Method for comprehensively recovering valuable metals from tungsten-containing waste residues difficult to treat

Abstract

The invention provides a method for comprehensively recovering valuable metals from tungsten-containing waste residues difficult to treat, and belongs to the field of metal waste recycling. The method comprises the steps that tungsten-containing waste residue slurry, inorganic acid and an oxidizing agent are subjected to an acid leaching cobalt dissolving reaction, and a cobalt salt solution and acid leaching tungsten-rich residues are obtained after solid-liquid separation; carrying out solid-liquid mixing on the acid-leached tungsten-rich slag, a sodium-containing molten salt medicament and water, and roasting to obtain roasted slag; the roasting slag and water are subjected to a heating water leaching reaction, and tungsten-containing water leaching liquid and water leaching slag are obtained after solid-liquid separation; and pulping the water leaching residues, carrying out acid leaching iron dissolving reaction with inorganic acid and an oxidizing agent, and carrying out solid-liquid separation to obtain an iron salt solution and tantalum-niobium enriched residues. The efficient treatment method for the tungsten-containing waste residues difficult to treat comprises the steps of acid leaching of the recycled residues, fused salt roasting of the acid leaching residues, water leaching of the roasting residues and secondary acid leaching of the water leaching residues, efficient extraction of various resources in the recycled residues is achieved, the recycling rate and the treatment cost are both considered, and wastewater discharge is reduced by recycling water resources.

Description

Technical Field

[0001] This invention belongs to the field of metal waste resource utilization technology, and specifically relates to a method for comprehensively recovering valuable metals from difficult-to-treat tungsten-containing waste slag. Background Technology

[0002] With the increasing depletion of global metal mineral reserves, the metal industry is accelerating its shift towards the utilization of recycled resources. The cemented carbide, wear-resistant materials, and tungsten product processing industries generate large quantities of tungsten-containing waste slag during production. This waste slag has a complex composition, typically containing valuable tungsten (W) and cobalt (Co), along with other valuable metals such as iron (Fe), tantalum (Ta), and niobium (Nb). Cobalt, as a strategic metal, is expensive and represents a crucial strategic resource. Therefore, the efficient and economical recovery of valuable metals from this tungsten-containing waste slag not only has significant economic benefits but also plays a vital role in achieving resource recycling and ensuring national strategic resource security.

[0003] However, such waste residues often have stable phase structures due to their complex origins and high-temperature sintering, making them typical difficult-to-treat solid waste resources. Existing technologies typically face a series of technical bottlenecks when processing this type of waste residue. Currently, many companies in the industry specialize in the recycling and processing of waste cemented carbide. Apart from the zinc smelting method for recycling clean cemented carbide, which is a pyrometallurgical process, other recycling methods mainly rely on hydrometallurgy. However, their target metal is primarily tungsten, which has the highest content and the simplest recycling process. This is because waste cemented carbide presents a diverse range of types but small quantities of each individual type, leading recycling companies to mix and process various types of waste cemented carbide together, discarding the resulting 5%-10% tungsten extraction slag. While this controls costs, it allows many other high-value metals to enter the tungsten extraction slag, mainly cobalt, tantalum, and niobium added during cemented carbide production. Furthermore, this slag, being a difficult-to-process waste, is usually sold mixed with other tungsten-containing wastes that are complex in composition and heavily polluting. This waste contains tungsten oxide, tungsten carbide, cemented carbide powder, and cemented carbide recycling tailings, often mixed with coarse iron filings and iron oxides generated from ferrotungsten recycling. Due to issues with the recycling process, the tungsten in the waste cemented carbide forms stable compounds with other metals during the roasting and oxidation process, along with some unoxidized tungsten carbide. This results in a large amount of tungsten (containing 20%-50% of the material) failing to leach completely. Moreover, the slag's complex and diverse elemental composition, uneven distribution, and similar physicochemical properties of some metal elements make it difficult to leach and separate valuable metals using conventional methods.

[0004] Patent CN109628744B discloses a method for recovering tungsten and cobalt from tungsten-containing cemented carbide waste. The method involves: firstly, the tungsten-containing cemented carbide waste undergoes a first-stage slurrying process followed by a first-stage atmospheric pressure acid leaching. After a first solid-liquid separation, cobalt is recovered from the first-stage leaching solution. The first-stage leaching residue is then added to a strong acid for a second-stage slurrying process, followed by a second-stage pressurized acid leaching. After a second solid-liquid separation, the second-stage leaching residue is returned to the alkaline leaching process in tungsten smelting to recover tungsten, and the second-stage leaching solution is returned to the first-stage slurrying process used for tungsten-containing cemented carbide waste. This invention employs a two-stage countercurrent acid leaching process, which can improve the cobalt leaching rate and achieve effective separation of tungsten and cobalt, resulting in high tungsten and cobalt recovery rates and significant economic benefits. However, this patent uses an acid method throughout the entire process to extract cobalt and enrich tungsten, and requires pressurized treatment during the second-stage acid leaching, which undoubtedly places a heavy burden on the acid leaching equipment. Furthermore, the conversion rate of tungsten into a form that can be processed by the alkaline leaching process using this method is not necessarily ideal. Furthermore, this patent only addresses the front-end processing of waste cemented carbide and does not solve the problem of re-extracting the residual high-value metals from the tailings after tungsten-cobalt extraction.

[0005] Therefore, there is an urgent need in this field to develop a new technical method that can achieve efficient, clean, and comprehensive recovery of various valuable metals such as tungsten, cobalt, and iron from difficult-to-treat tungsten-containing waste residues, in order to overcome the aforementioned shortcomings of existing technologies. Summary of the Invention

[0006] Therefore, the present invention aims to provide a method for the comprehensive recovery of valuable metals from difficult-to-process tungsten-containing waste slag, in order to solve at least one of the technical problems in the background art.

[0007] This invention is implemented as follows: This invention provides a method for the comprehensive recovery of valuable metals from difficult-to-treat tungsten-containing waste residue, the method comprising the following steps: S1. The tungsten-containing waste slurry obtained after crushing, screening and pulping pretreatment is subjected to cobalt leaching reaction with inorganic acid, and an oxidant is added to enhance the reaction to achieve selective leaching of cobalt. After solid-liquid separation, cobalt salt solution and acid-leached tungsten-rich slag are obtained. S2. The acid-leached tungsten-rich slag is mixed with sodium-containing molten salt agent and water solid-liquid mixture and then roasted to convert tungsten into soluble tungstate and iron into ferric oxide that is easily soluble in acid, thus obtaining roasted slag. S3. Mix the roasted residue with water and heat it for a water leaching reaction to allow tungsten to enter the solution. After solid-liquid separation, tungsten-containing water leaching solution and water leaching residue are obtained. S4. The water-leached residue is pulped and then subjected to an acid leaching reaction with an inorganic acid. An oxidant is added to enhance the reaction, thereby achieving iron leaching. After solid-liquid separation, an iron salt solution and tantalum-niobium enriched residue are obtained.

[0008] Preferably, the inorganic acid in S1 and S4 is hydrochloric acid and the oxidizing agent is hydrogen peroxide.

[0009] Preferably, in S1, the amount of oxidant added is not less than 15 wt% of the dry weight of the tungsten-containing waste residue; the acid leaching reaction for dissolving cobalt continues for at least 5 hours after the addition of the oxidant.

[0010] Preferably, in S1, the amount of inorganic acid used is not less than 1.5 times the theoretical amount used to dissolve cobalt.

[0011] Preferably, in step S4, the amount of oxidant added is not less than 15 wt% of the dry weight of the water-leached residue; the acid leaching iron dissolution reaction continues for at least 7 hours after the addition of the oxidant.

[0012] Preferably, in S4, the amount of inorganic acid used is not less than 1.5 times the theoretical amount of iron to be dissolved.

[0013] Preferably, in S2, the sodium-containing molten salt agent is sodium nitrate, or a combination of sodium nitrate and at least one of sodium carbonate and sodium hydroxide.

[0014] Preferably, in S2, the amount of sodium-containing molten salt agent is 40wt%~50wt% of the dry weight of the acid-leached tungsten-rich slag; and the amount of water used during solid-liquid mixing is 40wt%~60wt% of the dry weight of the acid-leached tungsten-rich slag.

[0015] Preferably, in step S2, the calcination temperature is 700℃~800℃, and the holding time is not less than 60 minutes.

[0016] Preferably, in S3, the temperature of the water immersion reaction is not lower than 70°C, and the water immersion time is not lower than 1 hour.

[0017] Compared with the prior art, the present invention has the following beneficial effects: 1. This invention provides an efficient treatment method for difficult-to-treat tungsten-containing waste residue, which involves acid leaching of recycled residue, molten salt roasting of acid leaching residue, water leaching of roasting residue, and secondary acid leaching of water leaching residue. This method achieves efficient extraction of various resources from the recycled residue, balances recovery rate and treatment cost, and reduces wastewater discharge by recycling water resources.

[0018] 2. This invention utilizes the different forms in which iron and cobalt exist in the recycling residue and the different priorities of their reactions with acid to extract cobalt by acid leaching, thereby enriching tungsten and controlling the leaching of iron.

[0019] 3. The recovered slag after molten salt roasting in this invention not only realizes the transformation of tungsten to a free state in the slag, but also completely oxidizes the iron in the slag into Fe2O3 and fixes it in the slag phase. This ensures the purity of the tungsten solution after water leaching and facilitates the subsequent tantalum and niobium enrichment by iron dissolution. At the same time, the solid-liquid mixing method can ensure uniform mixing and reduce the amount of molten salt reagent used.

[0020] 4. This invention utilizes the property that tantalum and niobium oxides are stable in general acids and alkalis. Through multiple acid and alkali treatments, most of the cobalt, tungsten, and iron in the slag are dissolved, thus extracting the main metal elements and enriching tantalum and niobium.

[0021] 5. The process of this invention balances recovery rate and treatment cost. The washing water from both water-leached and acid-leached slags can be recycled, saving water consumption and reducing the burden of wastewater treatment. Extensive experimental verification has shown that this process can achieve a stable WO3 leaching rate of over 97% and a cobalt leaching rate of over 95% when treating difficult-to-treat tungsten-containing waste slag, while increasing the percentage content of tantalum and niobium in the slag by 7 to 8 times. Attached Figure Description

[0022] Figure 1 This is a process flow diagram of the method for comprehensively recovering valuable metals from difficult-to-treat tungsten-containing waste slag according to the present invention. Detailed Implementation

[0023] To make the objectives, technical solutions, and advantages of this invention clearer, the invention will be further described in detail below with reference to embodiments. It should be understood that the specific embodiments described herein are merely illustrative and not intended to limit the invention.

[0024] like Figure 1 As shown, a method for the comprehensive recovery of valuable metals from refractory tungsten-containing waste slag includes steps S1 to S4.

[0025] S1. The tungsten-containing waste slurry obtained after crushing, screening and pulping pretreatment is subjected to cobalt leaching reaction with inorganic acid, and an oxidant is added to enhance the reaction to achieve selective leaching of cobalt. After solid-liquid separation, cobalt salt solution and acid-leached tungsten-rich slag are obtained. Specifically: After screening out large particles from the crushed tungsten-containing waste residue, it is mixed with an equal weight of tap water to form a slurry, which is then stirred using a stirring device. A certain amount of inorganic acid (preferably hydrochloric acid, which is more economical) is slowly added to the slurry. After the tungsten-containing waste residue slurry and inorganic acid have fully impregnated and initially reacted, a certain amount of oxidant (preferably industrial hydrogen peroxide) is continuously added to the mixed slurry to enhance cobalt dissolution. The amount of oxidant added is not less than 15% of the dry weight of the tungsten-containing waste residue before acid leaching, and it is added dropwise. After the dropwise addition is completed, the reaction continues for at least 5 hours to ensure complete cobalt leaching. After the reaction, solid-liquid separation is performed. The collected filter residue, after washing, yields acid-leached tungsten-rich slag, which achieves the enrichment of tungsten and iron, and will be further separated. The collected filtrate is the cobalt salt solution, which can be collected and sold or further processed. The washing water from the filter residue can be recycled and returned to the slurry preparation of tungsten-containing waste residue.

[0026] In this step, since cobalt acts as a binder in cemented carbide, it is typically added to tungsten carbide in nanoscale powder form before sintering. Therefore, the cobalt in the tungsten-containing waste slag has good acid solubility. Iron in the tungsten-containing waste slag mainly exists in the form of coarse iron filings (slow acid dissolution rate), Fe3O4 (trans-spinel structure, not easily reacting with acid), and a small amount of iron solid solution (e.g., CoFe2O4, stable properties). Therefore, only the amount of inorganic acid and the reaction time need to be controlled to ensure preferential dissolution of cobalt. In specific implementation, the amount of inorganic acid added depends on the total amount of Co in the reduction slag, and the amount of inorganic acid should not be less than 1.5 times the theoretical amount needed to dissolve cobalt.

[0027] S2. The acid-leached tungsten-rich slag is mixed with a sodium-containing molten salt agent and then roasted to convert tungsten into soluble tungstate and iron into ferric oxide that is easily soluble in acid, thus obtaining roasted slag. Specifically: the acid-leached tungsten-rich slag obtained in step S1 is added to sodium-containing molten salt agent and water and thoroughly mixed in a solid-liquid mixture. After the mixture is completed, the mixture is roasted in a roasting furnace at a temperature of 700℃~800℃ for a holding time of not less than 60 minutes.

[0028] The sodium-containing molten salt agent is made of solid sodium nitrate, or a combination of sodium nitrate with at least one of sodium carbonate and sodium hydroxide; the amount of sodium-containing molten salt agent is 40wt% to 50wt% of the dry weight of the acid-leached tungsten-rich slag; the amount of water used in solid-liquid mixing is 40wt% to 60wt% of the dry weight of the acid-leached tungsten-rich slag.

[0029] After roasting, the tungsten in the slag is converted into a free state, and the main reaction formula is: Fe(WO4). x / Cr(WO4) x +NaNO3→Na2WO4+Fe2O3+Na2CrO4; The remaining iron in the slag is fully oxidized to loose and acid-soluble Fe2O3 by high-temperature aerobic roasting with sodium nitrate, involving the following reaction formula: Fe / Fe3O4 / CoFe2O4+NaNO 3+ O2→Fe2O3+Co2O3.

[0030] S3. Mix the roasted residue with water and heat it for a water leaching reaction to allow tungsten to enter the solution. After solid-liquid separation, tungsten-containing water leaching solution and water leaching residue are obtained. Specifically: The calcined residue obtained in step S2 is poured into a reactor, and tap water is added to form a slurry. The liquid-to-solid ratio can be freely adjusted according to the required tungsten concentration in the leaching solution. Stirring and heating are started for leaching, with the leaching temperature not lower than 70°C, to begin tungsten dissolution. The leaching time is recorded from the moment the reactor temperature reaches the set value, and the leaching time is not less than 1 hour. After leaching, the slurry is separated into solid and liquid components to obtain filtrate and leached residue. The filtrate is a tungsten-containing leaching solution, which can be used to prepare tungstic acid or artificial scheelite depending on the tungsten concentration. The leached residue, with Fe2O3 as the main component, needs to be washed with water and dried for use in subsequent steps. The washing water can be recycled and returned to the leached calcined residue.

[0031] S4. The water-leached residue is pulped and then subjected to an acid leaching reaction with an inorganic acid. An oxidant is added to enhance the reaction, thereby achieving iron leaching. After solid-liquid separation, an iron salt solution and tantalum-niobium enriched residue are obtained.

[0032] Specifically: The water-leached residue obtained in step S3 is mixed with an equal weight of tap water to form a slurry, and the slurry is stirred using a stirring device. At this time, a certain amount of inorganic acid (preferably hydrochloric acid, which is more economical) is slowly added to the slurry. The amount of inorganic acid added depends on the total amount of Fe in the reduction residue (the amount of inorganic acid should not be less than 1.5 times the theoretical amount for iron dissolution). After the material is fully wetted by the acid and undergoes a preliminary reaction, a certain amount of oxidant (preferably industrial hydrogen peroxide) is continuously added to the mixed slurry to enhance iron dissolution. The amount of oxidant added should not be less than 15% of the dry weight of the water-leached residue, and the addition method is dropwise. After the dropwise addition is completed, the reaction continues for at least 7 hours to ensure the iron leaching reaction is complete. After the reaction is completed, solid-liquid separation is performed to obtain filtrate and filter residue. The filtrate is an iron salt solution; the filter residue is a tantalum-niobium enriched residue. After iron dissolution, tantalum and niobium are enriched. After washing with tap water, it can be sold externally or used for subsequent extraction. The washing water can be recycled and returned to the water-leached residue slurry preparation.

[0033] In the following examples and comparative examples, the tungsten-containing waste residue raw material is a difficult-to-treat tungsten-containing waste residue produced by preliminary tungsten recovery of cemented carbide. Its main components are shown in Table 1 below.

[0034] Table 1

[0035] Example 1 A method for the comprehensive recovery of valuable metals from difficult-to-treat tungsten-containing waste slag includes the following steps: S1. Tungsten-containing waste slag raw material is crushed and screened to obtain recycled slag. 100g of the screened recycled slag is weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (approximately 31% by mass concentration, theoretically 1.5 times the amount needed to dissolve cobalt) is added and stirred evenly. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled slag) is slowly added for acid leaching. After reacting for 5 hours, the slurry is filtered, and the filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich slag) are collected separately. The filter residue is washed, dried, and weighed to 78.8g. The cobalt content in the dried filter residue is tested to be 0.94% (C). 钴 The cobalt leaching rate in this step was calculated to be 96.03%. In the formula, The mass of cobalt in the raw materials; The quality of the dried filter residue; This represents the cobalt content in the filter residue.

[0036] S2. The acid-leached tungsten-rich slag obtained in step S1 is mixed evenly with 32g of sodium nitrate (40% of the dry weight of the acid-leached tungsten-rich slag) and 40mL of water. The mixture is then placed in a muffle furnace for oxidation roasting at a temperature of 700℃ and a holding time of 60min. After natural cooling, roasted slag is obtained.

[0037] S3. Take out the calcined residue obtained in step S2, add 150 mL of pure water for water leaching at 70℃ for 60 min. After water leaching, filter the slurry to obtain tungsten-containing water leaching solution and water leaching residue. The water leaching residue is washed, dried, and weighed 30.6 g. The WO3 content in the water leaching residue is tested to be 1.33%, and the WO3 leaching rate in this step is calculated to be 98.73%. In the formula, The mass of WO3 in the raw materials; The quality of the water-impregnated residue after drying; This refers to the WO3 content in the water-leached residue.

[0038] S4. Mix the water-leached residue obtained in step S3 with 30 mL of tap water to form a slurry. Add 114 mL (dissolved iron = 1.5 times the theoretical amount) of industrial hydrochloric acid (mass concentration approximately 31%) and stir until homogeneous. Then slowly add 4.59 g of hydrogen peroxide (15% of the dry weight of the water-leached residue) for acid leaching. After reacting for 7 hours, filter the slurry to obtain the iron salt solution and tantalum-niobium enriched residue. The tantalum-niobium enriched residue is washed, dried, and weighed 10.9 g. After drying, the tantalum-niobium enriched residue contains 5.77% iron (Fe), 4.71% tantalum (Ta), and 4.04% niobium (Nb).

[0039] Example 2 A method for the comprehensive recovery of valuable metals from difficult-to-treat tungsten-containing waste slag includes the following steps: S1. The tungsten-containing waste slag raw material is crushed and screened to obtain recycled slag. 100g of the screened recycled slag is weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (theoretically 1.5 times the amount needed to dissolve cobalt) is added and stirred evenly. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled slag) is slowly added for acid leaching. After reacting for 5 hours, the slurry is filtered, and the filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich slag) are collected separately. After washing and drying, the filter residue weighs 80.2g. The cobalt content in the dried filter residue is tested to be 0.76%, and the cobalt leaching rate in this step is calculated to be 96.73%.

[0040] S2. The acid-leached tungsten-rich slag obtained in step S1 is mixed evenly with 40g of sodium nitrate (50% of the dry weight of the acid-leached tungsten-rich slag) and 40mL of water. The mixture is then placed in a muffle furnace for oxidation roasting at a temperature of 700℃ and a holding time of 60min. After natural cooling, roasted slag is obtained.

[0041] S3. Take out the calcined residue obtained in step S2, add 150 mL of pure water for water leaching at 70℃ for 60 min. After water leaching, filter the slurry to obtain tungsten-containing water leaching solution and water leaching residue. After washing and drying, the water leaching residue weighs 31.5 g. The WO3 content in the water leaching residue is tested to be 0.72%, and the WO3 leaching rate in this step is calculated to be 99.29%.

[0042] The difference between this embodiment and Embodiment 1 is that the amount of reagent used in the molten salt roasting of the acid-leached tungsten-rich slag is increased from 40% to 50% of the dry weight, and only the water leaching step is performed; the remaining conditions are the same as in Embodiment 1. Increasing the amount of molten salt reagent to 50% increases the leaching rate of WO3 to 99.29%. Further increasing the reagent dosage would only increase costs and would be of no practical significance.

[0043] Example 3 A method for the comprehensive recovery of valuable metals from difficult-to-treat tungsten-containing waste slag includes the following steps: S1. The tungsten-containing waste slag raw material is crushed and screened to obtain recycled slag. 100g of the screened recycled slag is weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (theoretically 1.5 times the amount needed to dissolve cobalt) is added and stirred evenly. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled slag) is slowly added for acid leaching. After reacting for 5 hours, the slurry is filtered, and the filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich slag) are collected separately. After washing and drying, the filter residue weighs 77.9g. The cobalt content in the dried filter residue is tested to be 0.87%, and the cobalt leaching rate in this step is calculated to be 96.36%.

[0044] S2. The acid-leached tungsten-rich slag obtained in step S1 is mixed evenly with 16g sodium nitrate + 16g anhydrous sodium carbonate (40% of the dry weight of the acid-leached tungsten-rich slag) and 40mL water. The mixture is then placed in a muffle furnace for oxidation roasting at a temperature of 700℃ and a holding time of 60min. After natural cooling, roasted slag is obtained.

[0045] S3. Take out the calcined residue obtained in step S2, add 150 mL of pure water for water leaching at 70℃ for 60 min. After water leaching, filter the slurry to obtain tungsten-containing water leaching solution and water leaching residue. The water leaching residue is washed, dried, and weighed 34.1 g. The WO3 content in the water leaching residue is 2.02%, and the WO3 leaching rate in this step is calculated to be 97.85%.

[0046] The difference between this embodiment and Embodiment 1 is that the reagent used in the molten salt roasting of the acid-leached tungsten-rich slag is changed from pure sodium nitrate to sodium nitrate:sodium carbonate = 1:1, while the total amount of reagent remains the same. Furthermore, only the water leaching step is performed; the remaining conditions are the same as in Embodiment 1. After changing the roasting reagent to sodium nitrate + sodium carbonate, the leaching rate of WO3 was 97.85%, slightly lower but still within the ideal range.

[0047] Example 4 A method for the comprehensive recovery of valuable metals from difficult-to-treat tungsten-containing waste slag includes the following steps: S1. The tungsten-containing waste slag raw material is crushed and screened to obtain recycled slag. 100g of the screened recycled slag is weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (approximately 31% by mass concentration, theoretically 1.5 times the amount needed to dissolve cobalt) is added and stirred evenly. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled slag) is slowly added for acid leaching. After reacting for 5 hours, the slurry is filtered, and the filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich slag) are collected separately. After washing and drying, the filter residue weighs 78.4g. The cobalt content in the dried filter residue is tested to be 0.90%, and the cobalt leaching rate in this step is calculated to be 96.24%.

[0048] S2. The acid-leached tungsten-rich slag obtained in step S1 is mixed evenly with 16g sodium nitrate + 16g sodium hydroxide (40% of the dry weight of the acid-leached tungsten-rich slag) and 40mL water. The mixture is then placed in a muffle furnace for oxidation roasting at a temperature of 700℃ and a holding time of 60min. After natural cooling, roasted slag is obtained.

[0049] S3. Take out the calcined residue obtained in step S2, add 150 mL of pure water for water leaching at 70℃ for 60 min. After water leaching, filter the slurry to obtain tungsten-containing water leaching solution and water leaching residue. The water leaching residue is washed, dried, and weighed to 32.3 g. The WO3 content in the water leaching residue is tested to be 1.30%, and the WO3 leaching rate in this step is calculated to be 98.69%.

[0050] The difference between this embodiment and Embodiment 1 is that the reagent used in the molten salt roasting of the acid-leached tungsten-rich slag is changed from pure sodium nitrate to sodium nitrate:sodium hydroxide = 1:1, while the total amount of reagent remains the same. Furthermore, only the water leaching step is performed; the remaining conditions are the same as in Embodiment 1. After changing the roasting reagent to sodium nitrate + sodium hydroxide, the leaching rate of WO3 is 98.69%.

[0051] Example 5 A method for the comprehensive recovery of valuable metals from difficult-to-treat tungsten-containing waste slag includes the following steps: S1. The tungsten-containing waste residue is crushed and screened to obtain recycled residue. 100g of the screened recycled residue is weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (theoretically 1.5 times the amount needed to dissolve cobalt) is added and stirred evenly. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled residue) is slowly added for acid leaching. After reacting for 5 hours, the slurry is filtered. The filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich residue) are collected separately. The filter residue is washed, dried, and weighed to 79.8g.

[0052] S2. The acid-leached tungsten-rich slag obtained in step S1 is mixed evenly with 32g of sodium nitrate (40% of the dry weight of the acid-leached tungsten-rich slag) and 40mL of water. The mixture is then placed in a muffle furnace for oxidation roasting at a temperature of 800℃ and a holding time of 60min. After natural cooling, roasted slag is obtained.

[0053] S3. Take out the calcined residue obtained in step S2, add 150 mL of pure water for water leaching at 70℃ for 60 min. After water leaching, filter the slurry to obtain tungsten-containing water leaching solution and water leaching residue. After washing and drying, the water leaching residue weighs 30.0 g. The WO3 content in the water leaching residue is tested to be 0.99%, and the WO3 leaching rate in this step is calculated to be 99.07%.

[0054] S4. Mix the water-leached residue obtained in step S3 with 30 mL of tap water to form a slurry. Add 114 mL of industrial hydrochloric acid (1.5 times the theoretical amount for iron dissolution) and stir until homogeneous. Then slowly add 4.5 g of hydrogen peroxide (15% of the dry weight of the water-leached residue) for acid leaching. After reacting for 7 hours, filter the slurry and collect the iron salt solution and the tantalum-niobium enriched residue. The tantalum-niobium enriched residue is washed, dried, and weighed 9.88 g. After drying, the tantalum-niobium enriched residue contains 5.02% iron (Fe), only 5.26% tantalum (Ta), and only 4.45% niobium (Nb).

[0055] The difference between this embodiment and Embodiment 1 is that the roasting temperature of the acid-leached tungsten-rich slag + sodium-containing molten salt agent is increased from 700℃ to 800℃, while the other conditions remain the same as in Embodiment 1. Increasing the roasting temperature of the molten salt slightly improves the leaching rate of WO3 compared to Embodiment 1, while the iron content of the tailings remains essentially unchanged. At 700℃, the tungsten in the slag is already sufficiently decomposed, and the iron is completely oxidized. Further increasing the roasting temperature to 800℃ does not significantly improve the decomposition effect; instead, it increases energy consumption.

[0056] Example 6 A method for the comprehensive recovery of valuable metals from difficult-to-treat tungsten-containing waste slag includes the following steps: S1. The tungsten-containing waste residue is crushed and screened to obtain recycled residue. 100g of the screened recycled residue is weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (theoretically 1.5 times the amount needed to dissolve cobalt) is added and stirred evenly. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled residue) is slowly added for acid leaching. After reacting for 5 hours, the slurry is filtered, and the filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich residue) are collected separately. The filter residue is washed, dried, and weighed to 80.0g.

[0057] S2. The acid-leached tungsten-rich slag obtained in step S1 is mixed evenly with 32g of sodium nitrate (40% of the dry weight of the acid-leached tungsten-rich slag) and 40mL of water. The mixture is then placed in a muffle furnace for oxidation roasting at a temperature of 700℃ and a holding time of 120min. After natural cooling, roasted slag is obtained.

[0058] S3. Take out the calcined residue obtained in step S2, add 150 mL of pure water for water leaching at 70℃ for 60 min. After water leaching, filter the slurry to obtain tungsten-containing water leaching solution and water leaching residue. The water leaching residue is washed, dried, and weighed 31.1 g. The WO3 content in the water leaching residue is tested to be 1.04%, and the WO3 leaching rate in this step is calculated to be 98.99%.

[0059] S4. Mix the water-leached residue obtained in step S3 with 30 mL of tap water to form a slurry. Add 114 mL of industrial hydrochloric acid (1.5 times the theoretical amount for iron dissolution) and stir until homogeneous. Then slowly add 4.67 g of hydrogen peroxide (15% of the dry weight of the water-leached residue) for acid leaching. After reacting for 7 hours, filter the slurry and collect the iron salt solution and the tantalum-niobium enriched residue. The tantalum-niobium enriched residue is washed, dried, and weighed 10.94 g. After drying, the tantalum-niobium enriched residue contains 5.58% iron (Fe), only 4.75% Ta, and only 4.02% Nb.

[0060] The difference between this embodiment and Embodiment 1 is that the roasting and holding time for the acid-leached tungsten-rich slag + sodium-containing molten salt agent is extended from 60 min to 120 min, while the other conditions remain the same as in Embodiment 1. Extending the roasting and holding time with molten salt slightly increases the leaching rate of WO3 compared to Embodiment 1, while the iron content in the tailings remains essentially unchanged. The tungsten in the slag is already fully decomposed and the iron is completely oxidized within 60 min of holding time. Further extending the roasting and holding time to 120 min does not significantly improve the decomposition effect; instead, it increases energy consumption.

[0061] Comparative Example 1 The difference between this comparative example and Example 1 is that the amount of hydrochloric acid used for acid leaching of cobalt in S1 is reduced from the theoretical 1.5 times to 1.4 times, and only the first acid leaching of cobalt step is performed, while the remaining conditions are the same as in Example 1.

[0062] Specifically, in Comparative Example S1, tungsten-containing waste slag was crushed and screened to obtain recycled slag. 100g of the screened recycled slag was weighed and mixed with 100mL of tap water to form a slurry. 80mL of industrial hydrochloric acid (theoretically 1.4 times the amount needed to dissolve cobalt) was added and stirred until homogeneous. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled slag) was slowly added for acid leaching. After reacting for 5 hours, the slurry was filtered, and the filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich slag) were collected separately. The filter residue, after washing and drying, weighed 82.5g. Testing revealed that the residue contained 2.12% cobalt, with a cobalt leaching rate of 90.63%.

[0063] After reducing the amount of hydrochloric acid used for cobalt leaching to 1.4 times the theoretical amount, the cobalt leaching rate was only 90.63%, indicating insufficient hydrochloric acid and unsatisfactory cobalt leaching effect.

[0064] Comparative Example 2 The difference between this comparative example and Example 1 is that the amount of hydrogen peroxide used in S1 for acid leaching of cobalt is reduced from 15% to 10% of the dry weight, and only the first acid leaching step is performed; the remaining conditions are the same as in Example 1.

[0065] Specifically, in Comparative Example S1, tungsten-containing waste slag was crushed and screened to obtain recycled slag. 100g of the screened recycled slag was weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (theoretically 1.5 times the amount needed for cobalt dissolution) was added and stirred until homogeneous. Then, 10g of hydrogen peroxide (10% of the dry weight of the recycled slag) was slowly added for acid leaching. After reacting for 5 hours, the slurry was filtered, and the filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich slag) were collected separately. The filter residue, after washing and drying, weighed 81.7g. Testing revealed that the residue contained 3.24% cobalt, with a cobalt leaching rate of 85.81%.

[0066] In this comparative study, after reducing the amount of hydrogen peroxide used for acid leaching of cobalt to 10%, the cobalt leaching rate was only 85.81%. The insufficient amount of hydrogen peroxide failed to fully enhance the cobalt leaching effect, resulting in an unsatisfactory cobalt leaching effect.

[0067] Comparative Example 3 The difference between this comparative example and Example 1 is that the reaction time for acid leaching of cobalt in S1 is shortened from 5 hours to 4 hours, and only the first acid leaching step is performed; the remaining conditions are the same as in Example 1.

[0068] Specifically, in Comparative Example S1, tungsten-containing waste slag was crushed and screened to obtain recycled slag. 100g of the screened recycled slag was weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (theoretically 1.5 times the amount needed for cobalt dissolution) was added and stirred until homogeneous. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled slag) was slowly added for acid leaching. After reacting for 4 hours, the slurry was filtered, and the filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich slag) were collected separately. The filter residue, after washing and drying, weighed 80.4g. Testing revealed that the residue contained 1.97% cobalt, with a cobalt leaching rate of 91.51%.

[0069] In this comparative example, after shortening the acid leaching time for cobalt to 4 hours, the cobalt leaching rate was only 91.51%. The reaction time was insufficient, and the cobalt in the residue was not fully dissolved, resulting in an unsatisfactory cobalt leaching effect.

[0070] Comparative Example 4 The difference between this comparative example and Example 1 is that the batching step of acid leaching tungsten-rich slag and sodium-containing molten salt agent roasting in S2 does not adopt a solid-liquid mixing method (no water is added), but only simple mixing is carried out, and only up to the S3 water leaching step, the remaining conditions are the same as in Example 1.

[0071] Specifically, S1-S3 of this comparative example are: S1. The tungsten-containing waste residue is crushed and screened to obtain recycled residue. 100g of the screened recycled residue is weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (theoretically 1.5 times the amount needed to dissolve cobalt) is added and stirred evenly. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled residue) is slowly added for acid leaching. After reacting for 5 hours, the slurry is filtered, and the filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich residue) are collected separately. The filter residue is washed, dried, and weighed to 78.6g.

[0072] S2. The acid-leached tungsten-rich slag obtained in step S1 is mixed evenly with 32g of sodium nitrate (40% of the dry weight of the acid-leached tungsten-rich slag) and then placed in a muffle furnace for oxidation roasting at a temperature of 700℃ for 60min. After natural cooling, roasted slag is obtained.

[0073] S3. Take out the calcined residue obtained in step S2, add 150 mL of pure water for water leaching at 70℃ for 60 min. After water leaching, filter the slurry to obtain tungsten-containing water leaching solution and water leaching residue. The water leaching residue is washed, dried, and weighed (33.7 g). The WO3 content in the water leaching residue is tested to be 6.23%, and the WO3 leaching rate in this step is calculated to be 93.46%.

[0074] The difference between this comparative example and Example 1 is that the mixing method before molten salt roasting was changed to simple mixing, and the leaching rate of WO3 was reduced to 93.46%. The uniformity of the mixture in simple mixing is not as good as that in solid-liquid mixing, and the tungsten leaching effect is not ideal. A larger amount of reagent is required to ensure that the molten salt fully wets the recycled residue during roasting.

[0075] Comparative Example 5 The difference between this comparative example and Example 1 is that the amount of hydrochloric acid used in the acid leaching of iron from the S4 water leaching residue was reduced from 1.5 times to 1.0 times, while the other conditions were the same as in Example 1.

[0076] The specific steps for this comparative example are as follows: S1. The tungsten-containing waste residue is crushed and screened to obtain recycled residue. 100g of the screened recycled residue is weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (theoretically 1.5 times the amount needed to dissolve cobalt) is added and stirred evenly. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled residue) is slowly added for acid leaching. After reacting for 5 hours, the slurry is filtered, and the filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich residue) are collected separately. The filter residue is washed, dried, and weighed to 77.5g.

[0077] S2. The acid-leached tungsten-rich slag obtained in step S1 is mixed evenly with 32g of sodium nitrate (40% of the dry weight of the acid-leached tungsten-rich slag) and 40mL of water. The mixture is then placed in a muffle furnace for oxidation roasting at a temperature of 700℃ and a holding time of 60min. After natural cooling, roasted slag is obtained.

[0078] S3. Take out the calcined residue obtained in step S2, add 150 mL of pure water for water leaching at a leaching temperature of 70℃ for 60 min. After water leaching, filter the slurry to obtain tungsten-containing water leaching solution and water leaching residue. The water leaching residue is washed, dried, and weighed to 31.7 g.

[0079] S4. Mix the water-leached residue obtained in step S3 with 31 mL of tap water to form a slurry. Add 76 mL of industrial hydrochloric acid (1.0 times the theoretical amount for iron dissolution) and stir until homogeneous. Then slowly add 4.76 g of hydrogen peroxide (15% of the dry weight of the water-leached residue) for acid leaching. After reacting for 7 hours, filter the slurry and collect the iron salt solution and the tantalum-niobium enriched residue. The tantalum-niobium enriched residue is washed, dried, and weighed 24.4 g. After drying, the tantalum-niobium enriched residue contains 30.77% iron (Fe), the Ta percentage only increases to 2.13%, and the Nb percentage only increases to 1.80%.

[0080] This comparative example reduces the amount of hydrochloric acid used during iron leaching to 1.0 times. Insufficient hydrochloric acid leads to incomplete iron dissolution, increased dry weight, and poor enrichment of tantalum and niobium.

[0081] Comparative Example 6 The difference between this comparative example and Example 1 is that the amount of hydrogen peroxide used in the acid leaching of iron from the S4 water leaching residue was reduced from 15% to 10% of the dry weight, while the other conditions were the same as in Example 1.

[0082] The specific steps for this comparative example are as follows: S1. The tungsten-containing waste residue is crushed and screened to obtain recycled residue. 100g of the screened recycled residue is weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (theoretically 1.5 times the amount needed to dissolve cobalt) is added and stirred evenly. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled residue) is slowly added for acid leaching. After reacting for 5 hours, the slurry is filtered, and the filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich residue) are collected separately. The filter residue is washed, dried, and weighed to 78.1g.

[0083] S2. The acid-leached tungsten-rich slag obtained in step S1 is mixed evenly with 32g of sodium nitrate (40% of the dry weight of the acid-leached tungsten-rich slag) and 40mL of water. The mixture is then placed in a muffle furnace for oxidation roasting at a temperature of 700℃ and a holding time of 60min. After natural cooling, roasted slag is obtained.

[0084] S3. Take out the calcined residue obtained in step S2, add 150mL of pure water for water leaching at a leaching temperature of 70℃ for 60min. After water leaching, filter the slurry to obtain tungsten-containing water leaching solution and water leaching residue. The water leaching residue is washed, dried, and weighed to 29.8g.

[0085] S4. Mix the water-leached residue obtained in step S3 with 30 mL of tap water to form a slurry. Add 114 mL of industrial hydrochloric acid (1.5 times the theoretical amount for iron dissolution) and stir until homogeneous. Then slowly add 2.98 g of hydrogen peroxide (10% of the dry weight of the water-leached residue) for acid leaching. After reacting for 7 hours, filter the slurry. Collect the iron salt solution and the tantalum-niobium enriched residue. The tantalum-niobium enriched residue is washed, dried, and weighed 21.6 g. The residue contains 27.35% iron, with the Ta percentage increasing to only 2.41% and the Nb percentage increasing to only 2.04%.

[0086] In this comparative study, reducing the amount of hydrogen peroxide used in acid leaching to 10% failed to adequately enhance the iron-dissolving effect, resulting in incomplete iron dissolution, increased dry weight, and poor enrichment of tantalum and niobium.

[0087] Comparative Example 7 The difference between this comparative example and Example 1 is that the reaction time for acid leaching of iron from the S4 water leaching residue was shortened from 7 hours to 4 hours, while the other conditions were the same as in Example 1.

[0088] The specific steps for this comparative example are as follows: S1. The tungsten-containing waste residue is crushed and screened to obtain recycled residue. 100g of the screened recycled residue is weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (theoretically 1.5 times the amount needed to dissolve cobalt) is added and stirred evenly. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled residue) is slowly added for acid leaching. After reacting for 5 hours, the slurry is filtered. The filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich residue) are collected separately. The filter residue is washed, dried, and weighed to 79.7g.

[0089] S2. The acid-leached tungsten-rich slag obtained in step S1 is mixed evenly with 32g of sodium nitrate (40% of the dry weight of the acid-leached tungsten-rich slag) and 40mL of water. The mixture is then placed in a muffle furnace for oxidation roasting at a temperature of 700℃ and a holding time of 60min. After natural cooling, roasted slag is obtained.

[0090] S3. Take out the calcined residue obtained in step S2, add 150 mL of pure water for water leaching at a temperature of 70℃ for 60 min. After water leaching, filter the slurry to obtain tungsten-containing water leaching solution and water leaching residue. The water leaching residue is washed, dried, and weighed to 32.0 g.

[0091] S4. Mix the water-leached residue obtained in step S3 with 32 mL of tap water to form a slurry. Add 114 mL of industrial hydrochloric acid (1.5 times the theoretical amount for iron dissolution) and stir until homogeneous. Then slowly add 4.8 g of hydrogen peroxide (15% of the dry weight of the water-leached residue) for acid leaching. After reacting for 4 hours, filter the slurry. Collect the iron salt solution and the tantalum-niobium enriched residue. The tantalum-niobium enriched residue is washed, dried, and weighed 22.3 g. The residue contains 29.16% iron, with the Ta percentage increasing to only 2.33% and the Nb percentage increasing to only 1.97%.

[0092] Shortening the acid leaching time to 4 hours resulted in incomplete iron dissolution, high iron content in the tailings (29.16%), increased dry weight, and poor enrichment of tantalum and niobium.

[0093] Comparative Example 8 The difference between this comparative example and Example 1 is that the roasting temperature of the S2 acid-leached tungsten-rich slag + sodium-containing molten salt agent is reduced from 700℃ to 600℃, while the other conditions are the same as in Example 1.

[0094] The specific steps for this comparative example are as follows: S1. The tungsten-containing waste residue is crushed and screened to obtain recycled residue. 100g of the screened recycled residue is weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (theoretically 1.5 times the amount needed to dissolve cobalt) is added and stirred evenly. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled residue) is slowly added for acid leaching. After reacting for 5 hours, the slurry is filtered. The filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich residue) are collected separately. The filter residue is washed, dried, and weighed to 81.3g.

[0095] S2. The acid-leached tungsten-rich slag obtained in step S1 is mixed evenly with 32g of sodium nitrate (40% of the dry weight of the acid-leached tungsten-rich slag) and 40mL of water. The mixture is then placed in a muffle furnace for oxidation roasting at a temperature of 600℃ and a holding time of 60min. After natural cooling, roasted slag is obtained.

[0096] S3. Take out the calcined residue obtained in step S2, add 150 mL of pure water for water leaching at 70℃ for 60 min. After water leaching, filter the slurry to obtain tungsten-containing water leaching solution and water leaching residue. The water leaching residue was washed, dried, and weighed 36.7 g. The residue was tested and found to contain 10.08% WO3, with a WO3 leaching rate of 88.47%.

[0097] S4. Mix the water-leached residue obtained in step S3 with 35 mL of tap water to form a slurry. Add 114 mL of industrial hydrochloric acid (1.5 times the theoretical amount for iron dissolution) and stir until homogeneous. Then slowly add 5.5 g of hydrogen peroxide (15% of the dry weight of the water-leached residue) for acid leaching. After reacting for 7 hours, filter the slurry. Collect the iron salt solution and the tantalum-niobium enriched residue. The tantalum-niobium enriched residue is washed, dried, and weighed 29.5 g. The residue contains 50.85% iron, with the Ta percentage increasing to only 1.76% and the Nb percentage increasing to only 1.49%.

[0098] After lowering the molten salt roasting temperature, the leaching rate of WO3 decreased to 88.47%, and the iron content of the tailings reached 50.85%. The excessively low roasting temperature resulted in incomplete decomposition of tungsten compounds, and the tungsten leaching rate decreased after water leaching. The iron in the roasted material was not completely oxidized, and acid leaching could not fully dissolve it, leading to an increase in dry weight and poor enrichment of tantalum and niobium.

[0099] Comparative Example 9 The difference between this comparative example and Example 1 is that the roasting and holding time of the S2 acid-leached tungsten-rich slag + sodium-containing molten salt agent is shortened from 60 min to 40 min, while the other conditions are the same as in Example 1.

[0100] The specific steps for this comparative example are as follows: S1. The tungsten-containing waste residue is crushed and screened to obtain recycled residue. 100g of the screened recycled residue is weighed and mixed with 100mL of tap water to form a slurry. 96mL of industrial hydrochloric acid (theoretically 1.5 times the amount needed to dissolve cobalt) is added and stirred evenly. Then, 15g of hydrogen peroxide (15% of the dry weight of the recycled residue) is slowly added for acid leaching. After reacting for 5 hours, the slurry is filtered. The filtrate (cobalt salt solution) and filter residue (acid-leached tungsten-rich residue) are collected separately. The filter residue is washed, dried, and weighed to 80.8g.

[0101] S2. The acid-leached tungsten-rich slag obtained in step S1 is mixed evenly with 32g of sodium nitrate (40% of the dry weight of the acid-leached tungsten-rich slag) and 40mL of water. The mixture is then placed in a muffle furnace for oxidation roasting at a temperature of 700℃ and a holding time of 40min. After natural cooling, roasted slag is obtained.

[0102] S3. Take out the calcined residue obtained in step S2, add 150 mL of pure water for water leaching at 70℃ for 60 min. After water leaching, filter the slurry to obtain tungsten-containing water leaching solution and water leaching residue. The water leaching residue was washed, dried, and weighed 34.6 g. The residue was tested and found to contain 7.51% WO3, with a WO3 leaching rate of 91.89%.

[0103] S4. Mix the water-leached residue obtained in step S3 with 35 mL of tap water to form a slurry. Add 114 mL of industrial hydrochloric acid (1.5 times the theoretical amount for iron dissolution) and stir until homogeneous. Then slowly add 5.19 g of hydrogen peroxide (15% of the dry weight of the water-leached residue) for acid leaching. After reacting for 7 hours, filter the slurry. Collect the iron salt solution and the tantalum-niobium enriched residue. The tantalum-niobium enriched residue is washed, dried, and weighed 26.9 g. The residue contains 37.17% iron, with the Ta percentage increasing to only 1.93% and the Nb percentage increasing to only 1.64%.

[0104] Shortening the holding time of molten salt roasting reduced the leaching rate of WO3 to 91.89%, and the iron content of the tailings reached 37.17%. Insufficient roasting reaction time led to incomplete decomposition of tungsten compounds, and the tungsten leaching rate decreased after water leaching. Iron in the roasted material could not be fully oxidized, and acid leaching could not fully dissolve it, resulting in an increase in dry weight and poor enrichment effect of tantalum and niobium.

[0105] The reaction conditions and effects of each embodiment and comparative example are compared in Table 2.

[0106] Table 2

[0107] The embodiments described above are merely illustrative of several implementations of the present invention, and while the descriptions are specific and detailed, they should not be construed as limiting the scope of the present invention. It should be noted that those skilled in the art can make various modifications and improvements without departing from the concept of the present invention, and these modifications and improvements all fall within the scope of protection of the present invention. Therefore, the scope of protection of this patent should be determined by the appended claims.

Claims

1. A method for comprehensively recovering valuable metals from difficult-to-process tungsten-containing waste slag, characterized in that, The method includes the following steps: S1. The tungsten-containing waste slurry obtained after crushing, screening and pulping pretreatment is subjected to cobalt leaching reaction with inorganic acid, and an oxidant is added to enhance the reaction to achieve selective leaching of cobalt. After solid-liquid separation, cobalt salt solution and acid-leached tungsten-rich slag are obtained. S2. The acid-leached tungsten-rich slag is mixed with sodium-containing molten salt agent and water solid-liquid mixture and then roasted to convert tungsten into soluble tungstate and iron into ferric oxide that is easily soluble in acid, thus obtaining roasted slag. S3. Mix the roasted residue with water and heat it for a water leaching reaction to allow tungsten to enter the solution. After solid-liquid separation, tungsten-containing water leaching solution and water leaching residue are obtained. S4. The water-leached residue is pulped and then subjected to an acid leaching reaction with an inorganic acid. An oxidant is added to enhance the reaction, thereby achieving iron leaching. After solid-liquid separation, an iron salt solution and tantalum-niobium enriched residue are obtained.

2. The method for comprehensively recovering valuable metals from difficult-to-treat tungsten-containing waste slag according to claim 1, characterized in that, The inorganic acid in S1 and S4 is hydrochloric acid, and the oxidizing agent is hydrogen peroxide.

3. A method for comprehensively recovering valuable metals from difficult-to-treat tungsten-containing waste slag according to claim 1 or 2, characterized in that, In S1, the amount of oxidant added is not less than 15 wt% of the dry weight of the tungsten-containing waste residue; the acid leaching reaction for dissolving cobalt continues for at least 5 hours after the addition of the oxidant.

4. A method for comprehensively recovering valuable metals from difficult-to-treat tungsten-containing waste slag according to claim 1 or 2, characterized in that, In S1, the amount of inorganic acid used is not less than 1.5 times the theoretical amount used to dissolve cobalt.

5. A method for comprehensively recovering valuable metals from difficult-to-treat tungsten-containing waste slag according to claim 1 or 2, characterized in that, In S4, the amount of oxidant added is not less than 15 wt% of the dry weight of the water-leached residue; the acid leaching iron dissolution reaction continues for not less than 7 hours after the addition of the oxidant.

6. A method for comprehensively recovering valuable metals from refractory tungsten-containing waste slag according to claim 1 or 2, characterized in that, In S4, the amount of inorganic acid used is not less than 1.5 times the theoretical amount of iron to be dissolved.

7. The method for comprehensively recovering valuable metals from difficult-to-treat tungsten-containing waste slag according to claim 1, characterized in that, In S2, the sodium-containing molten salt agent is sodium nitrate, or a combination of sodium nitrate and at least one of sodium carbonate and sodium hydroxide.

8. A method for comprehensively recovering valuable metals from difficult-to-treat tungsten-containing waste slag according to claim 1 or 7, characterized in that, In S2, the amount of sodium-containing molten salt agent is 40wt%~50wt% of the dry weight of the acid-leached tungsten-rich slag; the amount of water used during solid-liquid mixing is 40wt%~60wt% of the dry weight of the acid-leached tungsten-rich slag.

9. The method for comprehensively recovering valuable metals from difficult-to-treat tungsten-containing waste slag according to claim 1, characterized in that, In S2, the calcination temperature is 700℃~800℃, and the holding time is not less than 60 minutes.

10. The method for comprehensively recovering valuable metals from difficult-to-treat tungsten-containing waste slag according to claim 1, characterized in that, In S3, the temperature of the water immersion reaction is not lower than 70°C, and the water immersion time is not lower than 1 hour.

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