A method of beneficiating phosphate concentrate from phosphate tailings
By crushing, grinding, and calcining phosphorus tailings, followed by jigging and flotation to remove magnesium and silicon, the problems of low phosphorus tailings separation efficiency and low phosphorus yield are solved, achieving efficient and economical phosphorus concentrate separation, which is suitable for resource recovery in phosphorus chemical enterprises.
Patent Information
- Authority / Receiving Office
- CN · China
- Patent Type
- Patents(China)
- Current Assignee / Owner
- GUANGDONG BRUNP RECYCLING TECH CO LTD
- Filing Date
- 2023-04-28
- Publication Date
- 2026-06-09
AI Technical Summary
Existing technologies for phosphorus tailings have low sorting efficiency, low phosphorus yield, and poor economic viability. They also suffer from secondary pollution and complex processes, making them unsuitable for widespread application.
The mineral is processed by crushing, grinding, and calcining to form a slurry for jigging beneficiation. Combined with flotation to remove magnesium and silicon, the jigging machine and flotation cell are used for gravity separation and reverse flotation to remove impurities such as dolomite and quartz, thereby improving the grade and yield of phosphate concentrate.
It improves the separation efficiency and yield of phosphate concentrate in phosphate tailings, the process is easy to operate and promotes its use, and reduces energy consumption and secondary pollution.
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Figure CN117136102B_ABST
Abstract
Description
Technical Field
[0001] This disclosure belongs to the field of phosphorus resource recovery technology, and specifically relates to a method for separating phosphorus concentrate from phosphorus tailings. Background Technology
[0002] Phosphate tailings are solid waste generated in the first stage of phosphate rock processing in phosphate chemical production. High-grade phosphate rock with a grade of 30 wt% or higher and an MgO content of less than 1.0 wt% can be directly used in phosphoric acid production after crushing and grinding. However, most phosphate rock has a grade of less than 30 wt% and an MgO content of more than 1.0 wt%. To meet the requirements of wet-process phosphoric acid production, it is necessary to go through the mineral processing stage.
[0003] Typically, phosphate chemical enterprises need to perform ore blending operations to balance the grade and MgO content of the raw ore within a certain range. After crushing, grinding, and flotation, phosphate concentrate and phosphate tailings are produced. The phosphate concentrate is then used in the next process of phosphoric acid production, while the phosphate tailings are discharged as waste into tailings ponds. The P2O5 content in the tailings is approximately 4wt%-10wt%, and the magnesium oxide content is approximately 16wt%-18wt%. How to fully recover and utilize the residual phosphorus in the tailings is an important way to improve the comprehensive utilization rate of resources.
[0004] Currently, technologies for the separation and reuse of phosphate tailings include acid dissolution and flotation, for the production of magnesium salts, magnesium-phosphate binary compound fertilizers, and as raw materials for building materials. However, these processes all have a series of problems. Acid dissolution of phosphate tailings easily generates large amounts of CO2, causes severe foaming, requires large amounts of acid, has high economic costs, and demands stringent equipment requirements. Direct flotation is inefficient in separating phosphate tailings. Overall, current separation processes suffer from low separation efficiency, low phosphorus yield, poor economics, and also suffer from serious secondary pollution and complex processes, hindering their widespread application. Summary of the Invention
[0005] This disclosure aims to address at least one of the technical problems existing in the related art. To this end, this disclosure proposes a method for separating phosphate concentrate from phosphate tailings, which has high separation efficiency, high phosphate recovery rate, and is easy to operate and promote.
[0006] The above-mentioned technical objective of this disclosure is achieved through the following technical solution:
[0007] A method for separating phosphate concentrate from phosphate tailings includes the following steps:
[0008] (1) The phosphate tailings are crushed, ground, and calcined, then mixed into a slurry for jigging beneficiation to obtain primary sediment, secondary sediment, and overflow material, wherein the density of the primary sediment is 2.95-3.25 t / m³. 3The density of the secondary precipitate is 2.85-2.95 t / m³. 3 The density of the overflow material is less than 2.85 t / m³. 3 ;
[0009] (2) After preparing the secondary precipitate material obtained in step (1) into a slurry, magnesium and silicon are removed by flotation to obtain the flotated slurry;
[0010] (3) The primary precipitate obtained in step (1) and the flotation slurry obtained in step (2) are filtered and dried to obtain the phosphate concentrate.
[0011] In one embodiment, in step (1), the calcination temperature is 300-400°C, and the calcination time is 3-8 hours.
[0012] In one embodiment, in step (1), the calcination temperature is 300-350°C and the calcination time is 4-6 hours.
[0013] In one embodiment, the calcination temperature is 310°C and the calcination time is 5 hours.
[0014] In one embodiment, the phosphorus tailings are phosphorus tailings produced in the reverse flotation magnesium removal process of a phosphorus chemical enterprise, with a -200 mesh fineness content of more than 80 wt%, where the -200 mesh fineness content refers to the content of materials with a fineness of less than 200 mesh.
[0015] In one embodiment, the flotation in step (2) includes the following steps:
[0016] S1 prepares the secondary precipitate into a slurry, adjusts the pH of the slurry to acidic, adds a fatty acid anion collector, and performs reverse flotation to remove magnesium, thus obtaining the slurry after primary flotation.
[0017] S2 adjusts the pH of the pulp obtained after the first flotation to alkaline, adds an amine cationic collector, and performs reverse flotation to remove silica, thus obtaining the pulp after the second flotation.
[0018] In one embodiment, in step S1, adjusting the pH of the slurry to acidic means adjusting the pH of the slurry to 4-6.
[0019] In one embodiment, in step S1, adjusting the pH of the slurry to acidic means adjusting the pH of the slurry to 4.5-5.5.
[0020] In one embodiment, in step S1, the fatty acid anion collector is oleic acid soap, and the ratio of the amount of fatty acid anion collector added to the phosphate tailings is (0.2-1.0) kg: 1 t. The oleic acid soap is prepared by mixing oleic acid and linoleic acid at a mass ratio of 2:3, and then adding sodium hydroxide powder at a mass ratio of fatty acid: sodium hydroxide: water = 9:1:10. The temperature is controlled at 80°C, and the stirring speed is 140±5 r / min for saponification.
[0021] In one embodiment, the saponification time is 90-120 min. In another embodiment, the ratio of the amount of fatty acid anionic collector added to the amount of phosphate tailings is (0.5-0.7) kg: 1 t.
[0022] In one embodiment, in step S2, adjusting the pH of the obtained pulp after primary flotation to alkaline means adjusting the pH of the pulp to 7.5-10.
[0023] In one embodiment, in step S2, adjusting the pH of the obtained pulp after primary flotation to alkaline means adjusting the pH of the pulp to 8-9.
[0024] In one embodiment, in step S2, the amine cationic collector is dodecylamine, and the ratio of the amount of the amine cationic collector added to the phosphate tailings is (0.1-0.5) kg: 1 t.
[0025] In one embodiment, the ratio of the amount of the amine cationic collector added to the amount of phosphate tailings is (0.3-0.4) kg: 1 t.
[0026] In one embodiment, in step (1), the particle size of the material obtained after crushing is less than 2 mm.
[0027] In one embodiment, in step (1), the material obtained after grinding has a -200 mesh content of 85wt%-95wt%, where the -200 mesh content refers to the content of material with a mesh size of less than 200.
[0028] In one embodiment, in step (1), the material obtained after grinding has a -200 mesh content of 90wt%-95wt%.
[0029] In one embodiment, in step (1), the concentration of the slurry is 20wt%-40wt%.
[0030] In one embodiment, in step (1), the concentration of the slurry is 28wt%-30wt%.
[0031] In one embodiment, in step (1), the amount of water flowing through the jigging process is 1-5 m³. 3 / h, stroke is 10-30mm, strokes are 300-450 times / min.
[0032] In one embodiment, in step (1), the amount of water flowing through the jigging process is 3-5 m³. 3 / h, stroke is 20-30mm, strokes are 320-420 times / min.
[0033] In one embodiment, in step (2), the concentration of the secondary precipitate material after being prepared into a slurry is 20wt%-40wt%.
[0034] In one embodiment, in step (2), the concentration of the secondary precipitate after being prepared into a slurry is 28wt%-30wt%.
[0035] In one embodiment, the phosphorus tailings are derived from phosphorus tailings produced in a reverse flotation magnesium removal process.
[0036] In one embodiment, a method for separating phosphate concentrate from phosphate tailings includes the following steps:
[0037] (1) Break up the phosphate tailings agglomerates to a particle size of less than 2 mm;
[0038] (2) Put the crushed phosphorus tailings into a rod mill, add an appropriate amount of water for grinding, so that the -200 mesh fineness content reaches 90wt%-95wt%, filter and dry.
[0039] (3) Place the dried tailings into an oven at 300-350℃ and calcine them to completely remove the reagents adhering to the surface of the tailings particles.
[0040] (4) After the calcined tailings have cooled, add water to prepare a slurry with a concentration of 30wt%, add it to a jig for jigging, adjust the amount of water under the screen and the number of flushes to ensure that the dolomite tailings flow out from the overflow outlet.
[0041] The jig produces three types of materials:
[0042] ① Primary sediment material, density 2.95-3.25 t / m³ 3 ;
[0043] ② Secondary precipitate material, with a density of 2.85-2.95 t / m³ 3 Because it contains impurities such as quartz and dolomite, it is a low-grade concentrate, which is filtered and dried.
[0044] ③ Overflow material with a density of less than 2.85 t / m³ 3 It is a fine-grained dolomite mineral, which is filtered and dried.
[0045] (5) Pour the secondary sediment into the flotation tank for resizing. The flotation concentration is 28wt%-30wt%. Add dilute phosphoric acid to adjust the pH to 4.5-5.5. Add fatty acid anionic collector, which is oleic acid soap, and carry out reverse flotation to remove magnesium. The foam generated by flotation is magnesium-containing tailings.
[0046] (6) Continue to add sodium carbonate to the slurry in the tank to adjust the pH to 8-9, add amine cationic collector, the amine cationic collector is dodecylamine, and carry out reverse flotation to remove silicon. The foam produced by flotation is silicon-containing tailings.
[0047] (7) The primary precipitate obtained in step (4) and the flotation slurry obtained in step (6) are filtered and dried to obtain phosphate concentrate.
[0048] The beneficial effects of this disclosure are:
[0049] (1) The phosphorus tailings mentioned in the method for separating phosphorus concentrate from phosphorus tailings in this disclosure are phosphorus tailings produced in the reverse flotation magnesium removal process of phosphorus chemical enterprises. The -200 mesh fineness content is above 80wt%, and the tailings are naturally agglomerated when piled up in the tailings pond. They need to be crushed and dispersed into fine particles before calcination, which is beneficial to improving calcination efficiency.
[0050] (2) In the method for separating phosphate concentrate from phosphate tailings disclosed herein, the calcination temperature is controlled at 300-400℃. Because the phosphate tailings produced by phosphate chemical enterprises have a large amount of fatty acid collectors attached to their surface, they still have a certain degree of hydrophobicity, which affects subsequent jigging and flotation. Therefore, they need to be removed. At this calcination temperature, the fatty acid collectors decompose and volatilize and are removed, while the phosphate tailings do not decompose. This achieves the purpose of removing organic matter while saving energy consumption and does not affect the surface properties of the minerals.
[0051] (3) In the method for separating phosphate concentrate from phosphate tailings disclosed herein, the use of a rod mill for grinding can effectively avoid over-grinding of phosphate tailings. Since dolomite minerals are easier to grind than phosphate minerals, dolomite minerals are distributed in a finer range, while phosphate minerals are distributed in a coarser range, which is beneficial for jigging operations.
[0052] (4) In the method of separating phosphate concentrate from phosphate tailings disclosed in this disclosure, a jig is used for mineral processing. The jig is a type of gravity separation equipment. Other gravity separation equipment includes shaking tables, sluices, and heavy media separation equipment. In this disclosure, the phosphate tailings have a fine particle size and are uniformly distributed. The theory of jig separation is to separate by specific gravity. The separation effect and recovery rate are higher than those of other equipment.
[0053] (5) In the method for separating phosphate concentrate from phosphate tailings disclosed herein, the grinding fineness to 200 mesh is preferably controlled at 90wt%-95wt%, and the dolomite and phosphate ore are basically liberated as monomers, wherein the density of the phosphate ore is 3.0-3.1 t / m³.3 Between these values, the specific gravity of quartz is 2.65 t / m³. 3 The specific gravity of dolomite is approximately 2.85-2.9 t / m³. 3 Between these, using a jig for gravity separation can effectively separate phosphate rock from quartz and dolomite.
[0054] (6) In the method for separating phosphorus concentrate from phosphorus tailings disclosed herein, low-grade concentrate is subjected to double reverse flotation to improve the grade of phosphorus concentrate, thereby increasing the phosphorus yield. At the same time, the process is easy to operate and convenient to promote and use. Attached Figure Description
[0055] Figure 1 This is a flowchart illustrating Embodiment 1 of the present disclosure. Detailed Implementation
[0056] The present disclosure will be further described below with reference to specific embodiments.
[0057] Example 1:
[0058] A method for separating phosphate concentrate from phosphate tailings, such as Figure 1 As shown, it includes the following steps:
[0059] (1) Take 1000g of phosphorus tailings lumps, crush them with a small crusher, and screen them with a 2mm aperture steel wire screen. The coarse particles are returned to the crusher for further crushing, and the final particle size is less than 2mm.
[0060] (2) Put the crushed phosphorus tailings into a rod mill, add an appropriate amount of water for grinding, so that the -200 mesh fineness content reaches 90wt%, filter and dry.
[0061] (3) Place the dried tailings in a 310℃ oven and calcine for 5 hours to completely remove the agent adhering to the surface of the tailings particles.
[0062] (4) After the calcined tailings have cooled, add water to prepare a slurry with a concentration of 28 wt%, and add it to a jig for jigging beneficiation. Adjust the water flow rate under the screen to 3 m³. 3 The stroke rate is 25 mm, and the stroke frequency is 320 strokes / min, yielding primary sediment, secondary sediment, and overflow material. The density of the primary sediment is 3.07 t / m³. 3 The density of the secondary precipitate is 2.92 t / m³. 3 The density of the overflow material is 2.75 t / m³. 3 The resulting overflow material is magnesium-containing dolomite tailings, which flow out from the overflow outlet.
[0063] (5) Pour the secondary precipitate into the flotation tank for re-slurrying. The flotation concentration is 28wt%. Add dilute phosphoric acid to adjust the pH to 5.0. Add fatty acid anionic collector oleic acid soap. The preparation method of oleic acid soap is to mix oleic acid and linoleic acid in a mass ratio of 2:3, and then add sodium hydroxide powder in a mass ratio of fatty acid: sodium hydroxide: water = 9:1:10. The temperature is controlled at 80℃ and the stirring speed is 140±5r / min for saponification for 90min. The amount of oleic acid soap added is 0.7kg / t of raw ore. Perform reverse flotation to remove magnesium. The flotation time is 3 minutes. The foam produced by flotation is magnesium-containing tailings.
[0064] (6) Continue to add sodium carbonate to the slurry in the tank to adjust the pH to 9.0, add dodecylamine, an amine cationic collector, at a dosage of 0.4 kg / t of raw ore, and carry out reverse flotation to remove silicon. The flotation time is 3 minutes, and the foam produced by flotation is silicon-containing tailings.
[0065] (7) The primary precipitate obtained in step (4) and the flotation slurry obtained in step (6) are filtered and dried to obtain phosphate concentrate.
[0066] Example 2:
[0067] A method for separating phosphate concentrate from phosphate tailings includes the following steps:
[0068] (1) Take 1000g of phosphorus tailings lumps, crush them with a small crusher, and screen them with a 2mm aperture steel wire screen. The coarse particles are returned to the crusher for further crushing, and the final particle size is less than 2mm.
[0069] (2) Put the crushed phosphorus tailings into a rod mill, add an appropriate amount of water for grinding, so that the -200 mesh fineness content reaches 95wt%, filter and dry.
[0070] (3) Place the dried tailings in a 310℃ oven and calcine for 5 hours to completely remove the agent adhering to the surface of the tailings particles.
[0071] (4) After the calcined tailings have cooled, add water to prepare a slurry with a concentration of 30 wt%, and add it to a jig for jigging beneficiation. Adjust the water flow rate under the screen to 5 m³. 3 The stroke rate is 25 mm, and the stroke frequency is 420 strokes / min, yielding primary sediment, secondary sediment, and overflow material. The density of the primary sediment is 3.11 t / m³. 3 The density of the secondary precipitate is 2.95 t / m³. 3 The density of the overflow material is 2.78 t / m³. 3 The resulting overflow material is magnesium-containing dolomite tailings, which flow out from the overflow outlet.
[0072] (5) Pour the secondary precipitate into the flotation tank for re-slurrying. The flotation concentration is 30wt%. Add dilute phosphoric acid to adjust the pH to 5.0. Add fatty acid anionic collector oleic acid soap. The preparation method of oleic acid soap is to mix oleic acid and linoleic acid in a mass ratio of 2:3, and then add sodium hydroxide powder in a mass ratio of fatty acid: sodium hydroxide: water = 9:1:10. The temperature is controlled at 80℃ and the stirring speed is 140±5r / min for saponification for 90min. The amount of oleic acid soap added is 0.5kg / t of raw ore. Perform reverse flotation to remove magnesium. The flotation time is 5 minutes. The foam produced by flotation is magnesium-containing tailings.
[0073] (6) Continue to add sodium carbonate to the slurry in the tank to adjust the pH to 9.0, add dodecylamine, an amine cationic collector, at a dosage of 0.3 kg / t of raw ore, and carry out reverse flotation to remove silicon. The flotation time is 5 minutes, and the foam produced by flotation is silicon-containing tailings.
[0074] (7) The primary precipitate obtained in step (4) and the flotation slurry obtained in step (6) are filtered and dried to obtain phosphate concentrate.
[0075] Example 3:
[0076] A method for separating phosphate concentrate from phosphate tailings includes the following steps:
[0077] (1) Take 1000g of phosphorus tailings lumps, crush them with a small crusher, and screen them with a 2mm aperture steel wire screen. The coarse particles are returned to the crusher for further crushing, and the final particle size is less than 2mm.
[0078] (2) Put the crushed phosphate tailings into a rod mill, add an appropriate amount of water for grinding, so that the -200 mesh fineness content reaches 93wt%, filter and dry.
[0079] (3) Place the dried tailings in a 310℃ oven and calcine for 5 hours to completely remove the agent adhering to the surface of the tailings particles.
[0080] (4) After the calcined tailings have cooled, add water to prepare a slurry with a concentration of 30 wt%, and add it to a jig for jigging beneficiation. Adjust the water flow rate under the screen to 5 m³. 3 The stroke rate is 25 mm, and the stroke rate is 420 strokes / min, yielding primary sediment, secondary sediment, and overflow material. The density of the primary sediment is 3.10 t / m³. 3 The density of the secondary precipitate is 2.95 t / m³. 3 The density of the overflow material is 2.80 t / m³. 3 The resulting overflow material is magnesium-containing dolomite tailings, which flow out from the overflow outlet.
[0081] (5) Pour the secondary precipitate into the flotation tank for re-slurrying. The flotation concentration is 29wt%. Add dilute phosphoric acid to adjust the pH to 5.0. Add fatty acid anionic collector oleic acid soap. The preparation method of oleic acid soap is to mix oleic acid and linoleic acid in a mass ratio of 2:3, and then add sodium hydroxide powder in a mass ratio of fatty acid: sodium hydroxide: water = 9:1:10. The temperature is controlled at 80℃ and the stirring speed is 140±5r / min for saponification for 90min. The collector dosage is 0.7kg / t of raw ore. Perform reverse flotation to remove magnesium. The flotation time is 4 minutes. The foam produced by flotation is magnesium-containing tailings.
[0082] (6) Continue to add sodium carbonate to the slurry in the tank to adjust the pH to 9.0, add dodecylamine, an amine cationic collector, at a dosage of 0.4 kg / t of raw ore, and carry out reverse flotation to remove silicon. The flotation time is 4 minutes, and the foam produced by flotation is silicon-containing tailings.
[0083] (7) The primary precipitate obtained in step (4) and the flotation slurry obtained in step (6) are filtered and dried to obtain phosphate concentrate.
[0084] Comparative Example 1: (The only difference from Example 2 is the calcination temperature of 298°C)
[0085] A method for separating phosphate concentrate from phosphate tailings includes the following steps:
[0086] (1) Take 1000g of phosphorus tailings lumps, crush them with a small crusher, and screen them with a 2mm aperture steel wire screen. The coarse particles are returned to the crusher for further crushing, and the final particle size is less than 2mm.
[0087] (2) Put the crushed phosphorus tailings into a rod mill, add an appropriate amount of water for grinding, so that the -200 mesh fineness content reaches 95wt%, filter and dry.
[0088] (3) Place the dried tailings in a 298℃ oven and calcine for 5 hours to completely remove the reagents adhering to the surface of the tailings particles.
[0089] (4) After the calcined tailings have cooled, add water to prepare a slurry with a concentration of 30 wt%, and add it to a jig for jigging beneficiation. Adjust the water flow rate under the screen to 5 m³. 3 The stroke rate is 25 mm, and the stroke frequency is 420 strokes / min, yielding primary sediment, secondary sediment, and overflow material. The density of the primary sediment is 3.01 t / m³. 3 The density of the secondary precipitate is 2.90 t / m³. 3 The density of the overflow material is 2.80 t / m³. 3 The resulting overflow material is magnesium-containing dolomite tailings, which flow out from the overflow outlet.
[0090] (5) Pour the secondary precipitate into the flotation tank for re-slurrying. The flotation concentration is 30wt%. Add dilute phosphoric acid to adjust the pH to 5.0. Add fatty acid anionic collector oleic acid soap. The preparation method of oleic acid soap is to mix oleic acid and linoleic acid in a mass ratio of 2:3, and then add sodium hydroxide powder in a mass ratio of fatty acid: sodium hydroxide: water = 9:1:10. The temperature is controlled at 80℃ and the stirring speed is 140±5r / min for saponification for 90min. The amount of oleic acid soap added is 0.5kg / t of raw ore. Perform reverse flotation to remove magnesium. The flotation time is 5 minutes. The foam produced by flotation is magnesium-containing tailings.
[0091] (6) Continue to add sodium carbonate to the slurry in the tank to adjust the pH to 9.0, add dodecylamine, an amine cationic collector, and add dodecylamine collector at a rate of 0.3 kg / t of raw ore. Perform reverse flotation to remove silicon. The flotation time is 5 minutes. The foam produced by flotation is silicon-containing tailings.
[0092] (7) The primary precipitate obtained in step (4) and the flotation slurry obtained in step (5) are filtered and dried to obtain phosphate concentrate.
[0093] Comparative Example 2: (The only difference from Example 2 is that the calcination temperature is 402°C)
[0094] A method for separating phosphate concentrate from phosphate tailings includes the following steps:
[0095] (1) Take 1000g of phosphorus tailings lumps, crush them with a small crusher, and screen them with a 2mm aperture steel wire screen. The coarse particles are returned to the crusher for further crushing, and the final particle size is less than 2mm.
[0096] (2) Put the crushed phosphorus tailings into a rod mill, add an appropriate amount of water for grinding, so that the -200 mesh fineness content reaches 95wt%, filter and dry.
[0097] (3) Place the dried tailings in a 402℃ oven and calcine for 5 hours to completely remove the reagents adhering to the surface of the tailings particles.
[0098] (4) After the calcined tailings have cooled, add water to prepare a slurry with a concentration of 30 wt%, and add it to a jig for jigging beneficiation. Adjust the water flow rate under the screen to 5 m³. 3 The stroke rate is 25 mm, and the stroke frequency is 420 strokes / min, yielding primary sediment, secondary sediment, and overflow material. The density of the primary sediment is 3.11 t / m³. 3 The density of the secondary precipitate is 2.95 t / m³. 3 The density of the overflow material is 2.78 t / m³. 3 The resulting overflow material is magnesium-containing dolomite tailings, which flow out from the overflow outlet.
[0099] (5) Pour the secondary precipitate into the flotation tank for re-slurrying. The flotation concentration is 30wt%. Add dilute phosphoric acid to adjust the pH to 5.0. Add fatty acid anionic collector oleic acid soap. The preparation method of oleic acid soap is to mix oleic acid and linoleic acid in a mass ratio of 2:3, and then add sodium hydroxide powder in a mass ratio of fatty acid: sodium hydroxide: water = 9:1:10. The temperature is controlled at 80℃ and the stirring speed is 140±5r / min for saponification for 90min. The amount of oleic acid soap added is 0.5kg / t of raw ore. Perform reverse flotation to remove magnesium. The flotation time is 5 minutes. The foam produced by flotation is magnesium-containing tailings.
[0100] (6) Continue to add sodium carbonate to the slurry in the tank to adjust the pH to 9.0, add dodecylamine, an amine cationic collector, at a dosage of 0.3 kg / t of raw ore, and carry out reverse flotation to remove silicon. The flotation time is 5 minutes, and the foam produced by flotation is silicon-containing tailings.
[0101] (7) The primary precipitate obtained in step (4) and the flotation slurry obtained in step (6) are filtered and dried to obtain phosphate concentrate.
[0102] The phosphate tailings used in Examples 1-3 and Comparative Examples 1-2 were from the same batch. The contents of P2O5, MgO, SiO2, Fe2O3, and Al2O3 in the phosphate tailings used in Examples 1-3 and Comparative Example 1, as well as in the prepared phosphate concentrate, were determined. The P2O5 content in the phosphate rock was determined according to the quinoline phosphomolybdate gravimetric method in GB / T 1871.1-1995; the MgO content was determined according to the EDTA titration method in GB / T 1871.5-1995; the SiO2 content was determined according to the perchloric acid dehydration gravimetric method in GB / T 1873-1995; and the Fe2O3 content was determined according to the ferric phosphate (aluminum) separation-EDTA titration method in GB / T 1871.2-1995 and the Al2O3 content in the phosphate rock according to the ferric phosphate (aluminum) separation-EDTA titration method in GB / T 1871.3-1995.
[0103] Phosphorus yield = (Concentrate quality * Concentrate P2O5 grade) / (Original ore quality * Original ore P2O5 grade) * 100%
[0104] The test results are shown in Table 1.
[0105] Table 1. Test Results:
[0106]
[0107] As shown in Table 1, in Examples 1-3, by calcining the phosphorus tailings to remove the surface-adhered reagents, performing jigging to remove most of the dolomite minerals, and then performing double reverse flotation to remove some of the quartz and dolomite minerals, the residual phosphorus in the tailings can be enriched to about 27.5%, and the MgO content in the enriched concentrate can be reduced to below 1%, which can be used as a ore blending material for phosphoric acid production.
[0108] When the calcination temperature of Comparative Example 1 was less than 300℃, the fatty acid collectors could not be completely decomposed and removed, resulting in low phosphate concentrate grade and phosphate yield. This indicates that the collector residue on the surface of phosphate tailings has a significant impact on the subsequent jigging and flotation of phosphate ore. When the calcination temperature of Comparative Example 2 was greater than 400℃, the phosphate tailings decomposed, affecting the phosphate yield.
Claims
1. A method for separating phosphate concentrate from phosphate tailings, characterized in that: Includes the following steps: (1) The phosphate tailings are crushed, ground, calcined, and then mixed into a slurry for jigging to obtain primary sediment, secondary sediment, and overflow material, wherein the density of the primary sediment is 2.95-3.25 t / m³. 3 The density of the secondary precipitate is 2.85-2.95 t / m³. 3 The density of the overflow material is less than 2.85 t / m³. 3 ; (2) After preparing the secondary precipitate material obtained in step (1) into a slurry, magnesium and silicon are removed by flotation to obtain the flotated slurry; (3) The primary precipitate obtained in step (1) and the flotation slurry obtained in step (2) are filtered and dried to obtain the phosphate concentrate; in step (1), the calcination temperature is 300-400℃ and the calcination time is 3-8h; the phosphate tailings are produced from the phosphate tailings produced in the reverse flotation magnesium removal process.
2. The method for separating phosphate concentrate from phosphate tailings according to claim 1, characterized in that: In step (1), the calcination temperature is 300-350℃ and the calcination time is 4-6h.
3. The method for separating phosphate concentrate from phosphate tailings according to claim 1, characterized in that: The flotation process described in step (2) includes the following steps: S1 prepares the secondary precipitate into a slurry, adjusts the pH of the slurry to acidic, adds a fatty acid anion collector, and performs reverse flotation to remove magnesium, thus obtaining the slurry after primary flotation. S2 adjusts the pH of the pulp obtained after the first flotation to alkaline, adds an amine cationic collector, and performs reverse flotation to remove silica, thus obtaining the pulp after the second flotation.
4. A method for separating phosphate concentrate from phosphate tailings according to claim 3, characterized in that: In step S1, adjusting the pH of the slurry to acidic means adjusting the pH of the slurry to 4-6.
5. A method for separating phosphate concentrate from phosphate tailings according to claim 4, characterized in that: In step S1, adjusting the pH of the slurry to acidic means adjusting the pH of the slurry to 4.5-5.
5.
6. A method for separating phosphate concentrate from phosphate tailings according to claim 3, characterized in that: In step S1, the fatty acid anionic collector is oleic acid soap, and the ratio of the amount of fatty acid anionic collector added to the amount of phosphate tailings is (0.2-1.0) kg: 1 t.
7. A method for separating phosphate concentrate from phosphate tailings according to claim 3, characterized in that: In step S2, adjusting the pH of the pulp obtained after the first flotation to alkaline means adjusting the pH of the pulp to 7.5-10.
8. A method for separating phosphate concentrate from phosphate tailings according to claim 3, characterized in that: In step S2, the amine cationic collector is dodecylamine, and the ratio of the amount of the amine cationic collector added to the amount of phosphorus tailings is (0.1-0.5) kg: 1 t.
9. A method for separating phosphate concentrate from phosphate tailings according to claim 1, characterized in that: In step (1), the particle size of the material obtained after crushing is less than 2 mm.
10. A method for separating phosphate concentrate from phosphate tailings according to claim 1, characterized in that: In step (1), the content of -200 mesh fineness in the material obtained after grinding is 85wt%-95wt%.
11. A method for separating phosphate concentrate from phosphate tailings according to claim 1, characterized in that: In step (1), the concentration of the slurry is 20wt%-40wt%.
12. A method for separating phosphate concentrate from phosphate tailings according to claim 1, characterized in that: In step (1), the amount of water flowing through the jigging process is 1-5 m³. 3 / h, stroke is 10-30mm, strokes are 300-450 times / min.
13. A method for separating phosphate concentrate from phosphate tailings according to claim 1, characterized in that: In step (2), the concentration of the secondary precipitate after being prepared into a slurry is 20wt%-40wt%.