Method for extracting tantalum and niobium from tungsten tailings

Abstract

The invention provides a method for extracting tantalum and niobium from tungsten tailings, and belongs to the technical field of rare metal recovery, and the method comprises the following steps: adding a reducing agent, a fluxing agent, a covering agent and a slag former into the tungsten tailings, and heating and smelting under the protection of inert gas to obtain an alloy ingot; the alloy ingot is crushed and then leached through hydrochloric acid, and acid leaching residues containing tantalum, niobium and tungsten are obtained; the acid leaching residues are subjected to alkaline leaching, and alkaline leaching residues containing tantalum and niobium are obtained; the alkaline leaching residues are leached through hydrofluoric acid and sulfuric acid, and decomposition liquid is obtained; and extracting and separating the decomposition liquid to obtain tantalum hydroxide and niobium hydroxide, and respectively calcining the tantalum hydroxide and niobium hydroxide to obtain tantalum oxide and niobium oxide. The method is simple in technological process, convenient to operate, low in cost and suitable for industrial production, tantalum and niobium can be efficiently extracted from tungsten tailings, and good economic benefits and environmental benefits are achieved.

Description

Technical Field

[0001] This invention relates to the field of rare metal recycling technology, specifically to a method for extracting tantalum and niobium from tungsten tailings. Background Technology

[0002] Tungsten tailings, a typical solid waste generated during the tungsten smelting process, are rich in various valuable metal resources. Its typical components include 0.4%–3% tantalum (Ta), 0.2%–0.1.2% niobium (Nb), and other strategic rare metals, along with 15%–18% iron (Fe), 5%–9% titanium (Ti), 2%–4% cobalt (Co), 0.8%–2% nickel (Ni), and 2%–3% tungsten (W), accounting for approximately 30% of the total metal content. The remainder consists of gangue minerals such as silicon, calcium, and aluminum. Direct landfilling of this type of tailings without treatment not only leads to resource waste due to the stable nature of tantalum and niobium, but also poses an environmental safety hazard due to the risk of heavy metal leaching. This contradicts the core objectives of solid waste reduction and resource utilization in the current "zero-waste city" construction, as well as the comprehensive conservation strategy.

[0003] Currently, the extraction of valuable metals tantalum and niobium from tungsten tailings mainly focuses on hydrometallurgical methods. Patent CN201711147174.0 proposes a method combining mineral processing and hydrometallurgy for recovery. This involves using centrifugal beneficiation to separate tungsten from tantalum and niobium, followed by hydrofluoric acid leaching for purification. However, due to the complex composition of the tungsten tailings used in this invention, containing large amounts of gangue minerals and other metals, direct hydrometallurgical leaching suffers from low leaching efficiency, high reagent consumption, and separation difficulties, resulting in high extraction costs and unsatisfactory results.

[0004] Patent CN202011367492.X discloses a method for enriching and purifying low-grade polymetallic pyrometallurgical slag containing tantalum and niobium. The process mainly includes steps such as crushing of tantalum and niobium slag, acid leaching, high-pressure alkaline leaching, hydrofluoric acid decomposition, tantalum and niobium extraction, preparation of tantalum hydroxide and niobium hydroxide, and high-purity tantalum and niobium oxide. It can efficiently enrich and purify valuable metals such as tantalum, niobium, tungsten, and tin from low-grade polymetallic pyrometallurgical slag containing tantalum and niobium. However, the object of its treatment is secondary slag generated during the smelting process, not primary tungsten tailings. Summary of the Invention

[0005] In view of the technical problems existing in the background art, this application provides a method for extracting tantalum and niobium from tungsten tailings, aiming to solve the technical problems of low leaching efficiency, large reagent consumption and difficult separation of tantalum and niobium from tungsten tailings.

[0006] In a first aspect, embodiments of this application provide a method for extracting tantalum and niobium from tungsten tailings, comprising the following steps: A reducing agent, flux, covering agent and slag-forming agent are added to tungsten tailings, and after heating and smelting under inert gas protection, an alloy ingot is obtained. The alloy ingot was crushed and then leached with hydrochloric acid to obtain an acid leaching residue containing tantalum, niobium, and tungsten. The acid leaching residue is subjected to alkaline leaching to obtain alkaline leaching residue containing tantalum and niobium; The alkaline leaching residue was leached with hydrofluoric acid and sulfuric acid to obtain a decomposition solution; The decomposition solution was extracted and separated to obtain tantalum hydroxide and niobium hydroxide, which were then calcined to obtain tantalum oxide and niobium oxide, respectively.

[0007] In some embodiments, the mass ratio of reducing agent, co-solvent, covering agent and slag-forming agent to tungsten tailings is (12~25):(5~10):(8~10):(10~20):100; The reducing agent includes at least one of elemental sodium, elemental calcium, and elemental carbon; Fluxes include at least one of borax, sodium carbonate, calcium fluoride, and sodium fluoride; Covering agents include glass powder; Slag-forming agents include at least one of calcium oxide, sodium carbonate, calcium sulfate, and silicon dioxide.

[0008] In some embodiments, the heating smelting includes a first stage and a second stage, wherein the first stage is to hold at 1600~1700℃ for 1~1.5h and the second stage is to hold at 1700~1800℃ for 1~2h.

[0009] In some embodiments, the concentration of hydrochloric acid in the hydrochloric acid leaching is 3~6 mol / L, the solid-liquid ratio of the alloy ingot to hydrochloric acid is 1:(5~8) g / mL, and the leaching time is 2~3 h.

[0010] In some embodiments, the alkaline leaching uses a sodium hydroxide solution with a concentration of 2-4 mol / L, the solid-liquid ratio of the acid leaching residue to the sodium hydroxide solution is 1:(4-6) g / mL, the alkaline leaching time is 1-3 h, and the temperature is 80-100℃.

[0011] In some embodiments, the alkaline leaching residue is leached with hydrofluoric acid and sulfuric acid, wherein the concentration of hydrofluoric acid is 2-3 mol / L, the concentration of sulfuric acid is 4-6 mol / L, the volume ratio of hydrofluoric acid to sulfuric acid is 1:(2-3), and the solid-liquid ratio of the alkaline leaching residue and the mixture of hydrofluoric acid and sulfuric acid is 1:(3-5) g / mL.

[0012] In some embodiments, the extraction and separation steps are as follows: The decomposition solution was extracted with an extractant to obtain an organic phase containing tantalum and niobium; The organic phase was back-extracted with dilute sulfuric acid to obtain a niobium back-extraction solution, and then the organic phase was back-extracted with deionized water to obtain a tantalum back-extraction solution. Ammonia gas was introduced into the niobium back-extraction solution and the tantalum back-extraction solution, respectively, to generate niobium hydroxide and tantalum hydroxide precipitates. Niobium hydroxide and tantalum hydroxide precipitates were calcined separately to obtain niobium oxide and tantalum oxide.

[0013] In some embodiments, the extractant includes at least one of MIBK, TBP, and 2-octanol, the volume ratio of the decomposition liquid to the extractant is 1:(1~5), and the extraction time is 5~15 min.

[0014] In some embodiments, ammonia gas is introduced into the niobium back-extraction solution and the tantalum back-extraction solution respectively until the pH of the niobium back-extraction solution and the tantalum back-extraction solution reaches 8 and no more precipitate is formed, at which point the ammonia gas is stopped and the reaction is stirred.

[0015] In some embodiments, the calcination temperature is 600°C, and the calcination is carried out in an air atmosphere for 2-3 hours.

[0016] The advantages of this application, which differ from existing technical solutions, include: 1. This invention innovatively uses a specific combination of reducing agent, fluxing agent, covering agent and slag-forming agent for pyrometallurgical processing, which can effectively reduce the smelting temperature (100~200℃ lower than the traditional process) and reduce energy consumption. At the same time, it achieves gradient reduction and aggregation of metal through segmented heating, and improves the recovery rate of Ta and Nb, which is significantly improved compared with the existing technology.

[0017] 2. During the leaching process, by rationally controlling parameters such as the concentration, temperature, and time of hydrochloric acid, alkaline solution, hydrofluoric acid, and sulfuric acid, efficient separation of metals such as cobalt, nickel, iron, and tungsten from tantalum and niobium was achieved, reducing the impact of impurities on subsequent extraction.

[0018] 3. By utilizing multiple extraction solvents and precisely adjusting the acidity, selective extraction and separation of tantalum and niobium were achieved, improving extraction efficiency and purity. During the back-extraction process, dilute sulfuric acid and water were used to back-extract niobium and tantalum separately, further enhancing the separation effect.

[0019] 4. The entire process of this invention is simple, easy to operate, and low in cost, making it suitable for industrial production. It can efficiently extract tantalum and niobium from tungsten tailings, resulting in good economic and environmental benefits.

[0020] The above description is merely an overview of the technical solution of this application. In order to better understand the technical means of this application and to implement it in accordance with the contents of the specification, and to make the above and other objects, features and advantages of this application more obvious and understandable, specific embodiments of this application are given below. Detailed Implementation

[0021] The embodiments of the technical solution of this application will be described in detail below. The following embodiments are only used to illustrate the technical solution of this application more clearly, and are therefore only examples, and should not be used to limit the scope of protection of this application.

[0022] A method for extracting tantalum and niobium from tungsten tailings includes the following steps: (1) Alloy ingots are obtained by pyrometallurgical refining: Tungsten tailings contain approximately 4% tantalum, 1% niobium, and other metals such as iron, titanium, cobalt, and nickel, with a total metal content of about 30%. The tungsten tailings are pretreated to remove impurities, and then a reducing agent, flux, covering agent, and slag-forming agent are added.

[0023] The reducing agent includes at least one of elemental sodium, elemental calcium, and elemental carbon. Fluxes include at least one of borax, sodium carbonate, calcium fluoride, and sodium fluoride; The covering agent includes glass powder and also requires inert gas protection; The slag-forming agent is at least one of calcium oxide, sodium carbonate, calcium sulfate, and silicon dioxide.

[0024] In the technical solution of this application embodiment, the reducing agent is used to reduce the metal oxides (such as Ta2O5, Nb2O5, FeO, etc.) in the tungsten tailings, converting them into elemental metals that enter the alloy ingot. The chemical equation is as follows: Ta₂O₅ + 5C = 2Ta + 5CO↑ Nb₂O₅ + 5Ca = 2Nb + 5CaO Fluxes are used to lower the melting temperature during the smelting process, promote the separation of slag from the metal phase, and improve smelting efficiency.

[0025] Covering agents are used to coat the surface of molten material, isolating it from air, preventing the oxidation of elemental metals, and reducing heat loss. Glass powder, mainly composed of SiO2 and Na2O, forms a viscous liquid layer at high temperatures, physically isolating it from air.

[0026] Slag-forming agents are used to react with impurity oxides (such as SiO2, Al2O3, P2O5, etc.) generated during the smelting process to form low-melting-point slag, which is easy to separate from the alloy phase.

[0027] The amounts of each reagent added are calculated based on the mass of the tungsten tailings slag. The amount of reducing agent added is 12%–25% of the tungsten tailings slag mass, the amount of flux added is 5%–10%, the amount of covering agent added is 8%–10%, and the amount of slagging agent added is 10%–20%. After the above materials are mixed evenly, they are added to an induction furnace for pyrometallurgical smelting. The temperature of the induction furnace is controlled at 1600–1900℃, and the smelting is carried out in stages. The total reaction time is 2–4 hours to ensure that the materials react fully and obtain alloy ingots. The main components of this alloy ingot are iron, cobalt, nickel, tungsten, tantalum, niobium, and other metals.

[0028] The segmented heating process and reaction principle are as follows: In the first stage (1600~1700℃), low-melting-point metal oxides are preferentially reduced to avoid excessive decomposition of high-melting-point Ta and Nb oxides; in the second stage (1700~1800℃), alloy melting and slag flow are promoted to improve separation efficiency.

[0029] In the first stage, the reducing agent (such as C, Na, Ca) undergoes a preliminary reduction reaction with the metal oxides in the tungsten tailings, generating elemental metals (Ta, Nb, Fe, Co, Ni, etc.). Fluxes (such as borax and sodium carbonate) begin to decompose and lower the system's melting point. Slag-forming agents (such as CaO and SiO2) react with impurity oxides (such as Al2O3 and SiO2) to form a low-melting-point slag phase. A preliminary metal solid solution (containing Ta, Nb, Fe, etc.) and liquid slag are formed, and the elemental metals begin to accumulate but are not completely melted.

[0030] In the second stage, at high temperature, the metal solid solution completely melts to form an alloy ingot, and unreacted metal oxides are further reduced. The slag-forming agent reacts fully with impurities, reducing slag viscosity and completely separating it from the alloy phase. The covering agent (glass powder) melts to form a dense protective layer, preventing alloy oxidation. This process results in a high-density alloy ingot (containing elemental metals such as Ta, Nb, Fe, Co, Ni, and W) and a highly fluid slag. The Ta and Nb enrichment rates in the alloy ingot exceed 90%, providing high-grade raw materials for subsequent wet separation.

[0031] After pyrometallurgical smelting, the alloy phase (mainly containing metals such as Fe, Co, Ni, W, Ta, and Nb) and the slag phase (mainly containing slag generated by the reaction of slag-forming agents, such as CaSiO3 and NaAlO2) are separated due to differences in density and physical state. After smelting, heating is stopped, and the materials in the furnace are allowed to cool to a solid state. The upper slag phase is then physically separated from the lower alloy ingot by crushing the furnace body or by decantation. The alloy ingot is crushed and ground before entering the subsequent wet leaching step, while the slag phase can be further treated or landfilled as industrial solid waste.

[0032] (2) Wet separation stage: multi-metal gradient leaching 1. Alloy powder preparation: After the alloy ingot is cooled, it is initially crushed by a crusher and then ground into powder by a ball mill. The particle size of the powder is controlled at 200~325 mesh for subsequent leaching operations.

[0033] 2. Leaching of metals such as cobalt, nickel, and iron with hydrochloric acid. Alloy powder is leached in hydrochloric acid solution at a concentration of 3-6 mol / L, with a solid-liquid ratio of 1:(5-8) g / mL. Under stirring conditions, the leaching temperature is controlled at 60-80℃, and the leaching time is 2-3 hours. During hydrochloric acid leaching, metals such as cobalt, nickel, and iron react with the hydrochloric acid to form soluble salts, which enter the solution. Metals such as tantalum, niobium, and tungsten remain relatively stable and are retained in the solid residue. After leaching, solid-liquid separation is performed to obtain hydrochloric acid leachate and acid leaching residue. Cobalt, nickel, and iron in the hydrochloric acid leachate can be recovered separately using existing publicly available methods.

[0034] 3. Alkali soaking The acid-leached residue was added to a sodium hydroxide solution for alkaline leaching. The concentration of sodium hydroxide was 2-4 mol / L, and the solid-liquid ratio of the acid-leached residue to the sodium hydroxide solution was 1:(4-6) g / mL. Under heating and stirring conditions, the leaching temperature was controlled at 80-100℃, and the leaching time was 1-3 hours. Tungsten reacts with the alkaline solution to form soluble tungstate, which enters the solution, while tantalum and niobium do not react with the alkaline solution and remain in the solid. After alkaline leaching, solid-liquid separation was performed to obtain alkaline leaching solution and alkaline leaching residue.

[0035] 4. Leaching of tantalum and niobium with hydrofluoric acid and sulfuric acid The alkaline leaching residue is added to a mixed solution of hydrofluoric acid and sulfuric acid for leaching. The concentration of hydrofluoric acid is 2-3 mol / L, and the concentration of sulfuric acid is 4-6 mol / L, with a volume ratio of hydrofluoric acid to sulfuric acid of 1:2 to 1:3. The solid-liquid ratio of the alkaline leaching residue to the mixed solution is 1:(3-5) g / mL. Under stirring conditions, the leaching temperature is controlled at 50-80℃, and the leaching time is 2-3 hours. Tantalum and niobium react with hydrofluoric acid and sulfuric acid to form soluble fluorotantalic acid and fluoroniobic acid, which enter the solution, resulting in a decomposition solution and a decomposition residue. The decomposition residue mainly consists of unreacted impurities such as silicates.

[0036] (3) Extraction and separation of tantalum and niobium The decomposed solution was transferred to an extraction apparatus. The extractant included at least one of MIBK, TBP, and 2-octanol, with a reaction ratio of 1:1 to 1:5 and an extraction time of 5 to 15 minutes. Niobium was back-extracted using dilute sulfuric acid, and tantalum was back-extracted using deionized water. By adding different back-extracting agents, tantalum and niobium were separated and enriched. Industrial ammonia gas was then introduced into the niobium and tantalum back-extracting solutions, respectively, controlling the final pH of the reaction system to 8. Gas introduction was stopped when no more precipitates formed. During gas introduction, stirring was maintained to ensure complete reaction between the ammonia gas and the solution. Tantalum and niobium formed white precipitates Ta(OH)5 and Nb(OH)5, respectively. After gas introduction was completed, solid-liquid separation was performed to obtain hydroxide precipitates of tantalum and niobium.

[0037] The obtained Ta(OH)5 and Nb(OH)5 precipitates were placed separately in a calcining furnace and calcined at 600℃ for 2-3 hours, with air circulation maintained during the calcination process. After calcination, Ta(OH)5 and Nb(OH)5 decomposed to form Ta2O5 and Nb2O5, respectively. After cooling, high-purity tantalum oxide and niobium oxide were obtained.

[0038] The following are some specific embodiments. It should be noted that the embodiments described below are exemplary and are only used to explain this application, and should not be construed as limiting this application. Where specific techniques or conditions are not specified in the embodiments, they shall be performed in accordance with the techniques or conditions described in the literature in this field or according to the product instructions. Reagents or instruments whose manufacturers are not specified are all conventional products that can be obtained commercially.

[0039] I. Preparation Method Example 1 (1) Pyrometallurgical stage 1000g of pretreated tungsten tailings (containing 4% Ta, 1% Nb, and 30% total metal content) was weighed and added according to the following mass ratios: reducing agent (15% sodium), flux (8% borax), covering agent (9% glass powder, with argon gas protection), and slagging agent (15% calcium oxide). The mixture was fed into an induction furnace and heated in stages: first, it was held at 1600℃ for 1 hour, then heated to 1700℃ and held for 2 hours, for a total smelting time of 3 hours, to obtain an alloy ingot. Testing showed that the Ta and Nb enrichment rates in the alloy ingot reached 96.2% and 94.8%, respectively, and the total Fe / Co / Ni content was 65%, forming a stable metallic solid solution.

[0040] (2) Wet separation stage ① Alloy powder preparation: After crushing the alloy ingot, ball mill it to 200 mesh, and the powder passing rate is ≥95%.

[0041] ② Hydrochloric acid leaching: Take 200g of alloy powder, add 1200mL of 5mol / L hydrochloric acid (solid-liquid ratio 1:6), and leach at 70℃ with stirring for 2.5 hours. After solid-liquid separation, Fe in the leaching solution... 2+ / Co 2+ / Ni 2+ The total concentration reached 85 g / L, and the leaching rate was >95%; the retention rate of Ta / Nb / W in the leaching residue was >98%.

[0042] ③ Alkaline leaching for tungsten extraction: The leaching residue is mixed with 3 mol / L sodium hydroxide solution at a solid-liquid ratio of 1:5 and leached at 90℃ with stirring for 2 hours. WO4 in the alkaline leaching solution... 2- The concentration reached 32 g / L, and the tungsten leaching rate was 92%; the alkaline leaching residue contained 24.7% (Ta) and 7.1% (Nb) (as metals).

[0043] ④ Fluorosulfuric acid leaching: The alkaline leaching residue was mixed with a mixed acid (HF 2 mol / L, H₂SO₄ 5 mol / L, volume ratio 1:2.5) at a solid-liquid ratio of 1:4, and leached at 65℃ for 2.5 hours. The TaF₇₇ in the decomposed solution... 2- NbF7 2- The concentrations were 18 g / L and 5 g / L, respectively, and the leaching rates were both >93%.

[0044] (3) Extraction, separation and calcination Using 20% ​​MIBK as the extractant (comparison: 1:2), extraction for 10 minutes resulted in a Ta / Nb partition ratio of 20:1 in the organic phase. In the back-extraction process, niobium was first back-extracted with dilute sulfuric acid (0.5 mol / L) (back-extraction rate 96%), followed by tantalum back-extraction with deionized water (back-extraction rate 98%). The back-extraction solution was purged with ammonia to adjust the pH to 8, yielding white precipitates Ta(OH)₅ and Nb(OH)₅. These precipitates were calcined at 600℃ for 2.5 hours to obtain Ta₂O₅ (purity 99.6%, recovery rate 88%) and Nb₂O₅ (purity 99.2%, recovery rate 85%).

[0045] Example 2 (1) Pyrometallurgical stage 1000g of tungsten tailings from the same batch were used. The reducing agent consisted of 20% elemental carbon, the flux was 7% sodium carbonate, the covering agent was 9% glass powder (under argon protection), and the slagging agent was 18% silica. The medium-frequency furnace smelting process was as follows: holding at 1700℃ for 1.5 hours, then heating to 1800℃ and holding for 1 hour, for a total time of 2.5 hours. The Ta / Nb metal recovery rates in the alloy ingot were 93% and 90%, respectively. Fe / Co / Ni formed a continuous solid solution phase, facilitating subsequent selective leaching.

[0046] (2) Wet separation stage ① Alloy powder preparation: crush and grind to 325 mesh, powder pass rate ≥98%.

[0047] ② Hydrochloric acid leaching: 200g of alloy powder was mixed with 1400mL of 4mol / L hydrochloric acid (solid-liquid ratio 1:7) and leached at 65℃ for 2 hours. The total leaching rate of Fe / Co / Ni in the leachate was >96%, and the Ta / Nb grade in the leaching residue was increased to 13% (based on metal content).

[0048] ③ Alkaline leaching of tungsten: The leaching residue is mixed with 3.5 mol / L sodium hydroxide solution at a solid-liquid ratio of 1:5.5 and leached at 95℃ for 1.5 hours. The tungsten leaching rate is 93%, and the purity of Ta / Nb in the alkaline leaching residue reaches 15%.

[0049] ④ Fluorine-sulfuric acid leaching: The alkaline leaching residue was mixed with a mixed acid (HF 2 mol / L, H2SO4 5.5 mol / L, volume ratio 1:2.8) at a solid-liquid ratio of 1:4.5 and leached at 70℃ for 2.2 hours. The leaching rates of Ta / Nb were both >94%. The impurity ions in the decomposed liquid (Si) 4+ Concentration < 1 g / L.

[0050] (3) Extraction, separation and calcination Using 20% ​​TBP as the extractant (comparison: 1:3), extraction was performed for 12 minutes to achieve efficient separation of Ta / Nb from residual impurities. During the back-extraction process, dilute sulfuric acid (1.5 mol / L) was used as the niobium back-extraction agent, and deionized water was used as the tantalum back-extraction agent, with back-extraction rates >97% for both. The precipitation process was controlled to an endpoint pH of 8. After calcination, Ta₂O₅ (purity 99.7%, recovery 89%) and Nb₂O₅ (purity 99.3%, recovery 86%) were obtained, all meeting the standards for electronic-grade oxides.

[0051] Example 3 (1) Pyrometallurgical stage 1000g of tungsten tailings (composition as before) was weighed. The reducing agent was 12% calcium + 5% carbon powder, the flux was 5% calcium fluoride + 3% borax, the covering agent was 8% glass powder (under nitrogen protection), and the slagging agent was 12% sodium carbonate + 8% calcium oxide. The medium-frequency furnace was used for segmented heating: 1650℃ for 1.5 hours, 1800℃ for 1.5 hours, for a total reaction time of 3 hours. The Ta / Nb recovery rates in the alloy ingot were 94% and 91%, respectively, and the Fe / Co / Ni content was 68%, forming a homogeneous alloy phase.

[0052] (2) Wet separation stage ① Alloy powder preparation: crush to 250 mesh (pass rate ≥ 96%).

[0053] ② Hydrochloric acid leaching: 200g of alloy powder was added to 1600mL of 3mol / L hydrochloric acid (solid-liquid ratio 1:8), and leached at 80℃ for 3 hours. The leaching rate of Fe / Co / Ni was >97%, and the Ta / Nb content in the leaching residue was increased to 14%.

[0054] ③Alkaline leaching of tungsten: The leaching residue is mixed with 4 mol / L sodium hydroxide solution at a solid-liquid ratio of 1:6 and leached at 100℃ for 1 hour. The tungsten leaching rate is 94%, and the purity of Ta / Nb in the alkaline leaching residue reaches 16%.

[0055] ④ Fluorine-sulfuric acid leaching: The alkaline leaching residue was mixed with a mixed acid (HF 3mol / L, H2SO4 6mol / L, volume ratio 1:3) at a solid-liquid ratio of 1:3 and leached at 80℃ for 2 hours. The leaching rates of Ta and Nb were both >95%, and the concentrations of Ta and Nb in the decomposition solution were 20g / L and 6g / L, respectively.

[0056] (3) Extraction, separation and calcination Using 20% ​​MIBK + 5% sec-octanol (volume ratio 4:1) as the extractant (volume ratio 1:4), extraction was performed for 15 minutes, resulting in a Ta / Nb partition ratio >25:1. During the back-extraction process, the niobium back-extraction rate was 98%, and the tantalum back-extraction rate was 99%. After precipitation, calcination yielded Ta₂O₅ (purity 99.8%, recovery 90%) and Nb₂O₅ (purity 99.5%, recovery 87%), meeting the requirements for ultra-high purity electronic materials.

[0057] All three embodiments, through different reagent combinations and optimized process parameters, achieved efficient extraction and separation of tantalum and niobium, with key indicators significantly superior to existing technologies. Embodiment 3 demonstrated outstanding performance in the preparation of ultra-high purity products, verifying the flexibility and adaptability of the process and providing diverse technical solutions for the industrial recovery of tantalum and niobium resources from tungsten tailings.

[0058] It should be noted that this application is not limited to the above-described embodiments. The above embodiments are merely examples, and any embodiments with the same structure and effect as the technical concept within the scope of this application are included in the technical scope of this application. Furthermore, various modifications that can be conceived by those skilled in the art to the embodiments, and other ways of constructing by combining some of the constituent elements of the embodiments, without departing from the spirit of this application, are also included in the scope of this application.

Claims

1. A method for extracting tantalum and niobium from tungsten tailings, characterized in that, Includes the following steps: A reducing agent, flux, covering agent and slag-forming agent are added to tungsten tailings, and after heating and smelting under inert gas protection, an alloy ingot is obtained. The alloy ingot was crushed and then leached with hydrochloric acid to obtain an acid leaching residue containing tantalum, niobium, and tungsten. The acid leaching residue is subjected to alkaline leaching to obtain alkaline leaching residue containing tantalum and niobium; The alkaline leaching residue was leached with hydrofluoric acid and sulfuric acid to obtain a decomposition solution; The decomposition solution was extracted and separated to obtain tantalum hydroxide and niobium hydroxide, which were then calcined to obtain tantalum oxide and niobium oxide, respectively.

2. The method for extracting tantalum and niobium from tungsten tailings according to claim 1, characterized in that, The mass ratio of the reducing agent, co-solvent, covering agent, and slag-forming agent to the tungsten tailings is (12~25):(5~10):(8~10):(10~20):100; The reducing agent includes at least one of elemental sodium, elemental calcium, and elemental carbon; The flux includes at least one of borax, sodium carbonate, calcium fluoride, and sodium fluoride; The covering agent includes glass powder; The slag-forming agent includes at least one of calcium oxide, sodium carbonate, calcium sulfate, and silicon dioxide.

3. The method for extracting tantalum and niobium from tungsten tailings according to claim 1, characterized in that, The heating and smelting process includes a first stage and a second stage. The first stage involves holding the temperature at 1600-1700℃ for 1-1.5 hours, and the second stage involves holding the temperature at 1700-1800℃ for 1-2 hours.

4. The method for extracting tantalum and niobium from tungsten tailings according to claim 1, characterized in that, The hydrochloric acid concentration in the leaching process is 3-6 mol / L, the solid-liquid ratio of the alloy ingot to hydrochloric acid is 1:(5-8) g / mL, and the leaching time is 2-3 h.

5. The method for extracting tantalum and niobium from tungsten tailings according to claim 1, characterized in that, The alkaline leaching uses a sodium hydroxide solution with a concentration of 2~4 mol / L, the solid-liquid ratio of the acid leaching residue to the sodium hydroxide solution is 1:(4~6) g / mL, the alkaline leaching time is 1~3 h, and the temperature is 80~100℃.

6. The method for extracting tantalum and niobium from tungsten tailings according to claim 1, characterized in that, The alkaline leaching residue is leached with hydrofluoric acid and sulfuric acid, wherein the concentration of hydrofluoric acid is 2-3 mol / L, the concentration of sulfuric acid is 4-6 mol / L, the volume ratio of hydrofluoric acid to sulfuric acid is 1:(2-3), and the solid-liquid ratio of the mixture of alkaline leaching residue and hydrofluoric acid and sulfuric acid is 1:(3-5) g / mL.

7. The method for extracting tantalum and niobium from tungsten tailings according to claim 1, characterized in that, The extraction and separation steps are as follows: The decomposition solution was extracted with an extractant to obtain an organic phase containing tantalum and niobium; The organic phase was back-extracted with dilute sulfuric acid to obtain a niobium back-extraction solution, and then the organic phase was back-extracted with deionized water to obtain a tantalum back-extraction solution. Ammonia gas is introduced into the niobium back-extraction solution and the tantalum back-extraction solution respectively to generate niobium hydroxide and tantalum hydroxide precipitates; The niobium hydroxide and tantalum hydroxide precipitates were calcined to obtain niobium oxide and tantalum oxide, respectively.

8. The method for extracting tantalum and niobium from tungsten tailings according to claim 1, characterized in that, The extractant includes at least one of MIBK, TBP, and 2-octanol, the volume ratio of the decomposition liquid to the extractant is 1:(1~5), and the extraction time is 5~15 min.

9. The method for extracting tantalum and niobium from tungsten tailings according to claim 1, characterized in that, Ammonia gas is introduced into the niobium and tantalum extraction solutions respectively until the pH of the niobium and tantalum extraction solutions reaches 8 and no more precipitates are formed. The ammonia gas is then introduced and the reaction is stirred.

10. The method for extracting tantalum and niobium from tungsten tailings according to claim 1, characterized in that, The calcination temperature is 600℃, and the calcination is carried out in air for 2-3 hours.