subjected to cyclone electrodeposition with the assistance of the composite auxiliary agent, so that metallic thallium can be efficiently recovered and its morphology can be improved.

[0013] In the present invention, the thallium-containing material is at least one waste material of liquid and solid containing Tl + and / or Tl 3+ . In particular, it is directed to a waste liquid containing Tl + and Tl 3+ . Tl 3+ and Tl + will reciprocally cycle between the anode and the cathode, and it is difficult for Tl to be plated on the plate. For this difficult-to-treat object, the process described in the present invention can also obtain good treatment effects.

[0014] In the present invention, the thallium-containing material is subjected to reduction alkali leaching in an alkali solution containing a reducing agent.

[0015] Preferably, the reducing agent includes at least one of sodium borohydride and ascorbic acid.

[0016] Preferably, the dosage of the reducing agent is 1-10 times, further preferably 7-9 times, of the theoretical dosage for reducing trivalent thallium in the thallium-containing material to monovalent thallium.

[0017] Preferably, the alkali solution is an aqueous solution of at least one of alkali metal hydroxides and carbonates.

[0018] Preferably, the pH in the reduction alkali leaching stage is 7-12, and further can be 9-9.5.

[0019] Preferably, the reduction leaching temperature is 50-100 °C, further can be 75-85 °C, and the time for reduction alkali leaching is 1-2 h.

[0020] In the present invention, the Tl + alkali leaching solution and a water-soluble sulfide are mixed for sulfide precipitation to obtain a Tl2S slag.

[0021] Preferably, the water-soluble sulfide includes at least one of sodium sulfide, potassium sulfide, and ammonium sulfide.

[0022] Preferably, the dosage of the water-soluble sulfide is 1 to 3 times, further preferably 1.5 to 2.5 times, of the theoretical amount for converting thallium in the alkali leaching solution into Tl2S. + The temperature in the sulfide precipitation stage is 25 to 85°C, further preferably 35 to 45°C.

[0023] Preferably, the sulfide precipitation reaction time is 0.5 to 2 h, further preferably 1 to 1.5 h.

[0024] In the present invention, the acid in the controlled-potential oxidation acid leaching process is a sulfuric acid solution. The controlled-potential oxidation acid leaching process also contains an oxidizing component. For example, it can be concentrated sulfuric acid, hydrogen peroxide, etc.

[0025] Preferably, the oxidation potential in the controlled-potential oxidation acid leaching process is 400 to 500 mV, further preferably 430 to 480 mV.

[0026] Preferably, the temperature in the controlled-potential oxidation acid leaching process is 30 to 90°C, further preferably 35 to 45°C;

[0027] Preferably, the time in the controlled-potential oxidation acid leaching process is 0.5 to 4 h, further preferably 2.5 to 3.5 h.

[0028] In the present invention, there is no special requirement for the concentration of thallium in the acid leaching solution. For example, it can be controlled below 80 g / L, such as 30 to 50 g / L; the free acid concentration is below 100 g / L, and further preferably 20 to 30 g / L.

[0029] In the present invention, the anode in the vortex electrodeposition is a titanium-based ruthenium-iridium rod or a graphite rod. + Thallium in the acid leaching solution + In the present invention, in the composite auxiliary agent, component A includes at least one of ascorbic acid, gallic acid, and tartaric acid.

[0030] Component A is 0.9 to 2 times, preferably 1 to 1.5 times, the molar amount of thallium in the acid leaching solution.

[0031] The concentration of component B in the solution system of the vortex electrodeposition is 0.01 to 2 g / L, and further preferably 0.1 to 1 g / L.

[0032] The component A is + 0.9 to 2 times, preferably 1 to 1.5 times, the molar amount of thallium in the acid leaching solution.

[0033] In the present invention, the current density of the vortex electrodeposition can be further 100 to 300 A / m

[0034] In the present invention, the current density of the vortex electrodeposition can be further 100 to 300 A / m2 。

[0035] In the present invention, the solution of the system in the swirl electrodeposition stage is in a swirl state.

[0036] In the present invention, the post-liquid of swirl electrodeposition is subjected to micro-current assisted replacement for thallium removal, and the steps are as follows: the post-liquid of swirl electrodeposition is introduced into a micro-current assisted replacement tank; a zinc plate is used as the working electrode, an inert material is used as the counter electrode, and a current is applied to the surface of the zinc plate for replacement to obtain sponge thallium and post-replacement liquid.

[0037] In the present invention, the micro-current is one of unidirectional or pulsed current; the micro-current density is 10-50 A / m 2 . The temperature of micro-current replacement is 25-50 °C.

[0038] Beneficial effects

[0039] The present invention provides a synergistic process of reduction alkali leaching treatment-Tl + sulfide precipitation-selective controlled potential acid leaching of Tl + acid leaching solution and swirl electrodeposition assisted by a composite additive, so that high-quality metallic thallium can be efficiently recovered. Brief description of the drawings

[0040] Figure 1 is a schematic process flow diagram of the present invention;

[0041] Figure 2 is a schematic diagram of the swirl electrodeposition device used in all examples and comparative examples;

[0042] Figure 3 XRD pattern of cathode thallium in Example 1;

[0043] Figure 4 is the macroscopic morphology of cathode thallium in Example 1;

[0044] Figure 5 is the macroscopic morphology and XRD characterization of the cathode product in Comparative Example 2;

[0045] Figure 6 is the morphology of the cathode product in Comparative Examples 5, 7, and 8; among them, a) the morphology of the cathode product in Comparative Example 4; b) the morphology of the cathode product in Comparative Example 6; c) the morphology of the cathode product in Comparative Example 7. Specific embodiments

[0046] The following examples are intended to further illustrate the present invention rather than limit the present invention.

[0047] An optional method for selectively extracting and preparing metallic thallium from thallium-containing materials according to the present invention, the steps of which include, for example:

[0048] (1) Alkaline reduction leaching. The thallium-containing material and a certain proportion of reducing agent are put into a weakly alkaline solution and leached for a period of time, and then filtered to obtain a thallium-containing alkaline leaching solution and alkaline leaching residue.

[0049] (2) Sulfide precipitation of thallium from the alkaline leaching solution. A certain amount of sodium sulfide is added to the alkaline leaching solution described in step (1) and reacted for a period of time, and then filtered to obtain a post-sulfidation solution and thallium sulfide residue.

[0050] (3) Controlled-potential oxidation acid leaching. The thallium sulfide residue obtained in step (2) is placed in an acidic solution, and a certain amount of oxidant is added at the same time and leached for a period of time, and then filtered to obtain an acid leaching solution and leaching residue.

[0051] (4) Cyclone electrodeposition of the acid leaching solution. A certain amount of component A and component B are added to the acid leaching solution obtained in step (3); then metal thallium and post-cyclone electrodeposition solution are obtained through cyclone electrodeposition.

[0052] In addition, the present invention selectively includes step (5) Micro-current assisted replacement for thallium removal from the post-cyclone electrodeposition solution. The post-electrodeposition solution obtained in step (4) is introduced into a micro-current assisted replacement tank; a zinc plate is used as the working electrode, and an inert material such as a ruthenium-iridium mesh is used as the counter electrode, and a specific current is applied to the surface of the zinc plate for a period of time to obtain sponge thallium and post-replacement solution.

[0053] In the present invention, the reducing agent described in step (1) is sodium borohydride, ascorbic acid, etc. The dosage of the reducing agent is 1 to 10 times, further preferably 1 to 3 times, of the theoretical dosage for reducing trivalent thallium in the thallium-containing material to monovalent thallium. The alkaline leaching solution is a sodium carbonate and / or sodium hydroxide solution.

[0054] The pH range of the alkaline solution described in step (1) is 7 to 12. The leaching temperature described in step (1) is 50 to 100 °C, and further preferably 65 to 85 °C considering actual operation; the leaching time is 1 to 2 h.

[0055] If the alkaline leaching solution contains halogen ions such as F, Cl, I, etc., the filtration temperature described in step (1) can be controlled above 50 °C, and further preferably 75 to 95 °C.

[0056] The dosage of sodium sulfide described in step (2) is 1 to 3 times, further preferably 1.5 to 2.5 times, of the theoretical amount for converting thallium in the alkaline leaching solution into Tl2S; or the concentration of free Na2S in the post-sulfidation solution is controlled to be 3 to 5 g / L, and the sulfide precipitation temperature is 25 to 85 °C; the sulfide precipitation reaction time is 0.5 to 2 h.

[0057] In the present invention, the acidic solution described in step (3) is concentrated sulfuric acid or sulfuric acid solution, and the oxidation potential is 400 to 500 mV. The purpose of controlled-potential oxidation leaching is to oxidize S2O3 2- to stable elemental S and SO42- and will not oxidize Tl + to Tl 3 + nor will it react with Tl + to form a precipitate, or a combination of one or more substances.

[0058] In the present invention, the temperature of the oxidative acid leaching reaction described in step (3) is 30 - 90 °C; the oxidation reaction time is 0.5 - 4 h.

[0059] In the present invention, the concentration of Tl in the electrolyte described in step (3) + is less than 80 g / L; the free acid concentration is less than 100 g / L, and more preferably less than 30 g / L.

[0060] In the present invention, the component A described in step (4) is a combination of one or more substances that can preferentially discharge at the anode with Tl + or (and) can form a complex that is not easily oxidized with Tl + For example, it can be a combination of one or more of ascorbic acid, gallic acid, and tartaric acid.

[0061] In the present invention, the component B described in step (4) is a combination of one or more of gallic acid, benzylideneacetone, benzophenone, bone glue, polyether, FES - 993, and quaternary ammonium salt.

[0062] In the present invention, the anode of the hydrocyclone electrodeposition tank described in step (4) is a titanium - based ruthenium - iridium rod or a graphite rod, and more preferably a titanium - based ruthenium - iridium rod; the cathode material is an aluminum plate, a copper plate, or a titanium plate. Considering the adhesion strength of thallium on the surface, it is more preferably an aluminum plate. The hydrocyclone electrodeposition temperature is 15 - 40 °C. The hydrocyclone electrodeposition current density is 50 - 500 A / m 2 .

[0063] In the present invention, according to the method for selectively extracting and preparing metallic thallium from thallium - containing materials described in claim 1, the micro - current in step (5) is one of unidirectional or pulsed current; the micro - current density is 10 - 50 A / m 2 . The micro - current replacement temperature is 25 - 50 °C.

[0064] The hydrocyclone electrodeposition equipment described in the present invention can be well - known. For example, its schematic diagram can be seen Figure 2 (unit: mm): 1 - hydrocyclone outlet; 2 - cathode of hydrocyclone device; 3 - valve; 4 - external circulation pump; 5 - circulating liquid storage tank; 6 - hydrocyclone inlet; 7 - anode of hydrocyclone device; 8 - hydrocyclone cavity.

[0065] In the following cases, the unit of the liquid - to - solid ratio is mL / g.

[0066] In the following cases (except Comparative Example 8), the solution of the system in the cyclone electrodeposition stage is in a cyclone state. For example, the input speed of the cyclone can be 1-2 L / min.

[0067] Example 1

[0068] The contents of lead, zinc, cadmium, and thallium in a certain thallium-containing zinc oxide dust are 21.3%, 28.9%, 7.23%, and 1.44% respectively. In thallium, it contains Tl + and Tl 3+ , among which, Tl + accounts for 92.25% of the total Tl.

[0069] Step 1: Take 2 kg of this dust, leach for 2 h under the conditions of a liquid-solid ratio of 5:1, a temperature of 80 °C, an end-point pH = 9 (regulated by sodium hydroxide), and a sodium borohydride dosage of 1.5 g / L (the dosage is 7.26 times the molar amount of thallium in the dust); then filter while it is hot to obtain an alkali leaching solution and an alkali leaching residue. 3+ After that, filter while it is hot to obtain an alkali leaching solution and an alkali leaching residue.

[0070] Step 2: Add sodium sulfide nonahydrate (the theoretical dosage is 1.45 times) to the alkali leaching solution, and then react at 40 °C for 1 h and filter to obtain a thallium sulfide residue and a post-sulfidation solution.

[0071] Step 3: Place the obtained thallium sulfide residue in 28 g of concentrated sulfuric acid, control the oxidation potential to be 430-450 mV, and react for 3 h (the temperature during the treatment process is 35-45 °C); immediately, add 486 mL of pure water to the leaching slurry to dissolve for 2 h, and then filter to obtain a leaching solution and a leaching residue.

[0072] Step 4: Add ascorbic acid (Component A, which is 1.1 times the molar amount of Tl in the leaching solution of Step 3) and benzylideneacetone (Component B, with a concentration of 0.5 g / L in the solution) to the above leaching solution, stir to dissolve fully to obtain an electrolyte; then pass the obtained electrolyte into a cyclone electrodeposition cell with an aluminum plate as the cathode and a titanium-based ruthenium-iridium rod as the anode (as shown in + ), and electrodeposit for 13 h under the conditions of a current density of 100 A / m Figure 2 , a temperature of 25 °C, and a cyclone speed of 1.5 L / min to obtain metallic thallium and a post-electrodeposition solution. 2 to obtain metallic thallium and a post-electrodeposition solution.

[0073] Step 5: Pass the above post-electrodeposition solution into a microcurrent-assisted displacement cell, and carry out microcurrent-assisted displacement for deep thallium removal with a zinc plate as the anode and a titanium-iridium-ruthenium mesh as the cathode. Under the conditions of a temperature of 35 °C and a current density of 15 A / m 2 , displace for 4 h to obtain sponge thallium and a deep displacement post-solution.

[0074] After detection and calculation, the leaching rates of lead, zinc, cadmium, and thallium in Step 1 were 6.32%, 0.011%, 0.012%, and 95.26% respectively. The precipitation rate of thallium in Step 2 was 99.25%, and the concentration of thallium in the post-sulfidation solution was 20.01 ppm. The leaching rate of thallium in Step 3 was 98.29%, the concentration of thallium in the leaching solution was 52.20 g / L, and the concentration of free acid in the leaching solution was 30 g / L. The output of metallic thallium after vortex electro-deposition was 25.03 g, the current efficiency was 90.12%, the concentration of thallium in the electrolyte decreased from 52.20 g / L to 1.97 g / L, and the crystal structure of the cathode thallium was complete (as shown in Figure 3 ), and the morphology was good (as shown in Figure 4 ). The replacement rate of thallium in Step 5 was 97.25%, the output of sponge thallium was 1.28 g, the purity of sponge thallium was 75.69%, and the concentration of thallium in the post-replacement solution decreased to 54.21 ppm.

[0075] Example 2

[0076] The contents of calcium, sulfur, titanium, and thallium (including Tl + and Tl 3+ ) in a certain waste flue gas catalyst were 3.29%, 1.20%, 48.15%, and 1.10% respectively. Among Tl + and Tl 3+ , Tl + accounted for 91.35% of the total Tl;

[0077] Step 1: Take 2 kg of this catalyst, leach for 1.5 h under the conditions of liquid-solid ratio of 5:1, temperature of 85 °C, end-point pH = 9.5 (regulated by sodium hydroxide), and sodium borohydride (the dosage is 8.5 times the molar amount of flue dust Tl 3+ ); then filter while it is hot to obtain the alkali leaching solution and alkali leaching residue.

[0078] Step 2: Add sodium sulfide nonahydrate (1.44 times the molar amount of Tl) to the alkali leaching solution, then react at 45 °C for 1 h and filter to obtain thallium sulfide residue and post-sulfidation solution.

[0079] Step 3: Place the obtained thallium sulfide residue in concentrated sulfuric acid, control the oxidation potential at 480 ± 10 mV, and react for 3 h; immediately, add 485 mL of pure water to the leaching slurry to dissolve for 2 h, then filter to obtain the leaching solution and leaching residue.

[0080] Step 4: Add ascorbic acid (Component A, 1.42 times the molar amount of Tl in the leaching solution of Step 3) and benzophenone (Component B, the concentration in the solution is 0.55 g / L) to the above leaching solution, stir and dissolve fully to obtain the electrolyte; then pass the obtained electrolyte into a vortex electro-deposition cell with an aluminum plate as the cathode and a titanium-based ruthenium-iridium rod as the anode, at a current density of 200 A / m 2、Under the conditions of a temperature of 30 °C and a swirling flow rate of 1.5 L / min, electrolytic deposition is carried out for 5 h to obtain thallium metal and the post-electrolytic solution.

[0081] Step 5: The above post-electrolytic solution is introduced into a micro-current assisted displacement tank, and micro-current assisted displacement is carried out to deeply remove thallium with a zinc plate as the anode and a titanium iridium ruthenium mesh as the cathode. Under the conditions of a temperature of 35 °C and a current density of 10 A / m 2 , the displacement time is 4 h to obtain sponge thallium and the post-depth displacement solution.

[0082] Upon detection and calculation, the leaching rates of calcium, sulfur, titanium, and thallium in Step 1 are 8.25%, 0.14%, 0.013%, and 96.35% respectively. The precipitation rate of thallium in Step 2 is 99.38%, and the concentration of thallium in the post-sulfidation solution is 20.69 ppm. The leaching rate of thallium in Step 3 is 98.65%, the concentration of thallium in the leaching solution is 41.20 g / L, and the free acid concentration of the leaching solution is 30 g / L. The output of thallium metal after vortex electrodeposition is 19.23 g, the current efficiency is 90.01%, and the concentration of thallium in the electrolyte is reduced from 41.20 g / L to 2.48 g / L. The displacement rate of thallium in Step 5 is 98.25%, the output of sponge thallium is 1.59 g, the purity of sponge thallium is 76.36%, and the concentration of thallium in the post-displacement solution is reduced to 43.62 ppm.

[0083] Comparative Example 1

[0084] Compared with Example 1, the difference is only that in Step 1, sodium borohydride is not added, and other operations and parameters are the same as those in Example 1.

[0085] Upon detection and calculation, the leaching rates of lead, zinc, and cadmium in Step 1 are basically the same as those in Example 1, which are 6.21%, 0.012%, and 0.011% respectively. However, the leaching rate of thallium is significantly reduced to 90.12%.

[0086] Comparative Example 2

[0087] Compared with Example 1, the difference is only that in Step 3, thallium sulfide slag is leached with 50 g / L dilute sulfuric acid (non-oxidizing), and the oxidation potential in the leaching stage is not controlled within 400 - 500 mV, and other operations and parameters are the same as those in Example 1.

[0088] The leaching rate of thallium in Step 3 is 97.89%, the concentration of thallium in the leaching solution is 51.62 g / L, and the free acid concentration of the leaching solution is 30 g / L. However, during the vortex electrodeposition process, yellow sulfur is precipitated at the cathode and anode (as Figure 5 shown), and no thallium metal is precipitated, which indicates that there is a large amount of low-valent sulfur in the dilute acid leaching solution, which will be preferentially precipitated during the electrolytic deposition process, resulting in the inability to carry out normal electrolytic deposition.

[0089] Comparative Example 3

[0090] Compared with Example 1, the only difference is that in Step 3, thallium sulfide slag is leached with dilute sulfuric acid at a concentration of 50 g / L, ozone is introduced at a rate of 100 mL / min during the leaching process, and the potential during the oxidation process is controlled at 1000 ± 20 mV. Other operations and parameters are the same as those in Example 1.

[0091] The results show that in Step 3, the leaching rate of thallium is 97.89%, the concentration of thallium in the leaching solution is 51.62 g / L, and the free acid concentration in the leaching solution is 30 g / L. Although no sulfur is produced at the anode, during the subsequent cyclone electrowinning process, a large number of bubbles are generated at the anode. After the electrowinning is completed, the mass of thallium on the cathode is only 2.56 g, and the corresponding current efficiency is only 9.22%. In addition, obvious back dissolution of thallium on the cathode is also found during the experiment.

[0092] Comparative Example 4

[0093] Compared with Example 1, the only difference is that in Step 4, Component A is changed. The experimental groups are as follows:

[0094] Group A: Ascorbic acid is not added to the leaching solution.

[0095] The result is that in Step 4, the yield of metallic thallium after cyclone electrodeposition is only 2.29 g, and the current efficiency is reduced to 8.25%.

[0096] Group B: Resorcinol is used as Component A in equal weight.

[0097] The result is that in Step 4, the yield of metallic thallium after cyclone electrodeposition is only 18.92 g, and the current efficiency is reduced to 68.12%.

[0098] Comparative Example 5

[0099] Compared with Example 1, the only difference is that in Step 4, FES-77 of equal weight is used as Component B. Other operations and parameters are the same as those in Example 1.

[0100] The result is that in Step 4, the yield of metallic thallium after cyclone electrodeposition is only 23.66 g, the current efficiency is 85.21%, the concentration of thallium in the electrolyte is reduced to 4.72 g / L, and the morphology of thallium on the cathode is extremely poor (as shown in Figure 6 a)).

[0101] Comparative Example 6

[0102] Compared with Example 1, the only difference is that in Step 4, the cathode during the cyclone electrodeposition process is replaced with a 304 stainless steel plate. Other operations and parameters are the same as those in Example 1.

[0103] The results were as follows: in Step 4, the output of metallic thallium after vortex electrodeposition was only 4.22 g, the current efficiency was 15.21%, and the concentration of thallium in the electrolyte decreased to 43.55 g / L. In addition, it was also found during the experiment that when a 304 stainless steel plate was used as the cathode, hydrogen evolution at the cathode was severe and it was difficult for thallium to deposit on the plate.

[0104] Comparative Example 7

[0105] Compared with Example 1, the only difference was that in Step 4, the current density during the vortex electrodeposition process was increased to 600 A / m 2 , and the electrolysis time was shortened to 2 h.

[0106] The results were as follows: in Step 4, the output of metallic thallium after vortex electrodeposition was 22.10 g, the current efficiency was 86.22%, and the concentration of thallium in the electrolyte decreased to 7.83 g / L. However, under this current density condition, the morphology of thallium on the cathode was poor (as shown in Figure 6 b)).

[0107] Comparative Example 8

[0108] Compared with Example 1, the only difference was that the vortex electrodeposition in Step 4 was replaced by conventional plate electrowinning.

[0109] The results were as follows: in Step 4, the output of metallic thallium after vortex electrodeposition was 23.40 g, the current efficiency was 84.25%, and the concentration of thallium in the electrolyte decreased to 5.24 g / L. However, under this current density condition, the morphology of thallium on the cathode was poor (as shown in Figure 6 c)).

[0110] In summary, by adopting the synergistic process of reduction alkali leaching treatment—Tl + sulfide precipitation—selective controlled-potential acid leaching—Tl + acid leaching solution and vortex electrodeposition assisted by a composite additive, high-quality metallic thallium can be efficiently recovered in this way.

Claims

1. A method for selectively extracting and preparing metallic thallium from thallium-containing materials, characterized in that, Reduce and alkali-leach thallium-containing materials to obtain Tl + alkali-leachate; subject the Tl + alkali-leachate to sulfide precipitation to obtain Tl2S residue; The Tl2S slag is subjected to controlled-potential oxidative acid leaching to selectively oxidize the negative-valence sulfur therein, obtaining Tl + acid leaching solution; Mix the thallium + acid leaching solution with a composite auxiliary agent and then carry out hydrocyclone electrodeposition treatment to obtain metallic thallium and the hydrocyclone electrodeposition post-liquid; The cathode material in vortex electroplating is aluminum plate, copper plate or titanium plate; the composite auxiliary agent includes component A and component B, wherein component A includes Tl liquefier; component B includes at least one of gallic acid, benzylidene acetone, benzophenone, bone glue, polyether, FES-993, quaternary ammonium salt; the current density of vortex electroplating is 50-500 A / m 2 .

2. The method for selectively extracting and preparing metallic thallium from thallium-containing materials according to claim 1, wherein The thallium-containing material is a waste material including at least one of liquid and solid containing Tl + and / or Tl 3+ .

3. The method for selectively extracting and preparing metallic thallium from thallium-containing materials as claimed in claim 1, characterized in that, The thallium-containing material is subjected to reduction alkali leaching in an alkali solution containing a reducing agent; Preferably, the reducing agent includes at least one of sodium borohydride and ascorbic acid; Preferably, the dosage of the reducing agent is 1 to 10 times the theoretical dosage for reducing trivalent thallium in the thallium-containing material to monovalent thallium; Preferably, the alkali solution is an aqueous solution of at least one of alkali metal hydroxides and carbonates; Preferably, the pH in the reduction alkali leaching stage is 7 to 12; Preferably, the reduction leaching temperature is 50 to 100 °C, and the time for reduction alkali leaching is 1 to 2 h.

4. The method for selectively extracting and preparing metallic thallium from thallium-containing materials as claimed in claim 1, wherein Mix the Tl + alkali leaching solution with a water-soluble sulfide for sulfide precipitation to obtain a Tl2S residue; Preferably, the water-soluble sulfide includes at least one of sodium sulfide, potassium sulfide, and ammonium sulfide; Preferably, the amount of water-soluble sulfide used is 1 to 3 times the theoretical amount required to convert thallium in the + alkali leaching solution into Tl2S; Preferably, the temperature in the sulfide precipitation stage is 25 to 85 °C; Preferably, the sulfide precipitation reaction time is 0.5 to 2 h.

5. The method for selectively extracting and preparing metallic thallium from thallium-containing materials as claimed in claim 1, wherein The acid in the controlled-potential oxidation acid leaching process is a sulfuric acid solution; Preferably, the oxidation potential in the controlled-potential oxidation acid leaching process is 400 to 500 mV; Preferably, the temperature in the controlled-potential oxidation acid leaching process is 30 to 90 °C; Preferably, the time for the controlled-potential oxidation acid leaching process is 0.5 to 4 h.

6. The method for selectively extracting and preparing metallic thallium from thallium-containing materials as claimed in claim 1, characterized in that, Tl + Tl in the acid leaching solution + The concentration is less than 80 g / L; the concentration of free acid is less than 100 g / L, and further preferably less than 30 g / L.

7. The method for selectively extracting and preparing metallic thallium from thallium-containing materials according to claim 1, characterized in that, The anode in the cyclone electrodeposition is a titanium-based ruthenium-iridium rod or a graphite rod.

8. The method for selectively extracting and preparing metallic thallium from thallium-containing materials according to claim 1, wherein, In the composite auxiliary agent, component A includes at least one of ascorbic acid, gallic acid, and tartaric acid; The component A described above is Tl + 0.9 to 2 times, preferably 1 to 1.5 times, the molar amount of Tl in the acid leaching solution; The concentration of component B in the solution system of the cyclone electrodeposition is 0.01 to 2 g / L, and further preferably 0.1 to 1 g / L.

9. The method for selectively extracting and preparing metallic thallium from thallium-containing materials according to claim 1, wherein The post-cyclone electrodeposition solution is subjected to microcurrent-assisted replacement for thallium removal, and the steps are as follows: The post-cyclone electrodeposition solution is introduced into a microcurrent-assisted replacement tank; a zinc plate is used as the working electrode, an inert material is used as the counter electrode, and a current is applied to the surface of the zinc plate for replacement to obtain sponge thallium and the post-replacement solution.

10. The method for selectively extracting and preparing metallic thallium from thallium-containing materials as described in claim 9, wherein The microcurrent is one of unidirectional or pulsed currents; the microcurrent density is 10 to 50 A / m 2 ; the microcurrent replacement temperature is 25 to 50 °C.

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