Phosphorus tailing comprehensive utilization method

Abstract

The invention discloses a comprehensive utilization method of phosphate tailings, and relates to the field of solid waste resource utilization. Comprising the following steps: S1, roasting the phosphate tailings under a medium-temperature condition; s2, the roasted product obtained in the step S1 and an ammonium salt solution are subjected to size mixing according to a certain proportion, and full stirring is conducted to enable ammonium salt to leach and extract magnesium; s3, carrying out solid-liquid separation on the ammonium leaching product obtained in S2 to obtain a magnesium-containing solution and a filter cake; and S4, carrying out direct flotation or reverse flotation on a filter cake obtained by filtering in the step S3 to recover apatite. Through the combined process of roasting, ammonium leaching and flotation, three valuable elements of phosphorus, magnesium and calcium in the phosphate tailings can be economically and efficiently separated one by one and converted into valuable products, the recycling degree is high, and the comprehensive utilization value of the phosphate tailings can be improved; and the method is low in energy consumption and high in operability, and has practical production and application values.

Description

Technical Field

[0001] This invention relates to the field of solid waste resource utilization technology, and in particular to a method for comprehensive utilization of phosphorus tailings. Background Technology

[0002] Phosphate rock is an important non-metallic strategic resource in my country, widely used in phosphate fertilizers, chemicals, and other fields. my country is rich in phosphate rock resources, ranking second in the world in phosphate reserves. However, phosphate rock is generally characterized by a predominance of low-grade ore and a scarcity of high-grade ore, with medium- and low-grade phosphate rock accounting for over 90%, and an average grade of only 17%. Most phosphate rock requires beneficiation processing before it can be used to produce phosphate chemical products. During the beneficiation process, a large amount of phosphate tailings is generated, which typically contains incompletely recovered minerals such as apatite, dolomite, and calcite. However, the comprehensive utilization rate of these phosphate tailings is very low; most are simply stockpiled, occupying land and posing environmental pollution risks. The high magnesium content in phosphate tailings is a key challenge restricting their comprehensive utilization, as magnesium impurities severely affect the quality of subsequent phosphate chemical products.

[0003] Existing technologies for treating phosphorus tailings mainly include re-selection, acid leaching, and high-temperature calcination. Re-selection recovers phosphorus from tailings through flotation, but due to the low grade and complex mineral distribution of the tailings, direct flotation efficiency is low, resulting in unsatisfactory concentrate grade and recovery rate, and it cannot solve the problem of magnesium impurities. Acid leaching uses strong acids such as sulfuric acid to leach impurities such as magnesium and calcium, but this method causes the dissolution of the target component, apatite, reducing the phosphorus recovery rate, and also generates a large amount of gypsum slag, causing secondary pollution. In addition, it consumes a large amount of acid, resulting in high costs. Although high-temperature calcination can decompose carbonates, it requires heating to 900~1000℃, resulting in high energy consumption. Furthermore, the product has high activity, subsequent separation steps are complex, and it fails to achieve the targeted separation and resource utilization of magnesium, calcium, and phosphorus. For example, Chinese invention patent CN104071818A discloses a method for separating calcium, magnesium, and phosphorus elements from phosphate tailings using water leaching and ammonium leaching. The phosphate tailings are calcined at a temperature of 850–1100°C. After calcination, the dolomite CaMg(CO3)2 and calcite CaCO3 in the phosphate tailings decompose into CaO and MgO. The CaO in the calcined product is then leached with circulating water, resulting in the chemical reaction CaO + H2O = Ca(OH)2. However, the generated Ca(OH)2 is slightly soluble and requires a large amount of water for leaching. The Ca(OH)2 solution is then obtained by filtration, and finally, CO2 is introduced to generate CaCO3 to separate calcium from the phosphate tailings. This method has high roasting energy consumption, large water consumption, and high power consumption to drive the water circulation, making it not very valuable for practical industrial production.

[0004] Therefore, it is necessary to develop a new method for the comprehensive recovery of phosphorus, magnesium, and calcium elements from phosphorus tailings that is economical, efficient, and has practical production application value. Summary of the Invention

[0005] The purpose of this invention is to provide a comprehensive utilization method for phosphorus tailings in response to the above-mentioned problems. Through a combination of roasting-ammonium leaching-flotation process, the three valuable elements of phosphorus, magnesium and calcium in phosphorus tailings can be separated one by one and converted into valuable products in an economical and efficient manner. The resource utilization level is high, which can improve the comprehensive utilization value of phosphorus tailings. Moreover, it has low energy consumption, strong operability, and practical production application value.

[0006] The technical solution adopted in this invention is as follows: A method for comprehensive utilization of phosphorus tailings, comprising the following steps: S1. Roast the phosphorus tailings under medium-temperature conditions; S2. Mix the roasted product obtained in S1 with the ammonium salt solution in a certain proportion and stir thoroughly to extract magnesium from the ammonium salt. S3. The ammonium leaching product obtained in S2 is subjected to solid-liquid separation to obtain a magnesium-containing solution and filter cake; S4. The filter cake obtained from S3 filtration is subjected to positive or reverse flotation to recover apatite.

[0007] Furthermore, the intermediate temperature condition of S1 is heating at 550-650°C for 1-3 hours.

[0008] Furthermore, the ammonium salt in S2 is one or more of ammonium chloride, ammonium sulfate, or ammonium nitrate.

[0009] Furthermore, the ammonium salt added in S2 contains NH4 + With the Mg in the calcination product 2+ The molar ratio is 2.0 to 2.2:1, and the amount of water used for slurry preparation in S2 is 1.5 to 3 times the mass of the roasted product.

[0010] Furthermore, the stirring conditions in S2 are: stirring reaction at a temperature of 60-90°C for 0.5-2 hours.

[0011] Furthermore, the calcination product of S1 includes CO2, and the ammonium leaching product of S2 includes NH3; the NH3 is collected by a condensation absorption device, and after reaching a certain concentration, CO2 is introduced to obtain a (NH4)2CO3 solution.

[0012] Furthermore, the prepared (NH4)2CO3 solution is added to the magnesium-containing solution filtered in S3 to react and generate MgCO3 precipitate and ammonium salt solution.

[0013] Furthermore, S4 includes: S41. Prepare a slurry with a mass concentration of 25-30% from the filter cake obtained from S3 filtration. S42. Add a pH adjuster to adjust the pH of the slurry in S41 to a weakly acidic environment of 5.0 to 6.5; S43. Add the inhibitor of apatite and stir for 3-5 minutes; then add the collector of calcium carbonate and stir for 2-3 minutes; finally add the foaming agent. S44. Aeration is used for rough selection. The foamy product from the rough selection is scraped off to obtain calcium carbonate. The product in the rough selection tank is the apatite crude ore.

[0014] Furthermore, S4 also includes: S45. The product obtained in the roughing tank from S44 is subjected to two separate fine purification processes. The fine foam is scraped off, and the final phosphate concentrate is obtained in the fine purification tank. S46. The roughing froth product obtained in S44 is scavenged, and the scavenged froth is scraped off as the final calcium carbonate tailings. S47. Combine the first and second selected foams obtained in S45 and the product in the scavenging tank obtained in S46 as raw materials, and perform a final selection. The final selected foam scraped off enters the final calcium carbonate tailings in S46, and the product in the final selection tank enters the final phosphate concentrate in S45.

[0015] Furthermore, the inhibitor of the apatite is one or more of phosphoric acid, phosphates, or organophosphonic acids; the collector of the calcium carbonate is one or more of fatty acids (salts) and their modified products, sulfonates, or sulfates; and the foaming agent is pine oil.

[0016] In summary, due to the adoption of the above technical solution, the beneficial effects of the present invention are: 1. The method for comprehensive utilization of phosphorus tailings provided by the present invention, through a combination process of roasting-ammonia leaching-flotation, can economically and efficiently separate the three valuable elements of phosphorus, magnesium and calcium in phosphorus tailings and convert them into valuable products. It has a high degree of resource utilization and can improve the comprehensive utilization value of phosphorus tailings. In addition, it has low energy consumption, strong operability and practical production application value.

[0017] 2. The method for comprehensive utilization of phosphorus tailings provided by this invention, the method of medium-temperature roasting of phosphorus tailings, not only saves energy consumption, but also avoids the decomposition of calcium carbonate, thereby eliminating the need for subsequent complex separation of calcium elements; by combining roasting and ammonium leaching processes, this invention first effectively removes and utilizes magnesium elements that affect the quality of phosphorus products in the form of soluble magnesium salts, laying the foundation for obtaining high-quality phosphorus concentrate in the future.

[0018] 3. The present invention provides a method for the comprehensive utilization of phosphorus tailings, in which carbon dioxide generated in the roasting step and ammonia generated in the ammonium leaching step are collected to generate ammonium carbonate. Ammonium carbonate can be used as a precipitant for magnesium-containing solutions to process magnesium carbonate products, and the resulting ammonium salt solution can be recycled in the ammonium leaching step, reducing production costs.

[0019] 4. The comprehensive utilization method for phosphorus tailings provided by this invention achieves the harmless and resource-based disposal of phosphorus tailings, resulting in significant economic benefits. After separation and recovery, the P2O5 grade can reach 29.5%, and the recovery rate can reach 92.6%, which can be applied to the production of phosphorus chemical products; the magnesium ion leaching rate can reach over 90%, and the magnesium salts can be further processed into chemical products such as magnesium hydroxide or magnesium carbonate; the CaCO3 content in the calcium carbonate tailings can reach over 90%, and the calcium carbonate product can be used as a building material. The recovery and separation process essentially achieves zero wastewater, waste residue, and harmful exhaust gas emissions, making it environmentally friendly. Attached Figure Description

[0020] The present invention will be described by way of example and with reference to the accompanying drawings, wherein: Figure 1 This is a process flow diagram of the present invention; Figure 2 This is a process flow diagram of the flotation process of the present invention. Detailed Implementation

[0021] The present invention will now be described in detail with reference to the accompanying drawings.

[0022] All features disclosed in this specification, or all steps in all disclosed methods or processes, may be combined in any way, except for mutually exclusive features and / or steps.

[0023] Any feature disclosed in this specification, unless otherwise stated, may be replaced by other equivalent or similar features. That is, unless otherwise stated, each feature is merely one example of a series of equivalent or similar features.

[0024] A method for comprehensive utilization of phosphorus tailings, such as Figure 1 As shown, it includes the following steps: S1. Roast the phosphorus tailings under medium-temperature conditions; S2. Mix the roasted product obtained in S1 with the ammonium salt solution in a certain proportion and stir thoroughly to extract magnesium from the ammonium salt. S3. The ammonium leaching product obtained in S2 is subjected to solid-liquid separation to obtain a magnesium-containing solution and filter cake; S4. The filter cake obtained from S3 filtration is subjected to flotation or reverse flotation to recover apatite.

[0025] In S1 of this invention, the main gangue mineral in the phosphate tailings, dolomite CaMg(CO3)2, undergoes selective thermal decomposition. The MgCO3 component decomposes into highly reactive MgO and releases CO2, while CaCO3 remains essentially unchanged at this temperature. Apatite Ca5(PO4)3F remains stable. Compared to the existing method of calcining phosphate tailings at high temperatures of 900-1000℃, this method of medium-temperature roasting not only saves energy but also avoids the decomposition of calcium carbonate, preventing calcium from affecting the subsequent separation of magnesium and eliminating the need for complex calcium separation. Furthermore, the CO2 generated during roasting can be recovered and reused, which helps reduce carbon emissions. The chemical reactions that occur are as follows: CaMg(CO3)2= CaCO3+ MgO + CO2↑ The key to S2 of this invention lies in the selective leaching of magnesium, while CaCO3 and apatite do not participate in the reaction and remain in the solid phase. This invention first effectively removes and utilizes magnesium, which affects the quality of phosphate products, in the form of soluble magnesium salts, laying the foundation for obtaining high-quality phosphate concentrate. Furthermore, the NH3 generated during this process can be collected and recycled using a condensation absorption device. The chemical reactions that occur are as follows: MgO + 2NH4Cl = MgCl2 + 2NH3↑ + H2O or MgO + (NH4)2SO4 = MgSO4 + 2NH3↑ + H2O or MgO + 2NH4NO3= Mg(NO3)2+ 2NH3↑ + H2O In S4 of this invention, phosphate concentrate can be recovered by adding a calcium carbonate depressant and an apatite collector through positive flotation. Alternatively, phosphate concentrate can be recovered by adding an apatite depressant and a calcium carbonate collector through reverse flotation.

[0026] This invention utilizes a combined roasting-ammonia leaching-flotation process to economically and efficiently separate phosphorus, magnesium, and calcium from phosphate tailings and convert them into valuable products. It achieves a high degree of resource utilization, enhancing the comprehensive utilization value of phosphate tailings. Furthermore, it features low energy consumption, strong operability, and practical production application value. After separation and recovery, the P2O5 grade can reach 29.5%, with a recovery rate of 92.6%, suitable for the production of phosphate chemical products. The magnesium ion leaching rate can reach over 90%, and magnesium salts can be further processed into magnesium hydroxide or magnesium carbonate and other chemical products. The CaCO3 content in the calcium carbonate tailings can reach over 90%, and the calcium carbonate product can be used as a building material. The recovery and separation process essentially achieves zero wastewater, waste residue, and harmful exhaust gas emissions, making it environmentally friendly.

[0027] In one alternative implementation, the intermediate temperature condition of S1 is heating at 550–650°C for 1–3 hours.

[0028] In one alternative embodiment, the ammonium salt in S2 is one or more of ammonium chloride, ammonium sulfate, or ammonium nitrate.

[0029] In one alternative embodiment, the ammonium salt added in S2 contains NH4 + With the Mg in the calcination product 2+ The molar ratio is 2.0 to 2.2:1, and the amount of water used for slurry preparation in S2 is 1.5 to 3 times the mass of the roasted product.

[0030] In one alternative implementation, the stirring conditions in S2 are: stirring the reaction at a temperature of 60–90°C for 0.5–2 hours.

[0031] In one alternative embodiment, the calcination product of S1 includes CO2, and the ammonium leaching product of S2 includes NH3; the NH3 is collected by a condensation absorption device, and after reaching a certain concentration, CO2 is introduced to obtain a (NH4)2CO3 solution. The chemical reaction that occurs is as follows: 2NH3 + H2O + CO2 = (NH4)2CO3 In one alternative embodiment, the prepared (NH4)2CO3 solution is added to the magnesium-containing solution filtered in step S3 to react and generate MgCO3 precipitate and an ammonium salt solution. The carbon dioxide produced in the roasting step and the ammonia produced in the ammonium leaching step are collected to generate ammonium carbonate. The ammonium carbonate can be used as a precipitant for the magnesium-containing solution in the processing into magnesium carbonate products, and the resulting ammonium salt solution can be recycled in the ammonium leaching step, reducing production costs. The reaction that occurs is as follows: MgCl2 + (NH4)2CO3 = MgCO3↓ + 2NH4Cl or MgSO4 + (NH4)2CO3 = MgCO3↓ + (NH4)2SO4 or Mg(NO3)2+ (NH4)2CO3= MgCO3↓ + 2NH4NO3 In one alternative implementation, S4 includes: S41. Prepare a slurry with a mass concentration of 25-30% from the filter cake obtained from S3 filtration. S42. Add a pH adjuster to adjust the pH of the slurry in S41 to a weakly acidic environment of 5.0 to 6.5; S43. Add the inhibitor of apatite and stir for 3-5 minutes; then add the collector of calcium carbonate and stir for 2-3 minutes; finally add the foaming agent. S44. Aeration is used for rough selection. The foamy product from the rough selection is scraped off to obtain calcium carbonate. The product in the rough selection tank is the apatite crude ore.

[0032] In one alternative implementation, S4 further includes: S45. The product obtained in the roughing tank from S44 is subjected to two separate refining processes. The refining foam is scraped off, and the final phosphate concentrate is obtained in the refining tank. A small amount of apatite inhibitor needs to be added before the first refining, and no reagent is added for the second refining. S46. The roughing froth product obtained in S44 is scavenged, and the scavenged froth is scraped off as the final calcium carbonate tailings. A small amount of calcium carbonate collector needs to be added before scavenging. S47. The first and second selected foams obtained in S45, and the product from the scavenging tank obtained in S46 are combined as raw materials for a final selection. The scraped-off final selected foam enters the final calcium carbonate tailings of S46, and the product from the final selection tank enters the final phosphate concentrate of S45. During the final selection, reagents are added according to the roughing reagent dosage ratio. As another implementation, the first and second selected foams and the product from the scavenging tank can also be returned to the roughing process.

[0033] Following a process flow of one roughing, two cleaning, one scavenging, and one final flotation, the P2O5 grade can reach 29.5%, the recovery rate can reach 92.6%, and the CaCO3 content in the calcium carbonate tailings can reach over 90%. In specific industrial production processes, if the phosphorus content in the raw material phosphate tailings is high, multiple flotation processes are unnecessary; the number of flotation stages can be selected based on the actual situation.

[0034] In one alternative embodiment, the inhibitor of the apatite is one or more of phosphoric acid, phosphates, or organophosphonic acids; the collector of the calcium carbonate is one or more of fatty acids (salts) and their modified products, sulfonates, or sulfates; and the foaming agent is pine oil.

[0035] Example 1 Phosphate tailings sample A has a P2O5 content of 8.5 wt%, an MgO content of 14.3 wt%, a CaO content of 30.6 wt%, and a moisture content of 9.8 wt%. It should be recycled according to the following steps: S1. Roasting: Take 1 kg of the phosphate tailings and place it in a muffle furnace. Roast at 650℃ for 1.0 h.

[0036] S2, Ammonium leaching: The mass of the mineral after roasting is 747.9g. (Molar ratio of NH4) + Mg 2+Take 471g of ammonium sulfate (AR) at a ratio of 2.0, then add 1496g of water (twice the mass of the calcined material), stir well, then add the calcined material, and place in an 80℃ water bath with stirring for 1 hour. The ammonia gas produced during the reaction is absorbed by water after condensation.

[0037] S3. Solid-liquid separation: After the reaction is complete, filter while hot, then wash the filter cake with water to obtain a magnesium sulfate solution with a magnesium ion concentration of 45.6 g / L and a magnesium ion leaching rate of 90.5%; the dried filter cake weighs 616.1 g.

[0038] S4. Flotation: Prepare 3 kg of filter cake raw material using the above experimental conditions, and conduct a simulated closed-circuit flotation experiment according to the process of one roughing, two cleaning, one scavenging, and one final selection.

[0039] S41. Slurry preparation: Prepare the filter cake into a slurry with a mass concentration of 30%.

[0040] S41. Adjust pH: Adjust the pH of the slurry to 5.0 using sulfuric acid.

[0041] S43. Adding reagents: Add phosphoric acid (50g / t) and sodium hexametaphosphate (100g / t) sequentially to suppress apatite and stir for 4 minutes. Then add tall oil (250g / t) and sodium petroleum sulfonate (60g / t) and stir for 2 minutes. Finally, add a small amount of pine oil (30g / t). The unit for flotation reagent dosage, g / t, indicates the amount of reagent consumed per ton of filter cake (excluding water).

[0042] S44. Roughing: Aeration is used for roughing. The foamy product from the roughing process is scraped off to obtain calcium carbonate. The product in the roughing tank is the apatite crude ore.

[0043] S45. Fine Concentration: The product in the roughing tank undergoes two fine concentration processes in sequence, and the fine concentration foam is scraped off. The final phosphate concentrate is obtained in the fine concentration tank. A small amount of sodium hexametaphosphate (20g / t) needs to be added before the first fine concentration. No reagents are added for the second fine concentration.

[0044] S46. Scavenging: Scavenging is performed on the roughing froth. Before scavenging, tall oil collector (50g / t) is added. The scavenged froth is the final calcium carbonate tailings.

[0045] S47. Final Selection: The product in the scavenging tank, the foam from the primary and secondary selection tanks are combined as raw materials and subjected to a final selection according to the ratio of roughing reagent dosage. The resulting foam and the final selection tank product are respectively fed into the final calcium carbonate tailings and the final phosphate concentrate.

[0046] The filter cake raw material, after being processed by the above flotation process, yielded a final phosphate concentrate with a dry weight of 1394.1g, a P2O5 grade of 25.8%, and a recovery rate of 86.9%; the calcium carbonate tailings had a dry weight of 1606.4g and a CaCO3 content of 89.1%.

[0047] Example 2 Phosphate tailings sample B has a P2O5 content of 7.2 wt%, an MgO content of 15.6 wt%, a CaO content of 31.5 wt%, and a moisture content of 9.5 wt%. It should be recycled according to the following steps: S1. Roasting: Take 1 kg of the phosphate tailings and place it in a muffle furnace. Roast at 550℃ for 3.0 h.

[0048] S2, Ammonium leaching: The mass of the calcined mineral is 750.5g. (Molar ratio of NH4) + Mg 2+ Take 458g of ammonium chloride (AR) at a ratio of 2.2, then add 2252g of water (3 times the weight of the calcined material), stir well, then add the calcined material, and place in a 60℃ water bath with stirring for 2 hours. The ammonia gas produced during the reaction is absorbed by water after condensation.

[0049] S3. Solid-liquid separation: After the reaction is complete, filter while hot, then wash the filter cake with water to obtain a magnesium chloride solution with a magnesium ion concentration of 34.0 g / L and a magnesium ion leaching rate of 88.6%; the dried filter cake weighs 610.7 g.

[0050] S4. Flotation: Prepare 3 kg of filter cake raw material using the above experimental conditions, and conduct a simulated closed-circuit flotation experiment according to the process of one roughing, two cleaning, one scavenging, and one final selection.

[0051] S41. Slurry preparation: Prepare the filter cake into a slurry with a mass concentration of 25%.

[0052] S41. Adjust pH: Use sulfuric acid to adjust the pH of the slurry to 4.5.

[0053] Subsequent steps S43-S47, the dosage ratio, and the stirring time are the same as in Example 1.

[0054] The filter cake raw material, after being processed by the above flotation process, yielded a final phosphate concentrate with a dried mass of 1194.2g, a P2O5 grade of 24.7%, and a recovery rate of 83.4%; the dried calcium carbonate tailings had a dried mass of 1808.7g and a CaCO3 content of 84.5%.

[0055] Example 3 Phosphate tailings sample C has a P2O5 content of 10.6 wt%, an MgO content of 13.8 wt%, a CaO content of 32.9 wt%, and a moisture content of 10.2 wt%. It should be recycled according to the following steps: S1. Roasting: Take 1 kg of the phosphate tailings and place it in a muffle furnace. Roast at 600℃ for 2.0 h.

[0056] S2, Ammonium leaching: The mass of the mineral after roasting is 757.9g. (Molar ratio of NH4) + Mg 2+ Take 386g of ammonium chloride (AR) at a ratio of 2.1, then add 1137g of water (1.5 times the weight of the calcined material), stir well, then add the calcined material, and place in a 90℃ water bath with stirring for 0.5h. The ammonia gas produced during the reaction is absorbed by water after condensation.

[0057] S3. Solid-liquid separation: After the reaction is complete, filter while hot, then wash the filter cake with water to obtain a magnesium chloride solution with a magnesium ion concentration of 54.7 g / L and a magnesium ion leaching rate of 87.4%; the dried filter cake weighs 635.5 g.

[0058] S4. Flotation: Prepare 3 kg of filter cake raw material using the above experimental conditions, and conduct a simulated closed-circuit flotation experiment according to the process of one roughing, two cleaning, one scavenging, and one final selection.

[0059] S41. Slurry preparation: Prepare the filter cake into a slurry with a mass concentration of 28%.

[0060] S41. Adjust pH: Use sulfuric acid to adjust the pH of the slurry to 5.5.

[0061] S43. Adding chemicals: Add phosphoric acid (50g / t) and aminotrimethylene phosphonic acid (70g / t) in sequence to inhibit apatite and stir for 4 minutes. Then add the collector tall oil (260g / t) and sodium petroleum sulfonate (50g / t) and stir for 2 minutes. Finally, add a small amount of pine oil (30g / t).

[0062] S44. Roughing: Aeration is used for roughing. The foamy product from the roughing process is scraped off to obtain calcium carbonate. The product in the roughing tank is the apatite crude ore.

[0063] S45. Fine Concentration: The product in the coarse separation tank undergoes two fine concentrations in sequence, and the fine concentration foam is scraped off. The final phosphate concentrate is obtained in the fine concentration tank. A small amount of aminotrimethylene phosphonic acid (20g / t) needs to be added before the first fine concentration. No reagent is added for the second fine concentration.

[0064] S46. Scavenging: Scavenging is performed on the roughing froth. Before scavenging, tall oil collector (50g / t) is added. The scavenged froth is the final calcium carbonate tailings.

[0065] S47. Final Selection: The product in the scavenging tank, the foam from the primary cleaning and the foam from the secondary cleaning are combined as raw materials and subjected to flotation again according to the ratio of roughing reagent dosage. The resulting foam and the product from the final selection tank are respectively fed into the final calcium carbonate tailings and the final phosphate concentrate.

[0066] The filter cake raw material, after being processed by the above flotation process, yielded a final phosphate concentrate with a dry weight of 1570.6g, a P2O5 grade of 29.5%, and a recovery rate of 92.6%; the calcium carbonate tailings had a dry weight of 1432.7g and a CaCO3 content of 90.6%.

[0067] Comparative Example 1 Phosphate tailings sample A has a P2O5 content of 8.5 wt%, an MgO content of 14.3 wt%, a CaO content of 30.6 wt%, and a moisture content of 9.8 wt%. It should be recycled according to the following steps: S1. Roasting: Take 1 kg of the phosphate tailings and place it in a muffle furnace. Roast at 500℃ for 3.0 h.

[0068] S2, Ammonium leaching: The mass of the calcined mineral is 850.3g. (Molar ratio of NH4) + Mg 2+ Take 471g of ammonium sulfate (AR) at a ratio of 2.0, then add 1701g of water (twice the mass of the calcined material), stir well, then add the calcined material, and place in an 80℃ water bath with stirring for 1 hour. The ammonia gas produced during the reaction is absorbed by water after condensation.

[0069] S3. Solid-liquid separation: After the reaction is complete, filter while hot, then wash the filter cake with water to obtain a magnesium sulfate solution with a magnesium ion concentration of 14.4 g / L and a magnesium ion leaching rate of 32.0%; the dried filter cake weighs 805.5 g.

[0070] S4. Flotation: Prepare 3 kg of filter cake raw material using the above experimental conditions, and conduct a simulated closed-circuit flotation experiment according to the process of one roughing, two cleaning, one scavenging, and one final selection.

[0071] Subsequent steps S41-S47, the dosage ratio, and the stirring time are the same as in Example 1.

[0072] The filter cake raw material, after being processed by the above flotation process, yielded a final phosphate concentrate with a dried mass of 1151.2g, a P2O5 grade of 14.6%, and a recovery rate of 53.1%; the dried calcium carbonate tailings had a dried mass of 1850.6g and a CaCO3 content of 52.3%.

[0073] Comparative Example 2 Phosphate tailings sample A has a P2O5 content of 8.5 wt%, an MgO content of 14.3 wt%, a CaO content of 30.6 wt%, and a moisture content of 9.8 wt%. It should be recycled according to the following steps: S1. Roasting: Take 1 kg of the phosphate tailings and place it in a muffle furnace. Roast at 700℃ for 1.0 h.

[0074] S2, Ammonium leaching: The mass of the calcined mineral is 693.7g. (Molar ratio of NH4) + Mg 2+ Take 471g of ammonium sulfate (AR) at a ratio of 2.0, then add 1387g of water (twice the mass of the calcined material), stir well, then add the calcined material, and place in an 80℃ water bath with stirring for 1 hour. The ammonia gas produced during the reaction is absorbed by water after condensation.

[0075] S3. Solid-liquid separation: After the reaction, the mixture was filtered while hot, and the filter cake was washed with water to obtain a magnesium sulfate solution with a magnesium ion concentration of 35.4 g / L and a magnesium ion leaching rate of 64.8%. The dried filter cake weighed 703.9 g. The low magnesium ion leaching rate is because some of the calcium carbonate in the phosphate tailings sample decomposed during roasting to form calcium oxide, consuming some ammonium salts.

[0076] S4. Flotation: Prepare 3 kg of filter cake raw material using the above experimental conditions, and conduct a simulated closed-circuit flotation experiment according to the process of one roughing, two cleaning, one scavenging, and one final selection.

[0077] Subsequent steps S41-S47, the dosage ratio, and the stirring time are the same as in Example 1.

[0078] The filter cake raw material, after being processed by the above flotation process, yielded a final phosphate concentrate with a dry weight of 1316.9g, a P2O5 grade of 18.1%, and a recovery rate of 65.8%; the calcium carbonate tailings had a dry weight of 1683.9g and a CaCO3 content of 54.2%.

[0079] Comparative Example 3 Phosphate tailings sample A has a P2O5 content of 8.5 wt%, MgO content of 14.3 wt%, CaO content of 30.6 wt%, and a moisture content of 9.8 wt%. It was recovered and utilized using direct flotation. A 3kg sample was taken, and a simulated closed-circuit flotation experiment was conducted following a process of one roughing, one scavenging, and two cleaning stages. The raw material was prepared into a 30% (w / w) slurry, and the pH was adjusted to 5.0 with sulfuric acid. Phosphoric acid (50g / t) and sodium hexametaphosphate (100g / t) were added sequentially to suppress the apatite, and the mixture was stirred for 4 minutes. Then, tall oil (250g / t) and sodium petroleum sulfonate (60g / t) were added as collectors, and the mixture was stirred for 2 minutes. Finally, a small amount of pine oil (30g / t) was added. After slurry preparation, the slurry entered the roughing stage. The rough concentrate obtained in the roughing cell underwent two cleaning stages to obtain the final phosphate concentrate. A small amount of sodium hexametaphosphate (20g / t) was added before the first cleaning stage, and no reagents were added during the second cleaning stage. The roughing froth was then scavenged, with tall oil (50g / t) added before scavenging. The scavenged froth was the final calcium carbonate tailings. The product from the scavenging tank, the foam from the primary cleaning and the foam from the secondary cleaning are combined as raw materials. A final cleaning is then performed according to the ratio of the roughing reagent dosage. The resulting foam and the final cleaning tank product are then fed into the final calcium carbonate tailings and the final phosphate concentrate, respectively.

[0080] The filter cake raw material, after being processed by the above flotation process, yielded a final phosphate concentrate with a dry weight of 1016.0g, a P2O5 grade of 12.7%, and a recovery rate of 50.6%; the flotation tailings had a dry weight of 1690.5g and a CaCO3 content of 51.0%.

[0081] This invention is not limited to the specific embodiments described above. The invention extends to any new feature or combination disclosed in this specification, as well as any new method or process step or combination disclosed herein.

Claims

1. A method for comprehensive utilization of phosphorus tailings, characterized in that: Includes the following steps: S1. Roast the phosphorus tailings under medium-temperature conditions; S2. Mix the roasted product obtained in S1 with the ammonium salt solution in a certain proportion and stir thoroughly to extract magnesium from the ammonium salt. S3. The ammonium leaching product obtained in S2 is subjected to solid-liquid separation to obtain a magnesium-containing solution and filter cake; S4. The filter cake obtained from S3 filtration is subjected to positive or reverse flotation to recover apatite.

2. The method for comprehensive utilization of phosphorus tailings as described in claim 1, characterized in that: The intermediate temperature condition for S1 is heating at 550–650°C for 1–3 hours.

3. The method for comprehensive utilization of phosphorus tailings as described in claim 1, characterized in that: The ammonium salt in S2 is one or more of ammonium chloride, ammonium sulfate, or ammonium nitrate.

4. The method for comprehensive utilization of phosphorus tailings as described in claim 1, characterized in that: The ammonium salt added in S2 contains NH4 + With the Mg in the calcination product 2+ The molar ratio is 2.0 to 2.2:1, and the amount of water used for slurry preparation in S2 is 1.5 to 3 times the mass of the roasted product.

5. The method for comprehensive utilization of phosphorus tailings as described in claim 1, characterized in that: The stirring conditions in S2 are: stirring reaction at 60-90°C for 0.5-2 hours.

6. The method for comprehensive utilization of phosphorus tailings as described in claim 1, characterized in that: The roasting product of S1 includes CO2, and the ammonium leaching product of S2 includes NH3; the NH3 is collected by a condensation absorption device, and after reaching a certain concentration, CO2 is introduced to obtain (NH4)2CO3 solution.

7. The method for comprehensive utilization of phosphorus tailings as described in claim 6, characterized in that: The prepared (NH4)2CO3 solution is added to the magnesium-containing solution filtered by S3 to react and generate MgCO3 precipitate and ammonium salt solution.

8. The method for comprehensive utilization of phosphorus tailings as described in claim 1, characterized in that: S4 includes: S41. Prepare a slurry with a mass concentration of 25-30% from the filter cake obtained from S3 filtration. S42. Add a pH adjuster to adjust the pH of the slurry in S41 to a weakly acidic environment of 5.0 to 6.5; S43. Add the inhibitor of apatite and stir for 3-5 minutes; then add the collector of calcium carbonate and stir for 2-3 minutes; finally add the foaming agent. S44. Aeration is used for rough selection. The foamy product from the rough selection is scraped off to obtain calcium carbonate. The product in the rough selection tank is the apatite crude ore.

9. The method for comprehensive utilization of phosphorus tailings as described in claim 8, characterized in that: S4 further includes: S45. The product obtained in the roughing tank from S44 is subjected to two separate fine purification processes. The fine foam is scraped off, and the final phosphate concentrate is obtained in the fine purification tank. S46. The roughing froth product obtained in S44 is scavenged, and the scavenged froth is scraped off as the final calcium carbonate tailings. S47. Combine the first and second selected foams obtained in S45 and the product in the scavenging tank obtained in S46 as raw materials, and perform a final selection. The final selected foam scraped off enters the final calcium carbonate tailings in S46, and the product in the final selection tank enters the final phosphate concentrate in S45.

10. The method for comprehensive utilization of phosphorus tailings as described in claim 8, characterized in that: The inhibitor of the apatite is one or more of phosphoric acid, phosphates, or organophosphonic acids; the collector of the calcium carbonate is one or more of fatty acids (salts) and their modified products, sulfonates, or sulfates; and the foaming agent is pine oil.

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