Method for regenerating silicon dioxide from ardealite by utilizing flotation and acid leaching coupled melting purification technology

Abstract

The invention provides a method for quality-divided regeneration of silicon dioxide from ardealite by using a flotation and acid leaching coupling melting purification technology, and relates to the technical field of regeneration of silicon dioxide from ardealite, the method comprises the following steps: S1, adding a nonionic collecting agent into ardealite ore pulp for reverse flotation to obtain rough machining ardealite; s2, carrying out direct flotation on the rough machining ardealite obtained in the step S1 to obtain silicon ardealite tailings; s3, drying and crushing the silicon phosphogypsum tailings, uniformly mixing the silicon phosphogypsum tailings with a washing agent, and acidizing to remove impurities, so as to obtain silicate; and S4, adding the silicate obtained in the step S3 into a dissolving agent, reacting to obtain silicon dioxide colloid, and calcining to obtain silicon dioxide with the purity as high as 99.9%.

Description

Technical Field

[0001] This invention relates to the field of phosphogypsum-derived silica technology, and more particularly to a method for fractional regeneration of silica from phosphogypsum using a coupled flotation and acid leaching melt purification technique. Background Technology

[0002] Phosphogypsum is an inevitable byproduct of wet-process phosphoric acid production, and its main component is calcium sulfate dihydrate (CaSO4·2H2O). Large-scale stockpiling of phosphogypsum not only occupies valuable land resources, but its soluble phosphorus, fluorine, heavy metals (such as arsenic, cadmium, and lead), and acidic substances are easily leached into the soil and groundwater by rainwater, causing serious environmental pollution and ecological risks. Therefore, the resource utilization and high-value utilization of phosphogypsum is a key issue in solving its environmental problems and achieving the green and sustainable development of the phosphorus chemical industry.

[0003] Besides its main component, calcium sulfate, phosphogypsum contains various impurities, including unreacted phosphate rock residues (such as phosphates and fluorides), acid-insoluble substances (mainly silicate minerals from phosphate rock), soluble salts (such as sulfates or phosphates of sodium, potassium, and magnesium), and trace amounts of heavy metals and radioactive elements. Notably, phosphogypsum contains a significant proportion of silicon (usually in the form of amorphous silica or silicates, reaching 15-25%). High-purity, high-value-added silica (SiO2) is an important industrial raw material, widely used in rubber, plastics, coatings, adhesives, ceramics, casting, electronic packaging materials, silicon compound synthesis (such as silica black and silica sol), and even the preparation of high-purity quartz sand. Recovering high-purity silica from phosphogypsum not only enables the recycling of silicon resources but also significantly enhances the comprehensive utilization value and economic viability of phosphogypsum.

[0004] Traditional chemical leaching methods (acid / alkali methods): Acid method (commonly hydrochloric acid or sulfuric acid): Acid is used to dissolve calcium sulfate and other soluble impurities (such as phosphates and some metal ions) in phosphogypsum, leaving acid-insoluble substances (mainly silicon dioxide) as residues. However, this method has significant drawbacks: (1) huge acid consumption: a large amount of acid (5-30 kg / t, i.e., the mass of acid added to each 1 t of material to be treated is 5-30 kg) is required to dissolve the main calcium sulfate, which is costly and generates a large amount of waste acid, making treatment difficult; (2) limited silicon purity: the residual silicon slag often encapsulates or adsorbs unwashed impurities (such as residual phosphorus, fluorine, iron, aluminum, etc.), making it difficult to obtain high-purity (e.g., >99%) silicon dioxide; (3) poor product performance: the obtained silicon dioxide is mostly amorphous, and the physical properties such as specific surface area and porosity often do not meet the requirements of high-end applications (such as white carbon black), and may contain a lot of impurities that affect its application performance; (4) high environmental pressure: a large amount of wastewater containing high concentrations of calcium salts (such as calcium chloride) and residual acid is generated, which is difficult to treat and easily causes secondary pollution. Alkaline methods (such as sodium hydroxide melting / leaching): While capable of dissolving silicon, these methods also suffer from high energy consumption, high alkali consumption, complex processes, and difficulties in subsequent separation (requiring precipitation of silicon dioxide from sodium silicate solution, and easily introducing sodium ion impurities). Furthermore, they are highly corrosive to equipment, resulting in poor economic and environmental performance. Physical separation methods (such as sieving and gravity separation): These methods struggle to effectively separate siliceous impurities tightly bound to calcium sulfate crystals or with similar particle sizes. Their pre-enrichment effect on silicon dioxide is limited, and the purity improvement is small. The application and limitations of flotation in phosphogypsum desilication: Flotation is widely used in mineral separation, utilizing differences in the physicochemical properties of mineral surfaces for separation. In phosphogypsum treatment, flotation is mainly used to remove siliceous gangue to improve gypsum purity (i.e., allowing silicon to enter the tailings), or to enrich other valuable elements (such as rare earth elements) in phosphogypsum.

[0005] However, there is limited research on flotation technology specifically for the efficient enrichment and purification of silica from phosphogypsum, and the results are not ideal. The main reasons are: (1) Silica in phosphogypsum often exists in the form of extremely fine particles, amorphous or tightly wrapped with gypsum and impurities, resulting in poor floatability; (2) It is difficult to find a flotation reagent system that can selectively collect siliceous minerals while effectively suppressing calcium sulfate and other impurities; (3) The silica-containing concentrate (tailings) produced during the flotation process is usually of low purity and still contains a large amount of impurities such as gypsum, phosphorus, and fluorine, making it difficult to directly obtain concentrates that can be used as high-value silica products; (4) A single flotation process is difficult to meet the high purity requirements of high-end applications.

[0006] In view of this, it is necessary to design an improved method for fractional regeneration of silica from phosphogypsum using a coupled flotation and acid leaching melt purification technique to solve the above problems. Summary of the Invention

[0007] The purpose of this invention is to provide a method for fractional regeneration of silica from phosphogypsum using a coupled flotation and acid leaching melt purification technique.

[0008] To achieve the above-mentioned objectives, in a first aspect, the present invention provides a method for fractional regeneration of silica from phosphogypsum using a coupled flotation and acid leaching melt purification technique, comprising the following steps:

[0009] S1. Add non-ionic collectors to the phosphogypsum slurry and perform reverse flotation to obtain coarse phosphogypsum.

[0010] S2. Perform positive flotation on the coarsely processed phosphogypsum obtained in step S1 to obtain silicophosphogypsum tailings.

[0011] S3. After drying and crushing the silicon phosphogypsum tailings, mix them evenly with washing agents and acidify them to remove impurities, thereby obtaining silicates.

[0012] S4. The silicate obtained in step S3 is added to a solvent to react and obtain silica gel. After calcination, silica is obtained. The solvent includes sodium pyrophosphate, sodium hydroxide, sodium carbonate, and potassium hydroxide.

[0013] Preferably, in step S4, the mass ratio of sodium pyrophosphate, sodium hydroxide, sodium carbonate, and potassium hydroxide in the solvent is 1:4:3:2, and the solid-liquid ratio of the silicate to the solvent is 1:3.

[0014] Preferably, in step S2, the positive flotation is carried out by the following method: adding a pH adjuster, a cationic collector, and a silicate inhibitor to the coarsely processed phosphogypsum, wherein the cationic collector is a mixture of dodecylamine and acetic acid, the pH adjuster is one or a mixture of sulfuric acid, hydrochloric acid, and nitric acid, and the silicate inhibitor is a mixed solution of sodium pyrophosphate, sodium hydroxide, sulfuric acid, and sodium silicate.

[0015] Preferably, the cationic collector contains 1% dodecylamine by mass and 99.8% acetic acid by mass. The addition rule for the cationic collector is: 120-650g of the cationic collector is added per 1t of the coarsely processed phosphogypsum.

[0016] Preferably, the silicate inhibitor comprises 10-60% sodium silicate, 10-60% sulfuric acid, 10-60% sodium pyrophosphate, and 10-60% sodium hydroxide, and the amount of silicate inhibitor added is 200g / t; the pH adjuster is one or a mixture of sulfuric acid, hydrochloric acid, and nitric acid, and the addition rule of the pH adjuster is: 0.1-2.5kg of the pH adjuster is added per 1t of the coarsely processed phosphogypsum.

[0017] Preferably, in step S3, the washing agent is hydrochloric acid with a mass fraction of 5-90% and an addition amount of 5-90g. The acidification and impurity removal temperature is 30-90℃ and the time is 0.25-1h.

[0018] Preferably, in step S4, the reaction process is carried out under stirring conditions, with a stirring speed of 100-200 r / min, a temperature of 25-90℃, and a reaction time of 1-3 h.

[0019] Preferably, in step S1, the nonionic collector is one of diesel oil, dodecyltrimethylammonium bromide, and acetic acid, and the addition rule of the nonionic collector is: the amount of nonionic collector added to each 1t of phosphogypsum slurry is 50-600g.

[0020] Preferably, in step S4, the calcination temperature is 600-1000℃ and the calcination time is 2h.

[0021] Secondly, the present invention provides a silicon dioxide prepared by the aforementioned preparation method, which has a purity greater than 98.5%.

[0022] The beneficial effects of this invention are:

[0023] 1. The present invention provides a method for fractional regeneration of silica from phosphogypsum using a coupled flotation and acid leaching melt purification technology. This method first separates silicon-containing impurities in phosphogypsum using flotation separation, and then uses a solvent to selectively extract silicon elements from the silicon-containing impurities into the liquid phase to achieve deep separation. The high-silicon tailings generated during the process can be recycled and reused to effectively dispose of the waste residue after impurity removal. The residual P, F and other impurity elements in the crystal lattice are removed by acid leaching washing treatment (i.e., acidification impurity removal), achieving deep removal of harmful elements and recycling of the phosphorus chemical system. Finally, silica with a purity of up to 99.90% is obtained through a high-temperature calcination process.

[0024] 2. The method provided by this invention aims to enrich high-silicon impurities through phosphogypsum flotation, enabling the resource reuse of the resulting high-silicon tailings. Residual insoluble lattice impurities are washed away, and silicon is then extracted and separated more completely using a dissolving agent. Finally, high-temperature calcination yields high-purity silicon dioxide. This method realizes the green application of phosphogypsum solid waste, develops a high-value utilization method for phosphogypsum, and provides new ideas and methods for the future high-end application of phosphogypsum resources.

[0025] 3. The method provided by this invention establishes a complete technology chain for phosphogypsum impurity removal, element recovery, and tailings resource utilization. It not only solves the environmental problems caused by phosphogypsum stockpiling, but also realizes the high-value utilization of silicon resources, providing a scientific basis and technical support for the comprehensive management of phosphogypsum.

[0026] 4. The method provided by this invention is simple to operate, consumes little energy, has a simple preparation process, and uses inexpensive reagents, thus reducing costs. The product obtained by this invention has a wide range of applications, including rubber, plastics, coatings, adhesives, ceramics, casting, electronic packaging materials, silicon compound synthesis (such as silica and silica sol), and even the preparation of high-purity quartz sand. Attached Figure Description

[0027] Figure 1 This is the process flow of the method proposed in this scheme for fractional regeneration of silica from phosphogypsum using a coupled flotation and acid leaching melt purification technique. Detailed Implementation

[0028] To make the objectives, technical solutions, and advantages of this invention clearer, the invention will be described in detail below with reference to the accompanying drawings and specific embodiments.

[0029] It should also be noted that, in order to avoid obscuring the present invention with unnecessary details, only the structures and / or processing steps closely related to the present invention are shown in the accompanying drawings, while other details that are not closely related to the present invention are omitted.

[0030] Additionally, it should be noted that the terms “comprising,” “including,” or any other variations thereof are intended to cover non-exclusive inclusion, such that a process, method, article, or apparatus that comprises a list of elements includes not only those elements but also other elements not expressly listed, or elements inherent to such process, method, article, or apparatus.

[0031] Please see Figure 1 As shown, the method for fractional regeneration of silica from phosphogypsum using a coupled flotation and acid leaching melt purification technique provided by the present invention includes the following steps:

[0032] S1. Add non-ionic collectors to the phosphogypsum slurry and perform reverse flotation to obtain coarse phosphogypsum.

[0033] S2. Perform positive flotation on the coarse phosphogypsum obtained in step S1 to obtain high-silica phosphogypsum tailings (silica content greater than 90%).

[0034] S3. After drying and crushing the above-mentioned high-silica phosphogypsum tailings, mix them evenly with washing agents and acidify them to remove impurities, thereby obtaining high-purity silicates.

[0035] S4. Add the high-purity silicate obtained in step S3 to the solvent and react under stirring conditions to obtain silica gel. After calcination, silica is obtained.

[0036] In the above technical solution, the mechanism of obtaining silicon dioxide from phosphogypsum by using flotation and acid leaching coupled melt purification technology is as follows: The method first uses forward and reverse flotation to remove most of the organic impurities and gypsum in phosphogypsum, thereby obtaining silicon slag with high silicon content. Then, high-purity silicate is obtained by washing with hydrochloric acid. The silicate is dissolved by a strong alkaline solvent. Sodium pyrophosphate is used to chelate metal cations to prevent precipitation. After filtration, Si gel is generated. Finally, silicon dioxide is obtained by high-temperature calcination in a muffle furnace.

[0037] In some embodiments, in step S1, the mass fraction of the phosphogypsum slurry is 20-40%. The nonionic collector is one of diesel oil, dodecyltrimethylammonium bromide, and acetic acid. When dodecyltrimethylammonium bromide is used as the nonionic collector, its mass fraction is 1%, and the addition amount is 50-600 g / t, that is, the amount of nonionic collector added to each 1 t of phosphogypsum slurry is 50-600 g.

[0038] In some embodiments, in step S2, the positive flotation of coarse phosphogypsum is achieved by first adding a pH adjuster to the coarse phosphogypsum, then adding a silicate inhibitor, and finally adding a cationic collector to separate silica slag from the coarse phosphogypsum. In this step, the addition of a pH adjuster allows the phosphogypsum and silicate minerals to dissociate first, and the addition of a silicate inhibitor suppresses the silicate slag to the bottom of the slurry, reducing the possibility of foam carrying silicate along with gypsum. Finally, a cationic collector is used to collect the gypsum in a directional manner. Specifically, the cationic collector is a mixture of dodecylamine and acetic acid, with an addition amount of 120-650 g / t, that is, the amount of cationic collector added per 1 t of crude phosphogypsum is 120-650 g, the mass fraction of dodecylamine in the cationic collector is 1%, and the concentration of acetic acid is 99.8%. In some embodiments, it is specifically obtained by mixing 100 mL of 1% dodecylamine with 0.5 mL of 99.8% acetic acid. The pH adjuster is a mixture of one or more of sulfuric acid, hydrochloric acid, and nitric acid, with a pH adjustment range of 1-4, and the amount of pH adjuster added per 1 t of crude phosphogypsum is 0.1-2.5 kg. The silicate inhibitor is a mixed solution obtained by mixing sodium pyrophosphate, sodium hydroxide, sulfuric acid and sodium silicate, with a solid-liquid ratio of (1:1)-(1:10). The mass fraction of sodium silicate is 10-60%, the mass fraction of sulfuric acid is 10-60%, the mass fraction of sodium pyrophosphate is 10-60%, the mass fraction of sodium hydroxide is 10-60%, and the amount of silicate inhibitor added is 200g / t.

[0039] In some embodiments, in step S3, the washing agent is hydrochloric acid, with a mass fraction of 5-90%, and an addition amount of 5-90g. The acidification and impurity removal temperature is 30-90℃, and the time is 0.25-1h. The main role of hydrochloric acid in this step is to wash away the small amount of gypsum and easily soluble substances in the product obtained after washing away most of the calcium sulfate and other easily soluble substances by forward and reverse flotation, so as to ensure that high-purity silica is subsequently obtained.

[0040] In some embodiments, in step S4, the solvent is obtained by mixing sodium pyrophosphate, sodium hydroxide, sodium carbonate, and potassium hydroxide in a mass ratio of 1:4:3:2, and the solid-liquid ratio of high-purity silicate (g) to solvent (g) is 1:3. It should be noted that the solid-liquid ratio described herein represents the mass ratio of the solid to the liquid substance.

[0041] In some embodiments, in step S4, the stirring speed is 100-200 r / min, and the mixture is vibrated or stirred at 25-90°C for 1-3 hours.

[0042] In some embodiments, in step S4, the calcination temperature is 600-1000℃ and the calcination time is 2h. Preferably, the calcination temperature is 800℃.

[0043] The method for fractional regeneration of silica from phosphogypsum using a coupled flotation and acid leaching melt purification technique, provided by the present invention, will be further illustrated below with specific embodiments:

[0044] Example 1

[0045] This embodiment provides a method for fractional regeneration of silica from phosphogypsum using a coupled flotation and acid leaching melt purification technique, comprising the following steps:

[0046] S1. Mix 166g of phosphogypsum (the main crystalline phase is calcium sulfate dihydrate) with water to prepare a slurry with a mass fraction of 33%.

[0047] Organic matter was separated by reverse flotation in a self-assembled variable frequency aerated single-cell flotation machine: 800 g / t of diesel oil was added to the slurry obtained in step S1, and after mixing for 3 minutes, 60 g / t of frother was added for flotation. The frothing product was organic tailings. Then, 210 g / t of dodecyltrimethylammonium bromide was added for reverse flotation to remove organic gypsum intercalation minerals, resulting in coarsely processed phosphogypsum. The frother was pine oil.

[0048] S2. Separation of silica slag by positive flotation in a self-assembled variable frequency aerated single-cell flotation machine: Add 10% sulfuric acid to the coarse phosphogypsum obtained in step S1 to ensure the initial pH of the solution is 2; then, add a silicate inhibitor at a rate of 200 g / t, mix for 3 min, and then add a cationic collector at a rate of 210 g / t. The froth product is concentrate gypsum, and the slurry product is silica slag. Take the tailings of high-silica phosphogypsum from the bottom tank (the silica content in the tailings is 90%). The silicate inhibitor is a mixture of 5 g sodium pyrophosphate, 10 g sodium hydroxide, 5 mL of 98% sulfuric acid, and 23.5 g sodium silicate. The cationic collector is a mixture of 1 g dodecylamine and 0.5 mL of 99.8% acetic acid.

[0049] S3. After drying the high-silica phosphogypsum tailings obtained in step S2 at 45℃ for 2 hours, crush and grind them to pass through a 200-mesh sieve to obtain 32.30g of dried high-silica phosphogypsum tailings (denoted as PG-Si material); mix the PG-Si material with hydrochloric acid solution at a solid-liquid ratio of 1:4, heat and stir at 90℃ for 0.3 hours at a stirring speed of 200 r / min, filter to obtain high-purity silicate (silicate content of 90%), and dry the product to obtain H-Si, with a mass of 28.25g; wherein, the mass fraction of hydrochloric acid solution is 25%;

[0050] S4. The H-Si obtained in step S3 is mixed with the solvent at a solid-liquid ratio of 1:3 and stirred at 90°C for 1 hour at a stirring speed of 200 r / min to obtain silica gel. Then, the silica gel is subjected to vacuum pressure to evaporate water, yielding Si gel. The Si gel is calcined at 800°C for 2 hours to obtain silica with a purity of 99.00%. The solvent has a mass fraction of 40% (the ratio of the sum of the masses of all solutes to the mass of the solvent), and is obtained by mixing sodium pyrophosphate, sodium hydroxide, sodium carbonate, and potassium hydroxide in a mass ratio of 1:4:3:2 (sodium pyrophosphate:sodium hydroxide:sodium carbonate:potassium hydroxide = 1:4:3:2), and then adding the corresponding amount of water according to the desired solvent concentration.

[0051] Example 2

[0052] The only difference between this embodiment and Example 1 is that the mass fraction of the solvent used in step S4 is different from that in Example 1. The specific preparation method is as follows:

[0053] S1. Mix 166g of phosphogypsum (the main crystalline phase is calcium sulfate dihydrate) with water to prepare a slurry with a mass fraction of 33%.

[0054] Organic matter was separated by reverse flotation in a self-assembled variable frequency aerated single-cell flotation machine: 800 g / t of diesel oil was added to the slurry obtained in step S1, and after mixing for 3 min, 60 g / t of frother was added for flotation. The froth product was organic tailings. Then, 210 g / t of dodecyltrimethylammonium bromide was added for reverse flotation to remove organic gypsum intercalation minerals and obtain coarsely processed phosphogypsum.

[0055] S2. Separation of silica slag by positive flotation in a self-assembled variable frequency aerated single-cell flotation machine: Add 10% sulfuric acid to the coarse phosphogypsum obtained in step S1 to ensure the initial pH of the solution is 2; then, add a silicate inhibitor at a rate of 200 g / t, mix for 3 min, and then add a self-proportioned cationic collector at a rate of 210 g / t. The froth product is concentrate gypsum, and the slurry product is silica slag. Take the tailings of high-silica phosphogypsum from the bottom tank (the silica content in the tailings reaches more than 90%).

[0056] S3. After drying the high-silica phosphogypsum tailings obtained in step S2 at 45℃ for 2 hours, crush and grind them to pass through a 200-mesh sieve to obtain 32.30g of dried high-silica phosphogypsum tailings (denoted as PG-Si material); mix the PG-Si material with hydrochloric acid solution at a solid-liquid ratio of 1:4, heat and stir at 90℃ for 0.3 hours at a stirring speed of 200 r / min, filter to obtain high-purity silicate (silicate content of 90%), and dry the product to obtain H-Si, with a mass of 28.25g; wherein, the mass fraction of hydrochloric acid solution is 25%;

[0057] S4. The H-Si obtained in step S3 is mixed with the solvent at a solid-liquid ratio of 1:3 and stirred at 90°C for 1 hour at a stirring speed of 200 r / min to obtain silica gel. Then, the silica gel is subjected to vacuum pressure in a vessel to evaporate water, obtaining Si gel. The Si gel is calcined at 800°C for 2 hours to obtain silica with a purity of 98.50%. The solvent has a mass fraction of 50% and is obtained by mixing sodium pyrophosphate, sodium hydroxide, sodium carbonate, and potassium hydroxide in a mass ratio of 1:4:3:2 (sodium pyrophosphate:sodium hydroxide:sodium carbonate:potassium hydroxide = 1:4:3:2), and then adding the corresponding amount of water according to the desired solvent concentration. The specific masses of the four substances are 5g, 20g, 15g, and 10g, respectively.

[0058] Example 3

[0059] The only difference between this embodiment and Example 1 is that the mass fraction of the solvent used in step S4 is different from that in Example 1. The specific preparation method is as follows:

[0060] S1. Mix 166g of phosphogypsum (the main crystalline phase is calcium sulfate dihydrate) with water to prepare a slurry with a mass fraction of 33%.

[0061] Organic matter was separated by reverse flotation in a self-assembled variable frequency aerated single-cell flotation machine: 800 g / t of diesel oil was added to the slurry obtained in step S1, and after mixing for 3 min, 60 g / t of frother was added for flotation. The froth product was organic tailings. Then, 210 g / t of dodecyltrimethylammonium bromide was added for reverse flotation to remove organic gypsum intercalation minerals and obtain coarsely processed phosphogypsum.

[0062] S2. Separation of silica slag by positive flotation in a self-assembled variable frequency aerated single-cell flotation machine: Add 10% sulfuric acid to the coarse phosphogypsum obtained in step S1 to ensure the initial pH of the solution is 2; then, add a silicate inhibitor at a rate of 200 g / t, mix for 3 min, and then add a self-proportioned cationic collector at a rate of 210 g / t. The froth product is concentrate gypsum, and the slurry product is silica slag. Take the tailings of high-silica phosphogypsum from the bottom tank (the silica content in the tailings is 90%).

[0063] S3. After drying the high-silica phosphogypsum tailings obtained in step S2 at 45℃ for 2 hours, crush and grind them to pass through a 200-mesh sieve to obtain 32.30g of dried high-silica phosphogypsum tailings (denoted as PG-Si material); mix the PG-Si material with hydrochloric acid solution at a solid-liquid ratio of 1:4, heat and stir at 90℃ for 0.3 hours at a stirring speed of 200 r / min, filter to obtain high-purity silicate (silicate content of 90%), and dry the product to obtain H-Si, with a mass of 28.25g; wherein, the mass fraction of hydrochloric acid solution is 25%;

[0064] S4. The H-Si obtained in step S3 is mixed with the solvent at a solid-liquid ratio of 1:3 and stirred at 90°C for 1 hour at a stirring speed of 200 r / min to obtain silica gel. Then, the silica gel is placed in a vacuum pressure vessel to evaporate the water, obtaining Si gel. The Si gel is calcined at 800°C for 2 hours to obtain silica with a purity of 99.90%. The solvent has a mass fraction of 60% and is obtained by mixing sodium pyrophosphate, sodium hydroxide, sodium carbonate, and potassium hydroxide in a mass ratio of 1:4:3:2 (sodium pyrophosphate:sodium hydroxide:sodium carbonate:potassium hydroxide = 1:4:3:2), and then adding the corresponding amount of water according to the desired solvent concentration.

[0065] Example 4

[0066] The only differences between this embodiment and Example 1 are: in step S3, the concentration of the hydrochloric acid solution is different from that in Example 1, and in step S4, the mass fraction of the solvent used is different from that in Example 1. The specific preparation method is as follows:

[0067] S1. Mix 166g of phosphogypsum (the main crystalline phase is calcium sulfate dihydrate) with water to prepare a slurry with a mass fraction of 33%.

[0068] Organic matter was separated by reverse flotation in a self-assembled variable frequency aerated single-cell flotation machine: 800 g / t of diesel oil was added to the slurry obtained in step S1, and after mixing for 3 min, 60 g / t of frother was added for flotation. The froth product was organic tailings. Then, 210 g / t of dodecyltrimethylammonium bromide was added for reverse flotation to remove organic gypsum intercalation minerals and obtain coarsely processed phosphogypsum.

[0069] S2. Separation of silica slag by positive flotation in a self-assembled variable frequency aerated single-cell flotation machine: Add 10% sulfuric acid to the coarse phosphogypsum obtained in step S1 to ensure the initial pH of the solution is 2; then, add a silicate inhibitor at a rate of 200 g / t, mix for 3 min, and then add a self-proportioned cationic collector at a rate of 210 g / t. The froth product is concentrate gypsum, and the slurry product is silica slag. Take the tailings of high-silica phosphogypsum from the bottom tank (the silica content in the tailings is 90%).

[0070] S3. After drying the high-silica phosphogypsum tailings obtained in step S2 at 45℃ for 2 hours, crush and grind them to pass through a 200-mesh sieve to obtain 32.30g of dried high-silica phosphogypsum tailings (denoted as PG-Si material); mix the PG-Si material with hydrochloric acid solution at a solid-liquid ratio of 1:4, heat and stir at 90℃ for 0.3 hours at a stirring speed of 200 r / min, filter to obtain high-purity silicate (silicate content of 90%), and dry the product to be denoted as H-Si, with a mass of 28.25g; wherein, the mass fraction of hydrochloric acid solution is 40%;

[0071] S4. The H-Si obtained in step S3 is mixed with the solvent at a solid-liquid ratio of 1:3 and stirred at 90°C for 1 hour at a stirring speed of 200 r / min to obtain silica gel. Then, the silica gel is subjected to vacuum pressure to evaporate water, yielding Si gel. The Si gel is calcined at 800°C for 2 hours to obtain silica with a purity of 99.95%. The solvent has a mass fraction of 70% and is obtained by mixing sodium pyrophosphate, sodium hydroxide, sodium carbonate, and potassium hydroxide in a mass ratio of 1:4:3:2 (sodium pyrophosphate:sodium hydroxide:sodium carbonate:potassium hydroxide = 1:4:3:2), and then adding the corresponding amount of water according to the desired solvent concentration.

[0072] Comparative Example 1

[0073] The only difference between this comparative example and Example 2 is that only flotation of phosphogypsum is performed, while acid leaching coupled with melt purification is omitted, i.e., steps S3 and S4 are omitted. The remaining experimental steps and parameters are the same as in Example 1, and will not be repeated here. The silica obtained in this comparative example has a purity of 91.20%.

[0074] Comparative Example 2

[0075] The only difference between this comparative example and Example 2 is that the acid leaching treatment is omitted, i.e., step S3 is omitted. The remaining experimental steps and parameters are the same as in Example 1 and will not be repeated here. The silica prepared in this comparative example has a purity of 95.50%.

[0076] Comparative Example 3

[0077] The only difference between this comparative example and Example 2 is that only phosphogypsum is subjected to flotation, omitting melt purification, i.e., step S4 is omitted. The remaining experimental steps and parameters are the same as in Example 1 and will not be repeated here. The silica obtained in this comparative example has a purity of 93.20%.

[0078] Comparative examples 4 to 7

[0079] The only difference between Comparative Examples 4 to 7 and Example 2 is that the composition of the solvent used in step S4 is different from that in Example 2. The remaining experimental steps and parameters are the same as those in Example 1, and will not be repeated here.

[0080] The composition of the solvent in Examples 2 and Comparative Examples 4 to 7 is shown in Table 1. The data in the table shows that the four components of the solvent have a synergistic effect. Sodium hydroxide and potassium hydroxide mainly provide strong alkalinity to dissolve silicates and destroy the silicon-oxygen framework. Sodium carbonate mainly acts as a buffer to adjust the pH. Sodium pyrophosphate mainly acts as a metal cation chelating agent to prevent the precipitation of unremoved metal cations. Under the synergistic effect among the four components, silicates can be dissolved to the maximum extent to obtain silica gel with a purity of 99.90%.

[0081] Table 1. Composition of the solvents in Example 2 and Comparative Examples 4 to 7

[0082]

[0083] The above embodiments are only used to illustrate the technical solutions of the present invention and are not intended to limit it. Although the present invention has been described in detail with reference to preferred embodiments, those skilled in the art should understand that modifications or equivalent substitutions can be made to the technical solutions of the present invention without departing from the spirit and scope of the technical solutions of the present invention.

Claims

1. A method for fractional regeneration of silica from phosphogypsum using a coupled flotation and acid leaching melt purification technique, characterized in that, Includes the following steps: S1. Add non-ionic collectors to the phosphogypsum slurry and perform reverse flotation to obtain coarse phosphogypsum. S2. Perform positive flotation on the coarsely processed phosphogypsum obtained in step S1 to obtain silicophosphogypsum tailings. S3. After drying and crushing the silicon phosphogypsum tailings, mix them evenly with washing agents and acidify them to remove impurities, thereby obtaining silicates. S4. The silicate obtained in step S3 is added to a solvent to react and obtain silica gel. After calcination, silica is obtained. The solvent includes sodium pyrophosphate, sodium hydroxide, sodium carbonate, and potassium hydroxide.

2. The method according to claim 1, characterized in that, In step S4, the mass ratio of sodium pyrophosphate, sodium hydroxide, sodium carbonate, and potassium hydroxide in the solvent is 1:4:3:2, and the solid-liquid ratio of the silicate to the solvent is 1:

3.

3. The method according to claim 1, characterized in that, In step S2, the positive flotation is carried out by the following method: a pH adjuster, a cationic collector, and a silicate inhibitor are added to the crude phosphogypsum. The cationic collector is a mixture of dodecylamine and acetic acid. The pH adjuster is one or a mixture of sulfuric acid, hydrochloric acid, and nitric acid. The silicate inhibitor is a mixed solution of sodium pyrophosphate, sodium hydroxide, sulfuric acid, and sodium silicate.

4. The method according to claim 3, characterized in that, The cationic collector contains 1% dodecylamine and 99.8% acetic acid. The addition rule for the cationic collector is that the amount of cationic collector added to each 1t of the crudely processed phosphogypsum is 120-650g.

5. The method according to claim 3, characterized in that, The silicate inhibitor comprises sodium silicate with a mass fraction of 10-60%, sulfuric acid with a mass fraction of 10-60%, sodium pyrophosphate with a mass fraction of 10-60%, and sodium hydroxide with a mass fraction of 10-60%. The amount of silicate inhibitor added is 200 g / t. The pH adjuster is one or a mixture of sulfuric acid, hydrochloric acid, and nitric acid. The addition rule for the pH adjuster is: 0.1-2.5 kg of the pH adjuster is added per 1 t of the coarsely processed phosphogypsum.

6. The method according to claim 1, characterized in that, In step S3, the washing agent is hydrochloric acid with a mass fraction of 5-90% and an addition amount of 5-90g. The acidification and impurity removal temperature is 30-90℃ and the time is 0.25-1h.

7. The method according to claim 1, characterized in that, In step S4, the reaction is carried out under stirring conditions, with a stirring speed of 100-200 r / min, a temperature of 25-90℃, and a reaction time of 1-3 h.

8. The method according to claim 1, characterized in that, In step S1, the nonionic collector is one of diesel oil, dodecyltrimethylammonium bromide, and acetic acid. The addition rule of the nonionic collector is: the amount of nonionic collector added to each 1t of phosphogypsum slurry is 50-600g.

9. The method according to claim 1, characterized in that, In step S4, the calcination temperature is 600-1000℃ and the calcination time is 2h.

10. A type of silicon dioxide, characterized in that, The sample is prepared by the method described in any one of claims 1-9 and has a purity greater than 98.5%.

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