A method of extracting valuable components from a refining slag

Abstract

This invention discloses a method for extracting valuable components from refining slag, belonging to the field of solid waste resource utilization technology. The invention includes steps such as refining slag leaching, CO2 mineralization, silica product separation, and alumina product separation. By utilizing ammonium bisulfate as a recycling aid and employing indirect mineralization, the method achieves efficient leaching of valuable components from solid waste, ultimately converting the refining slag solid waste into products such as light carbonates, silica, and alumina, which can be used as building materials and raw materials for steel smelting. This invention solves the problems of low utilization rate of refining slag, low reaction efficiency during valuable component extraction, difficulty in aid recovery, and high energy consumption in existing technologies. This invention not only enables the resource utilization of valuable components such as calcium, magnesium, aluminum, and silicon in refining slag but also achieves CO2 emission reduction in the steel industry, resulting in significant economic and environmental benefits.

Description

Technical Field

[0001] This invention belongs to the field of solid waste resource utilization technology, specifically relating to a method for extracting valuable components from refining slag. Background Technology

[0002] Refining slag is a residual waste generated during the steel refining process. Its main components are unreacted calcium oxide, alumina, and silicon dioxide, along with potentially small amounts of other metals such as iron, manganese, and chromium, as well as their oxides, sulfides, and phosphides. As can be seen, refining slag has a complex composition and contains various harmful elements and impurities, making its treatment challenging. Currently, China discharges as much as 20 million tons of refining slag annually, of which only about 7 million tons are effectively utilized for the production of slag cement, slag powder, and other building materials; the remaining 13 million tons are stockpiled and have not been effectively disposed of or utilized. Currently, nearly 100 million tons of refining slag are piled up nationwide. If not properly treated, heavy metals in the refining slag may seep into the soil and water bodies, causing heavy metal pollution; tailings stockpiles may also occupy large amounts of land resources, impacting the ecological environment. Therefore, given the significant needs of the steel industry for low-carbon development and solid waste management, the rational treatment and utilization of refining slag is crucial for environmental protection and the sustainable use of resources.

[0003] Common methods for treating refining tailings include direct utilization, such as using them as paving materials, brick making, and foundation backfill, which has a significant environmental impact; stabilization treatment, such as high-pressure steam stabilization or direct carbon fixation, but even after stabilization, the refining tailings are still difficult to utilize; and separation and resource utilization, such as alkali separation and treatment to prepare 4A molecular sieves, but the market capacity for this product is small and its application is limited. Existing methods for treating and utilizing refining tailings have certain limitations; therefore, developing a more efficient and green resource utilization technology for refining tailings is an urgent problem to be solved. Summary of the Invention

[0004] In view of the above-mentioned prior art, the present invention provides a method for extracting valuable components from refining slag, which solves the problems of low utilization rate of refining slag, low reaction efficiency, difficulty in recovering additives, and high energy consumption in the extraction process of valuable components in the prior art.

[0005] To achieve the above objectives, the technical solution adopted by the present invention is: to provide a method for extracting valuable components from refining slag, comprising the following steps:

[0006] (1) Leaching of refining slag: Prepare a refining slag slurry by mixing the refining slag with bisulfate and water at a liquid-solid ratio of 0.70-1.60 mL:1 g, and then add the mineralized liquid phase product obtained in step (2). Stir at 400-600 rpm for 10-60 min at 40-60℃ to obtain a leaching slurry. Separate the leaching slurry into solid and liquid phases to obtain a leaching filtrate and a leaching filter cake. The refining slag is a refining slag containing calcium oxide, magnesium oxide, aluminum oxide and silicon oxide. The mass ratio of bisulfate to refining slag is 2.5-4.0:1.

[0007] (2) CO2 mineralization: Under the conditions of 20-30℃ and stirring speed of 400-600rpm, a CO2 gas with a volume concentration of 10%-100% is passed into the leaching filtrate for 0.5-2.0h to obtain a mineralized slurry; the mineralized slurry is subjected to solid-liquid separation to obtain a light carbonate product and a mineralized liquid phase product.

[0008] (3) Separation of silica products: The sulfuric acid solution and the leaching filter cake are prepared into a dissolving slurry with a liquid-solid ratio of 2-5 mL:1 g; the dissolving slurry is subjected to solid-liquid separation to obtain silica products and filtrate; the concentration of the sulfuric acid solution is 1.2-1.8 mol / L;

[0009] (4) Separation of alumina products: Under the conditions of 20-30℃ and stirring speed of 400-600rpm, add alkaline solvent to the filtrate until the pH of the solution is 7-9 to obtain alkaline precipitation slurry; perform solid-liquid separation on the alkaline precipitation slurry, collect the solid product and calcine to obtain alumina product; the concentration of alkaline solvent is 2.5-3.5mol / L, and the alkaline solvent is ammonia water, sodium bicarbonate solution or organic amine solution.

[0010] The beneficial effects of the above-mentioned technical solution of this invention are as follows: This invention utilizes ammonium bisulfate as a recycling aid and employs indirect mineralization to achieve efficient leaching of valuable components from solid waste. This includes steps such as leaching of valuable components from refining slag, mineralization separation of calcium and magnesium components, separation of silica, and separation of alumina. Ultimately, the refining slag solid waste is converted into products such as light carbonates, silica, and alumina, which can be used as building materials and raw materials for steel smelting. This method not only realizes the resource utilization of magnesium oxide, calcium oxide, alumina, and silica in waste slag but also achieves carbon emission reduction in the steel industry, featuring green and low-carbon characteristics, high economic benefits, and recycling of process by-products and raw materials. This method mainly involves two main reactions: mineralization reaction to prepare carbonates and reaction of ammonium aluminum sulfate with ammonia to prepare alumina.

[0011] (1) The principle of preparing carbonates by mineralization reaction includes the following three steps:

[0012] ①Mg in refining slag 2+ Ca 2+Dissolved in the solution,

[0013] CaO + 2H+ + →Ca 2+ +H2O (1)

[0014] MgO + 2H + →Mg 2+ +H2O (2)

[0015] ② CO2 in the gas phase dissolves into the aqueous phase.

[0016] CO2 + H2O → H2CO3 (3)

[0017]

[0018]

[0019] ③ CO3 in the aqueous phase 2- and Mg 2+ Ca 2+ Magnesium carbonate and calcium carbonate are produced.

[0020]

[0021]

[0022] (2) The principle of preparing alumina by reacting ammonium aluminum sulfate with ammonia is as follows:

[0023] NH4Al(SO4)2+3NH3·H2O→AlOOH+2(NH4)2 SO4+H2O (8)

[0024] AlOOH→Al2O3+H2O (9)

[0025] Based on the above technical solution, the present invention can be further improved as follows.

[0026] Furthermore, the total content of calcium oxide and magnesium oxide in the refining slag exceeds 45 wt%; its particle size does not exceed 200 mesh; the bisulfate is ammonium bisulfate, potassium bisulfate or sodium bisulfate, the liquid-solid ratio of the mixture of refining slag and bisulfate and water is 0.94-1.20 mL:1 g, and the mass ratio of bisulfate to refining slag is 3.2-4.0:1.

[0027] Furthermore, the bisulfate is ammonium bisulfate, and the liquid-solid ratio of the mixture of refining residue and ammonium bisulfate with water is 0.94 mL: 1 g.

[0028] Furthermore, CO2-containing gas is introduced into the leachate via bubbling at a flow rate of 0.1–0.2 m³ / s. 3 / h.

[0029] Furthermore, the calcination temperature is 650–750℃, and the time is 1.5–2.5 h; the organic amine solution is an ethanolamine solution or an N-methyldiethanolamine solution.

[0030] Furthermore, the roasting temperature was 700℃ and the time was 2 hours.

[0031] Furthermore, the solid-liquid separation methods in steps (1) to (4) are all natural sedimentation, filtration, centrifugation, or vacuum filtration.

[0032] Furthermore, the method for liquid-solid separation is vacuum filtration.

[0033] Further, after solid-liquid separation in steps (2) to (4), the solid product is washed and dried at a temperature of 60°C for 24 hours.

[0034] Further, in step (1), the stirring temperature is 50℃, the stirring speed is 500rpm, the time is 30min, and the mass ratio of hydrogen sulfate to refining slag is 3.2:1; in step (2), the stirring temperature is 30℃, the stirring speed is 500rpm; the volume concentration of CO2 in the gas introduced into the leaching filtrate is 12%, and the time is 0.5h; in step (3), the concentration of sulfuric acid solution is 1.5mol / L, and the liquid-solid ratio of sulfuric acid solution to leaching filter cake is 2mL:1g; in step (4), the stirring temperature is 30℃, the stirring speed is 500rpm; the alkaline solvent is ammonia water with a concentration of 3mol / L, and when the pH of the solution is 9, it is stirred for another 10min to obtain an alkaline precipitate slurry.

[0035] The beneficial effects of the above-mentioned further technical solutions of this invention are as follows: This method efficiently extracts calcium, magnesium, silicon, and aluminum components from refining slag, while fully utilizing the high-calcium and magnesium components to absorb and mineralize CO2, thus simultaneously achieving resource utilization of solid waste and CO2 mineralization and emission reduction. Since the direct CO2 mineralization reaction rate of refining slag is slow, the addition of additives during the mineralization process is unavoidable. This method uses bisulfate as a recycling additive, employing indirect mineralization to carbonate the active calcium and magnesium components in the refining slag, achieving efficient CO2 sequestration while allowing for the recycling of ammonium bisulfate, reducing the loss of additives.

[0036] The beneficial effects of this invention are:

[0037] (1) This invention makes comprehensive resource utilization of refining slag solid waste generated by metallurgy. The light carbonate products prepared can be used as building coatings after modification treatment, which improves the economic benefits of this process technology. The silicon oxide products and aluminum oxide products prepared can be used as raw materials for iron and steel smelting, which reduces the external demand for raw materials and realizes a green, circular and low-carbon economy.

[0038] (2) The calcium and magnesium components in the refining slag are relatively high. This method uses CO2 in industrial flue gas to carry out mineralization reaction with calcium and magnesium components. It has high storage rate and low energy consumption, and realizes carbon emission reduction and solid waste treatment of refining slag. It has positive significance for energy conservation, carbon reduction and green production in the steel industry.

[0039] (3) In the leaching process of refining slag, hydrogen sulfate is used as a recycling aid and the carbonation reaction of refining slag is carried out by indirect mineralization. This achieves efficient CO2 sequestration while allowing ammonium bisulfate to be recycled, reducing the amount of aid lost.

[0040] (4) In the process of preparing alumina products, ammonium aluminum sulfate is used as the aluminum salt, and an alkaline solvent, such as ammonia, is used as the precipitant to generate boehmite (AlOOH) by precipitation. In this reaction, nitrogen remains in the solution in the form of ammonium bisulfate, which can avoid the large amount of harmful gases such as SO3 and NH3 released by the direct roasting of ammonium aluminum sulfate and the loss of nitrogen. Boehmite is roasted to generate Al2O3 and H2O. The roasting temperature is low, the energy consumption is small, and it is environmentally friendly. Attached Figure Description

[0041] Figure 1 This is a process flow diagram for the extraction of valuable components from refining slag, wherein: 1. Ball mill; 2. First constant-temperature heating magnetic stirrer; 3. First three-necked flask; 4. First vacuum pump; 5. First suction flask; 6. N2 gas cylinder; 7. CO2 gas cylinder; 8. First mass flow meter; 9. Second mass flow meter; 10. Second constant-temperature heating magnetic stirrer; 11. Second three-necked flask; 12. CO2 online infrared analyzer; 13. Second suction flask; 14. Second vacuum pump; 15. Beaker; 16. Third vacuum pump; 17. Third suction flask; 18. Third constant-temperature heating magnetic stirrer; 19. Third three-necked flask; 20. Fourth vacuum pump; 21. Fourth suction flask.

[0042] Figure 2 Thermogravimetric-mass spectrum of the solid products after mineralization of calcium and magnesium leaching solution in Example 2;

[0043] Figure 3 The diagrams show the specific surface area and pore size distribution of AlOOH at different pH values; where A is the specific surface area of ​​AlOOH at pH 9, B is the pore size distribution of AlOOH at pH 9, C is the specific surface area of ​​AlOOH at pH 10, and D is the pore size distribution of AlOOH at pH 10.

[0044] Figure 4 SEM images of AlOOH at different pH values; where A represents pH 9 and B represents pH 10.

[0045] Figure 5The diagram shows the specific surface area and pore size distribution of Al2O3 after calcination in Example 2, where A represents the specific surface area and B represents the pore size distribution.

[0046] Figure 6 This is an SEM image of Al2O3, the product after calcination in Example 2. Detailed Implementation

[0047] The process flow for extracting valuable components from refining slag is as follows: Figure 1 As shown, the process is as follows:

[0048] (1) Refining slag leaching process: The refining slag is crushed in ball mill 1 to obtain refining slag powder; the refining slag powder is mixed with ammonium bisulfate and water to prepare refining slag slurry; the refining slag slurry and the mineralized liquid phase product obtained in step (2) are added to the first three-necked flask 3, and the mixture is heated in a water bath and stirred by the first constant temperature heating magnetic stirrer 2 to obtain leaching slurry; the leaching slurry is filtered by the first vacuum pump 4 and the first suction filtration flask 5 to separate the solid and liquid, and obtain leaching filtrate rich in calcium and magnesium and leaching filter cake rich in silicon and aluminum.

[0049] During the first leaching process of refining slag, the mineralized liquid phase product obtained in step (2) was not obtained. Therefore, the refining slag slurry was directly added to the first three-necked flask 3 and subsequent operations were carried out. During the second and subsequent leaching processes of refining slag, the mineralized liquid phase product obtained in step (2) was obtained. Therefore, the refining slag slurry and the mineralized liquid phase product were added together to the first three-necked flask 3 and subsequent operations were carried out.

[0050] (2) CO2 mineralization process: The leaching filtrate is added to the second three-necked flask 11. N2 gas and CO2 gas are introduced into the leaching filtrate of the second three-necked flask 11 in the form of bubbling using a microporous aerator. The gas output of the N2 gas cylinder 6 and the CO2 gas cylinder 7 is controlled by the first mass flow meter 8 and the second mass flow meter 9. While introducing CO2-containing gas, the second constant temperature heating magnetic stirrer 10 is used for water bath heating and stirring to obtain a mineralized slurry. During the mineralization process, the temperature change of the solution is measured by a thermometer, and the concentration change of CO2 at the outlet of the second three-necked flask 11 is measured by a CO2 infrared online analyzer 12. The mineralized slurry is filtered by the second vacuum pump 14 and the second suction filtration flask 13 to separate the solid and liquid phases. The obtained solid phase product is dried to obtain a light carbonate product. The obtained mineralized liquid phase product is introduced into the first three-necked flask 3 for use in the refining slag leaching process.

[0051] (3) Separation process of silicon oxide products: Add sulfuric acid solution and leaching filter cake to beaker 15 to dissolve and obtain dissolved slurry; use third vacuum pump 16 and third suction filtration bottle 17 to filter the dissolved slurry to separate solid and liquid, and dry the obtained solid product to obtain silicon oxide product, and the obtained liquid product is filtrate, which is then processed in the next step.

[0052] (4) Separation process of alumina products: The filtrate is placed in the third three-necked flask 19, and ammonia water is added dropwise to the filtrate. At the same time, the solution is heated and stirred by the third constant temperature magnetic stirrer 18 to obtain an alkaline precipitation slurry. The pH of the solution is detected by a pH meter during the process. The alkaline precipitation slurry is filtered by the third vacuum pump 20 and the third suction filtration flask 21 to separate the solid and liquid. The obtained solid product is dried and then calcined to obtain the alumina product.

[0053] The specific embodiments of the present invention will be described in detail below with reference to examples.

[0054] Example 1

[0055] A method for extracting valuable components from refining slag includes the following steps:

[0056] (1) Refining slag leaching process: The refining slag is crushed in ball mill 1 to obtain refining slag powder with a particle size of less than 200 mesh; a mixture of refining slag powder and ammonium bisulfate solid and water are prepared into refining slag slurry with a liquid-solid ratio of 0.94 mL:1 g, wherein the refining slag powder is 100 g and the ammonium bisulfate is 320 g; the refining slag slurry and the mineralized liquid phase product obtained in step (2) are added together to the first three-necked flask 3 and heated in a 50°C water bath for 30 min, while the first constant temperature heating magnetic stirrer 2 is stirred at a speed of 500 rpm to mix and disperse the refining slag slurry to obtain leaching slurry; the leaching slurry is separated into solid and liquid phases by vacuum filtration, and the obtained liquid phase product is a leaching filtrate rich in calcium and magnesium. The obtained solid phase product is washed with water and dried at a drying temperature of 60°C for 24 h to obtain a leaching filter cake rich in silicon and aluminum.

[0057] During the first leaching process of refining slag, the mineralized liquid phase product obtained in step (2) was not obtained. Therefore, the refining slag slurry was directly added to the first three-necked flask 3 and subsequent operations were carried out. During the second and subsequent leaching processes of refining slag, the mineralized liquid phase product obtained in step (2) was obtained. Therefore, the refining slag slurry and the mineralized liquid phase product were added together to the first three-necked flask 3 and subsequent operations were carried out.

[0058] (2) CO2 mineralization process: The leaching filtrate is added to the second three-necked flask 11. The output of N2 cylinder 6 and CO2 cylinder 7 is controlled by the first mass flow meter 8 and the second mass flow meter 9 to obtain simulated flue gas with a CO2 volume concentration of 12%. The simulated flue gas is then bubbled into the leaching filtrate of the second three-necked flask 11 at a flow rate of 0.15 m³ / s. 3 / h, while purging simulated flue gas, a second constant-temperature heating magnetic stirrer is used to heat the solution in a 30℃ water bath and stir at 500rpm for 30 minutes to carry out a mineralization reaction, resulting in a mineralized slurry. During the mineralization process, a thermometer is used to measure the temperature change of the solution, and a CO2 infrared online analyzer 12 is used to measure the change of CO2 concentration at the outlet of the second three-necked flask 11 over time. The mineralized slurry is separated into solid and liquid phases by vacuum filtration. The solid phase product obtained is washed with water and dried to obtain a light carbonate product. The drying temperature is 60℃ and the drying time is 24h. The resulting liquid phase product is the mineralized liquid phase product.

[0059] (3) Separation process of silicon oxide products: Add sulfuric acid solution with a concentration of 1.5 mol / L and leaching filter cake to beaker 15 for dissolution. The liquid-solid ratio is 2 mL: 1 g to obtain a dissolved slurry. The dissolved slurry is separated into solid and liquid phases by vacuum filtration. The obtained solid phase product is washed with water and dried to obtain silicon oxide product. The drying temperature is 60℃ and the drying time is 24 h. The obtained liquid phase product is the filtrate, which is used for the next step of alumina product separation process.

[0060] (4) Separation process of alumina product: The filtrate was placed in the third three-necked flask 19. Under the condition of heating in a water bath at 30°C, an ammonia solution with a concentration of 3.0 mol / L was slowly added dropwise to the solution. At the same time, the third constant temperature heating magnetic stirrer 18 was stirred at a speed of 500 rpm to disperse and mix the solution. During the process, the pH of the reaction solution was detected by a pH meter. When the pH reached 9, the addition of ammonia solution was stopped. The mixture was stirred for another 10 min to obtain an alkaline precipitate slurry. The alkaline precipitate slurry was separated into solid and liquid phases by vacuum filtration. The obtained solid product was washed with water and anhydrous ethanol in sequence and dried at a drying temperature of 60°C for 24 h. Then, it was calcined at a calcination temperature of 700°C for 2 h to obtain the alumina product.

[0061] Example 2

[0062] A method for extracting valuable components from refining slag involves preparing a refining slag slurry by mixing a mixture of refining slag powder and ammonium bisulfate solid with water at a liquid-to-solid ratio of 0.71 mL:1 g, with the remaining steps being the same as in Example 1.

[0063] Example 3

[0064] A method for extracting valuable components from refining residue, wherein the refining residue is leached in a 40°C water bath for 30 minutes, and the remaining steps are the same as in Example 1.

[0065] Example 4

[0066] A method for extracting valuable components from refining residue, wherein the refining residue is leached in a 50°C water bath for 60 minutes, and the remaining steps are the same as in Example 1.

[0067] Example 5

[0068] A method for extracting valuable components from refining slag, wherein 100g of refining slag powder and 400g of ammonium bisulfate are used in the leaching process of refining slag, and the remaining implementation steps are the same as in Example 1.

[0069] Example 6

[0070] A method for extracting valuable components from refining slag, wherein the CO2 mineralization process uses simulated flue gas with a CO2 volume concentration of 90%, and the remaining implementation steps are the same as in Example 1.

[0071] Example 7

[0072] A method for extracting valuable components from refining slag, comprising a CO2 mineralization process, a mineralization reaction of 2 hours, and the remaining implementation steps being the same as in Example 1.

[0073] Example 8

[0074] A method for extracting valuable components from refining slag involves stopping the addition of ammonia solution when the pH reaches 10 during the alumina product separation process, followed by stirring for 10 minutes. The remaining steps are the same as in Example 1.

[0075] Comparative Example

[0076] A method for extracting valuable components from refining slag includes the following steps:

[0077] (1) Refining slag leaching process: The refining slag is crushed in ball mill 1 to obtain refining slag powder with a particle size of less than 200 mesh; 100g of refining slag powder and water are mixed with water at a liquid-solid ratio of 0.94mL:1g to prepare refining slag slurry. The refining slag slurry is added to the first three-necked flask 3 and heated in a 50℃ water bath for 30min. At the same time, the first constant temperature heating magnetic stirrer 2 is stirred at a speed of 500rpm to mix and disperse the refining slag slurry to obtain leaching slurry; the leaching slurry is separated into solid and liquid phases by vacuum filtration. The obtained liquid phase product is a leaching filtrate rich in calcium and magnesium. The obtained solid phase product is washed with water and dried at a drying temperature of 60℃ for 24h to obtain a leaching filter cake rich in silicon and aluminum.

[0078] (2) CO2 mineralization process: The leaching filtrate is added to the second three-necked flask 11. The output of N2 cylinder 6 and CO2 cylinder 7 is controlled by the first mass flow meter 8 and the second mass flow meter 9 to obtain simulated flue gas with a CO2 volume concentration of 12%. The simulated flue gas is then bubbled into the leaching filtrate of the second three-necked flask 11 at a flow rate of 0.15 m³ / s. 3 / h, while purging simulated flue gas, a second constant-temperature heating magnetic stirrer is heated in a 30℃ water bath and stirred at 500rpm for 30min to carry out a mineralization reaction, resulting in a mineralized slurry; during the mineralization process, a thermometer is used to measure the temperature change of the solution, and a CO2 infrared online analyzer 12 is used to measure the change of CO2 concentration at the outlet of the second three-necked flask 11 over time; the mineralized slurry is separated into solid and liquid phases by vacuum filtration, the obtained solid product is washed with water and dried to obtain a light carbonate product, the drying temperature is 60℃ and the drying time is 24h.

[0079] (3) Separation process of silica products: Add sulfuric acid solution with a concentration of 1.5 mol / L to beaker 15 to dissolve the leaching filter cake. The liquid-solid ratio is 2 mL: 1 g to obtain a dissolved slurry. The dissolved slurry is separated into solid and liquid phases by vacuum filtration. The obtained solid product is washed with water and dried to obtain silica products. The drying temperature is 60℃ and the drying time is 24 h. The obtained liquid product is the filtrate, which is used for the next step of alumina product separation process.

[0080] (4) Separation process of alumina products: The filtrate was placed in the third three-necked flask 19. Under the condition of heating in a water bath at 30°C, an ammonia solution with a concentration of 3.0 mol / L was slowly added dropwise to the solution. At the same time, the third constant temperature heating magnetic stirrer 18 was stirred at a speed of 500 rpm to disperse and mix the solution. During the process, the pH of the reaction solution was detected by a pH meter. When the pH reached 9, the addition of ammonia solution was stopped. The solution was stirred for another 10 min to obtain an alkaline precipitate slurry. The alkaline precipitate slurry was separated into solid and liquid phases by vacuum filtration. The obtained solid product was washed with water and anhydrous ethanol in sequence and dried at a drying temperature of 60°C for 24 h. Then, it was calcined at a calcination temperature of 700°C for 2 h to obtain the alumina product.

[0081] Experimental Example

[0082] 1. The phase composition of the refining slag was analyzed by X-ray fluorescence spectroscopy, and the results are shown in Table 1. It can be seen that the total mineralizable calcium and magnesium content of the refining slag reaches 45 wt%, the aluminum content reaches 16 wt%, and the silicon content reaches 33 wt%. Selective and efficient separation of these valuable components with high content can realize the resource utilization of refining slag solid waste.

[0083] Table 1 Main components of refining slag samples

[0084]

[0085] 2. The calcium and magnesium ion content in the leaching filtrate was measured by inductively coupled plasma atomic emission spectrometry. The mass change of the refining residue before and after leaching and the volume of the leaching filtrate were measured by weighing method. The conversion rates (R) of magnesium and calcium were calculated by equations (10) and (11), respectively.

[0086]

[0087]

[0088] Where m1 is the mass (g) of the refined residue taken in the leaching experiment, ω1 and ω2 are the mass percentages (wt%) of magnesium and calcium in the refined residue, respectively; V is the volume (L) of the leaching filtrate, c1 and c2 are the concentrations (g / L) of magnesium ions and calcium ions in the leaching filtrate, respectively; m2 is the mass (g) of the leaching filter cake; and ω3 is the mass fraction (wt%) of calcium in the form of calcium sulfate in the leaching residue.

[0089] The extraction rates of magnesium and calcium in Example 1 were calculated to be 98.2% and 100%, respectively, while those in the comparative example were 68.7% and 73.4%, respectively. The results indicate that while using water as the leaching solvent in the comparative example allows CO2 to directly react with the refining slag slurry for mineralization, thus avoiding the energy costs associated with introducing additives, the leaching efficiency is low. The ammonium bisulfate solution used in this invention can fully leach the calcium and magnesium components from the refining slag, achieving efficient and selective separation of calcium and magnesium components from silicon and aluminum components in the refining slag.

[0090] The extraction rates of magnesium and calcium in Examples 2-5 were calculated, and the effects of different liquid-solid ratios, reaction times, and reaction temperatures on the extraction rates of calcium and magnesium ions were analyzed.

[0091] In Example 2, the extraction rates of magnesium and calcium were 92.1% and 89.8%, respectively. Compared with Example 1, it can be seen that the higher the liquid-to-solid ratio in the leaching process, the lower the viscosity of the slurry, the higher the mass transfer rate, and the more complete the reaction.

[0092] In Example 3, the extraction rates of magnesium and calcium were 92.3% and 92.8%, respectively. Compared with Example 1, it can be seen that the reaction temperature has a certain influence on the extraction of calcium and magnesium ions. Within a certain range, the higher the temperature, the greater the extraction rate of calcium and magnesium ions in the same time period, and the faster the reaction rate. This is because the increase in temperature increases the number of activated molecules of ammonium bisulfate, making the dissociation reaction of ammonium bisulfate easier to proceed. When the temperature is increased to 50°C, the reaction is basically completed. Further increasing the temperature will lead to an increase in heat exchange energy consumption.

[0093] In Example 4, the extraction rates of magnesium and calcium were 98.4% and 100%, respectively. Compared with Example 1, it can be seen that there is no significant difference in the extraction rate of calcium and magnesium ions between a reaction time of 30 min and a reaction time of 60 min. This indicates that the reaction of ammonium bisulfate leaching the refining residue is very rapid. Within just a few minutes, a large amount of calcium and magnesium are converted into the corresponding ions. When the time reaches 30 min, the reaction has been basically completed.

[0094] In Example 5, the extraction rates of magnesium and calcium were 98.9% and 100%, respectively. Compared with Example 1, it can be seen that the higher the content of ammonium bisulfate during the leaching process, the more fully the calcium and magnesium are leached from the refining residue. However, when the ratio of refining residue to refining residue is increased from 3.2:1 to 4.0:1, the calcium and magnesium extraction rate in the refining residue does not increase significantly. Therefore, while ensuring the extraction rate, the amount of ammonium bisulfate can be reduced as much as possible.

[0095] 3. The calcium carbonate and magnesium carbonate contents in the light carbonate product obtained in Example 2 were analyzed using a thermogravimetric analyzer. The results are as follows: Figure 2 As shown in the figure, the thermogravimetric curve has two weight loss regions. The first weight loss region occurs below 100℃, which is attributed to the desorption of adsorbed water from the mineralization products. The second weight loss region occurs between 550 and 800℃, corresponding to the mass spectrometry curve, where a large amount of CO2 is released, which is caused by the decomposition of calcium carbonate at high temperatures. Calculations show that the amount of light carbonate product obtained per ton of refining slag is 795.32 kg, which can fix approximately 263.27 kg of CO2, indicating that this method gives the refining slag a significant CO2 fixation potential. Thermogravimetric analysis of the light carbonate products obtained in Examples 6 and 7 showed that the amount of light carbonate product prepared per ton of refining slag was 802.15 kg and 805.23 kg, respectively. Compared with Example 1, in the CO2 mineralization process, using a gas with a CO2 volume concentration of 12% for 30 minutes, more than 97% of calcium and magnesium ions could be reacted. Increasing the CO2 concentration and the mineralization reaction time can improve the mineralization utilization rate.

[0096] 4. The specific surface area and pore volume / pore size of the precipitates obtained after adding ammonia in Examples 1 and 8 were analyzed using BET analysis. The results are as follows: Figure 3 As shown in Table 2; the morphology of the precipitated products was observed using SEM, and the results are as follows. Figure 4 As shown, the precipitate product has a boehmite (AlOOH) structure, and the higher its crystallinity, the more complete the crystal form. When the pH reaches 9 ( Figure 3 A and 3B) particles are the smallest and most uniform, far from their isoelectric point, exhibiting some repulsion between particles, resulting in relatively light aggregation. They also have larger specific surface area, pore volume, and average pore size, making them suitable as raw materials for calcination preparation of alumina products. When the pH reaches 10 (… Figure 3Although C and 3D have the highest crystallinity, the resulting products have a dense, plate-like, blocky structure with indistinct grain edges and exhibit aggregation. This is because the increased electrolyte concentration and the very thin thickness of the opposing ion layer surrounding the aggregates allow for contact and growth, thus intensifying aggregation and reducing the surface area to 377.72 m². 2 The product has a low pore volume and is highly viscous at higher pH values, making filtration difficult and increasing filtration costs in industry.

[0097] Table 2. Specific surface area, average pore volume, and pore size of AlOOH products under different pH conditions.

[0098]

[0099] 5. The specific surface area and pore volume / pore size of the calcined solid product in Example 1 were analyzed using BET analysis, and the results are as follows: Figure 5 As shown in Table 3, Figure 5 The 700 indicates calcination at 700℃; the morphology of the solid product was observed using SEM, and the results are as follows. Figure 6 As shown, the alumina product has a type IV mesoporous structure, the hysteresis loop is type H3, and the specific surface area of ​​Al2O3 is 130.1 m². 2 ·g.

[0100] Table 3. Specific surface area, average pore volume, and pore size of Al2O3 products calcined at 700℃

[0101]

[0102] The composition of the solid product after calcination in Example 1 was analyzed by ICP-OES, as shown in Table 4. It can be seen that the alumina product has a high purity, reaching 99.96%, and the γ-Al₂O₃ has a large specific surface area. Therefore, boehmite, as a precursor, can produce high-purity γ-Al₂O₃ with high specific surface area, large pore size and volume, and a bimodal distribution. γ-Al₂O₃ can be recycled as a raw material for refining furnaces, achieving a green circular economy.

[0103] Table 4. Impurity content and purity of alumina products

[0104]

[0105]

[0106] Although specific embodiments of the present invention have been described in detail with reference to examples, they should not be construed as limiting the scope of protection of this patent. Various modifications and variations that can be made by those skilled in the art without inventive effort within the scope described in the claims are still within the scope of protection of this patent.

Claims

1. A method for the extraction of valuable components from a refining slag, characterized in that, Includes the following steps: (1) Leaching of refining slag: Prepare a refining slag slurry by mixing the refining slag with bisulfate and water at a liquid-solid ratio of 0.70~1.60 mL:1 g, and then add the mineralized liquid phase product obtained in step (2). Stir at 400~600 rpm for 10~60 min at 40~60 ℃ to obtain the leaching slurry. The leaching slurry is subjected to solid-liquid separation to obtain leaching filtrate and leaching filter cake; the refining slag is a refining slag containing calcium oxide, magnesium oxide, aluminum oxide and silicon oxide; the mass ratio of the hydrogen sulfate to the refining slag is 2.5~4.0:1; (2) CO2 mineralization: Under the conditions of temperature of 20~30 ℃ and stirring speed of 400~600 rpm, a CO2 gas with a volume concentration of 10%~100% is introduced into the leaching filtrate for 0.5~2.0 h to obtain a mineralized slurry; the mineralized slurry is subjected to solid-liquid separation to obtain a light carbonate product and a mineralized liquid phase product; (3) Separation of silica products: The sulfuric acid solution and the leaching filter cake are prepared into a dissolving slurry with a liquid-solid ratio of 2~5 mL:1 g; the dissolving slurry is subjected to solid-liquid separation to obtain silica products and filtrate; the concentration of the sulfuric acid solution is 1.2~1.8 mol / L; (4) Separation of alumina products: Under the conditions of a temperature of 20~30 ℃ and a stirring speed of 400~600 rpm, an alkaline solvent is added to the filtrate until the solution pH is 7~9 to obtain an alkaline precipitate slurry; the alkaline precipitate slurry is subjected to solid-liquid separation, the solid phase is collected and calcined to obtain the alumina product; the concentration of the alkaline solvent is 2.5~3.5 mol / L, and the alkaline solvent is ammonia water, sodium bicarbonate solution or organic amine solution; The total content of calcium oxide and magnesium oxide in the refining slag exceeds 45 wt%, and its particle size does not exceed 200 mesh; the bisulfate is ammonium bisulfate, potassium bisulfate or sodium bisulfate, the liquid-solid ratio of the mixture of the refining slag and bisulfate and water is 0.94~1.20 mL:1 g, and the mass ratio of the bisulfate to the refining slag is 3.2~4.0:

1.

2. The method for extraction of valuable components from refining slag according to claim 1, characterized in that: The hydrogen sulfate is ammonium bisulfate, and the liquid-solid ratio of the mixture of refining slag, ammonium bisulfate, and water is 0.94 mL:1 g.

3. The method for extraction of valuable components from refining slag according to claim 1, characterized in that: The gas containing CO2 is bubbled into the leaching filtrate in a bubbling manner, and the gas flow rate is 0.1-0.2 m 3 / h.

4. The method for extraction of valuable components from refining slag according to claim 1, characterized in that: The calcination temperature is 650~750 ℃, and the time is 1.5~2.5 h; The organic amine solution is an ethanolamine solution or an N-methyldiethanolamine solution.

5. The method for extraction of valuable components from refining slag according to claim 4, characterized in that: The roasting temperature was 700 ℃ and the time was 2 h.

6. The method for extraction of valuable components from refining slag according to claim 1, characterized in that: The solid-liquid separation methods in steps (1) to (4) are all natural sedimentation, filtration, centrifugation or vacuum filtration.

7. The method for extraction of valuable components from refining slag according to claim 6, characterized in that: The solid-liquid separation method is vacuum filtration.

8. The method for extracting valuable components from refining slag according to claim 1, characterized in that: After solid-liquid separation in steps (1) to (4), the solid product is washed and dried at a temperature of 60 °C for 24 h.

9. The method for extracting valuable components from refining slag according to claim 1, characterized in that: In step (1), the stirring temperature is 50 ℃, the stirring speed is 500 rpm, and the time is 30 min. The mass ratio of the hydrogen sulfate to the refining slag is 3.2:

1. In step (2), the stirring temperature is 30 ℃ and the stirring speed is 500 rpm. The volume concentration of CO2 in the gas introduced into the leaching filtrate is 12%, and the time is 0.5 h. In step (3), the concentration of the sulfuric acid solution is 1.5 mol / L, and the liquid-solid ratio of the sulfuric acid solution to the leaching filter cake is 2 mL:1 g. In step (4), the stirring temperature is 30 ℃ and the stirring speed is 500 rpm. The alkaline solvent is ammonia water with a concentration of 3 mol / L. When the pH of the solution is 9, it is stirred for another 10 min to obtain an alkaline precipitate slurry.

Images