Method for reducing magnesium and aluminum in medium-low grade collophanite phosphorus enrichment and gravity-flotation combined concentration concentrate

Abstract

The invention relates to the technical field of phosphate ore single reverse flotation, in particular to a method for reducing magnesium and aluminum in medium-low grade collophanite phosphorus enrichment and heavy flotation combined concentration concentrate. Aiming at a phosphorus compound fertilizer enterprise for producing medium-low-grade silicon-calcium phosphate ore through ore blending, firstly, all medium-low-grade phosphate ore raw ores are uniformly mixed according to a specific proportion, and the mixed ores are ground into raw ore pulp with specific concentration and fineness through a ball mill; and adding the obtained raw ore pulp into the gravity-flotation combined concentrate, adjusting the pH value to a specific range by using phosphoric acid, and carrying out normal production after adding an anion collecting agent. By utilizing the method provided by the invention, the low grade of the concentrate produced by blending the medium-low-grade silico-calcareous phosphorite can be effectively improved, and the magnesium or aluminum of the high-magnesium concentrate can be effectively reduced by the gravity-flotation combined separation.

Description

Technical Field

[0001] This invention relates to the field of single-flotation and reverse flotation technology for phosphate rock, specifically to a method for reducing magnesium and aluminum in medium- and low-grade phosphate rock concentrate obtained through combined phosphorus enrichment and gravity flotation. Background Technology

[0002] Phosphate rock is an indispensable strategic mineral for the production of fertilizers and pesticides. Although my country ranks second in the world in phosphate rock resources, its phosphate rock is mainly low- to medium-grade siliceous calcium phosphate rock, characterized by complex crystal composition, fine particle size, and a high proportion of lean ore, making it difficult to beneficiate. Currently, the main beneficiation methods for low- to medium-grade phosphate rock in China are flotation and gravity separation. Flotation is excellent for separating fine-grained ores, but its process of removing magnesium gangue impurities using fatty acid anionic collectors is complex, and some phosphorus-containing fine particles are easily carried out with the froth, resulting in poor phosphorus enrichment in low-grade phosphate rock concentrates. Gravity separation is suitable for the pre-concentration of ores with simple mineral composition and coarse particle size, but the fine particle size of magnesium and aluminum gangue leads to high magnesium and aluminum impurity content in the gravity separation concentrate.

[0003] To fully leverage the advantages of flotation and gravity separation, the industry has developed a combined gravity-flotation separation process. For example, Chinese invention patent CN105880032A discloses a combined gravity-flotation separation method for medium- and low-grade collophane. This method involves crushing and finely grinding medium- and low-grade collophane, followed by gravity separation using a sulfuric acid slag separation and purification equipment to obtain gravity concentrate and gravity tailings. Depending on the MgO grade in the gravity concentrate, it is either used directly as concentrate or treated with carbonates via reverse flotation as concentrate I. The gravity tailings are subjected to forward flotation to obtain a rough concentrate, which is then subjected to reverse flotation to obtain concentrate II. Finally, concentrate I and concentrate II are combined to obtain the phosphate concentrate product. The process utilizes specialized sorting equipment with a fine feed particle size (50-90 wt% less than 0.074 mm), offering advantages such as large ore throughput, simple operation, low cost, and no pollution. It can reduce the amount of ore processed by direct flotation by 30-60%, lower the reagent dosage and heating costs of direct flotation, and produce phosphate concentrate with a P2O5 grade of 30-33% and an MgO grade of 0.6-1.0%, with a P2O5 recovery rate greater than 80%. Significant progress has been made in terms of sorting efficiency and cost control.

[0004] However, the existing gravity-flotation combined process still focuses on the separate gravity separation and flotation of a single ore, without addressing the synergistic flotation technology for blending low- and medium-grade phosphate rock with gravity-flotation concentrate. This fails to effectively address the issues of phosphorus enrichment and upgrading after blending low-grade phosphate rock, as well as the targeted removal of residual aluminum impurities from gravity-flotation concentrate, making it difficult to simultaneously achieve the comprehensive goals of phosphorus enrichment, magnesium reduction, and aluminum reduction. Therefore, to meet the actual needs of low- and medium-grade siliceous-calcareous phosphate rock blending production, there is an urgent need to develop an optimized process that can simultaneously improve concentrate grade and remove multiple impurities. Summary of the Invention

[0005] This invention provides a method for reducing phosphorus content in low- and medium-grade phosphate rock and reducing magnesium and aluminum content in the combined gravity flotation concentrate. This method allows low- and medium-grade phosphate rock to be blended and then separated with the gravity flotation concentrate to achieve the goal of enriching phosphorus, reducing magnesium, and reducing aluminum content in the final concentrate.

[0006] To achieve the above objectives, the technical solution adopted by the present invention is as follows: This invention provides a method for reducing magnesium and aluminum in medium- and low-grade phosphate rock concentrate obtained through combined phosphorus and gravity flotation processes, comprising the following steps: S1. Ore blending and grinding: Mix at least two types of medium and low grade phosphate rock of different grades in a predetermined ratio, and grind them together with water to obtain raw ore slurry. S2. Grading: The slurry obtained in step S1 is graded to separate the overflow slurry that meets the predetermined concentration and fineness requirements. S3, Mixed concentrate: The gravity flotation concentrate is mixed with the overflow slurry obtained in step S2 to obtain a mixed slurry; S4. pH adjustment: Add an acidic regulator to the mixed slurry obtained in step S3 to adjust the pH value of the slurry to the acidic range; S5. Slurry conditioning and reagent addition: Add flotation return water to the pH-adjusted slurry to adjust the slurry concentration to the predetermined flotation concentration, then add anionic collector and stir to condition the slurry; S6. Flotation Separation: The slurry after conditioning is subjected to flotation operations that include at least roughing, cleaning and scavenging stages, wherein a conditioning agent is added in the cleaning and / or scavenging stages to finally obtain phosphate concentrate.

[0007] The aforementioned scheme defines the core technological framework for enriching phosphorus in low- and medium-grade phosphate rock and reducing magnesium and aluminum in the gravity flotation concentrate. Through a complete process including ore blending and grinding, classification, mixing concentrates, pH adjustment, slurry preparation and reagent addition, and flotation separation, it innovatively combines ore blending technology with synergistic flotation of the gravity flotation concentrate. This overcomes the limitations of single flotation or gravity separation, achieving synergistic separation of low- and medium-grade phosphate rock blends and gravity flotation concentrates, simultaneously achieving the three objectives of enriching phosphorus, reducing magnesium, and reducing aluminum. It effectively improves the concentrate grade of low- and medium-grade phosphate rock blends while removing magnesium and aluminum impurities from the gravity flotation concentrate, ultimately achieving a P2O5 grade of ≥30% and a recovery rate of ≥90% in the phosphate concentrate.

[0008] As a preferred embodiment of the present invention, in step S1, the ore blending and grinding has at least one of the following features a1)-a3): a1) The P2O5 grade in the mixed phosphate rock is 22-25%, and the MgO content is 3-6%; a2) The average grade of the medium and low grade phosphate rock is expressed as the mass fraction of phosphorus pentoxide, including three different grades of phosphate rock: 22% phosphate rock, 24% phosphate rock, and 25% phosphate rock, with a blending ratio of 1:1:1 for the three different grades of phosphate rock. a3) The water mentioned is the return water from the thickener production process, with a pH value of 3.5-4.8.

[0009] The above scheme clearly defines the key parameters for the ore blending and grinding stages. The blending ratio (1:1:1) and grade range (P2O5 22-25%, MgO 3-6%) are designed reasonably to ensure the stability of raw material properties. Thickening equipment is used to recycle wastewater, achieving resource utilization and combining environmental protection with economic efficiency. Multiple characteristic parameters can be flexibly selected to adapt to different raw material conditions. Controlling the phosphorus grade and magnesium impurity content of the initial raw material avoids the impact of raw material fluctuations on flotation efficiency, while reducing fresh water consumption and lowering production costs, laying the foundation for achieving subsequent phosphorus enrichment and magnesium reduction targets.

[0010] As a preferred technical solution of the present invention, in step S2, the overflow slurry that meets the predetermined concentration and fineness requirements has a predetermined concentration of 30-42 wt% and a fineness of -0.074 mm (fineness of -0.074 mm refers to a particle size of less than 0.074 mm) accounting for 72-85 wt%.

[0011] In the above scheme, the concentration and fineness parameters are matched to the requirements of subsequent flotation operations, avoiding insufficient dissociation of coarse particles or excessive loss of fine particles, and ensuring that the pulp has good separation conditions. This improves the contact efficiency between minerals and reagents in subsequent flotation processes, reduces ineffective separation, and provides a guarantee for improving concentrate grade and recovery rate, avoiding a decrease in separation efficiency due to improper pulp particle size or concentration.

[0012] As a preferred embodiment of the present invention, in step S3, the mixed concentrate has at least one of the following characteristics b1)-b4): b1) The mass concentration of the gravity flotation concentrate is 55-65 wt%; b2) The gravity flotation concentrate and the overflow slurry are mixed at a volume ratio of 8:1 to 1:1; b3) Before mixing, adjust the concentration of the heavy flotation concentrate to match that of the overflow slurry; b4) The mixing time is 5-15 min.

[0013] In the above scheme, the mixing ratio (volume ratio 8:1-4:1) and time (5-15 min) are scientifically designed to ensure full integration of the two slurries; the concentration consistency is adjusted before mixing to avoid uneven separation due to concentration differences; the concentration of the gravity flotation concentrate (55-65wt%) is adapted to the mixing requirements, balancing separation efficiency and raw material utilization. This achieves the synergistic effect of low-grade phosphate rock blending and gravity flotation concentrate, improving the separation performance of the mixed slurry, ensuring the stability of phosphorus enrichment, magnesium reduction, and aluminum reduction effects, and avoiding index fluctuations caused by improper mixing.

[0014] As a preferred embodiment of the present invention, in step S4, the pH adjustment has at least one of the following features c1)-c3): c1) The dosage of the acidic regulator is 5-9 kg / t of feed; c2) The acid regulator is phosphoric acid; c3) The pH value of the adjusted slurry is 4.0-4.8.

[0015] In the above scheme, phosphoric acid is selected as the acidity regulator, which does not introduce new impurity ions and can supplement some phosphorus elements; the dosage (5-9 kg / t feed) and pH value (4.0-4.8) are precisely controlled to match the action conditions of the subsequent anionic collector; an acidic flotation environment conducive to the separation of phosphorus minerals from magnesium and aluminum impurities is created, enhancing the selective collection ability of the subsequent collector on phosphorus minerals, and providing key process conditions for magnesium and aluminum reduction.

[0016] As a preferred embodiment of the present invention, in step S5, the slurry preparation and chemical addition have at least one of the following characteristics d1)-d4): d1) The predetermined flotation concentration is 30-35 wt%; d2) The dosage of the anionic collector is 1.1-1.6 kg / t of feed. d3) The anionic collector is a fatty acid saponified product.

[0017] In the above scheme, the flotation concentration (30-35 wt%) and collector dosage (1.1-1.6 kg / t feed) are well matched, avoiding reagent waste or insufficient dosage; fatty acid saponification is selected as an anionic collector, which has good selective collection performance for phosphate minerals; the parameter design takes into account both separation efficiency and cost control. This improves the collection efficiency of phosphate minerals, reduces the adsorption of magnesium and aluminum impurities, and controls reagent consumption, thereby reducing production costs and ensuring the achievement of phosphorus enrichment and impurity removal targets in the flotation separation stage.

[0018] As a preferred embodiment of the present invention, the fatty acid saponified product is prepared by a constant-temperature saponification reaction of mixed fatty acids and caustic soda solution at a mass fraction ratio of 1:0.8-2.5 at 65-75 °C. The mixed fatty acids are composed of palmitic acid, linolenic acid, arachidic acid and oleic acid mixed at a mass fraction ratio of 22-28:8-12:5-9:51-57.

[0019] In the above scheme, the synthesized fatty acid saponified compounds are highly targeted and have superior collecting performance compared to conventional collectors. This enhances the selectivity of the collector for phosphorus minerals, further improving concentrate grade, while reducing the entrainment of magnesium and aluminum impurities. It solves the problem of poor separation effect of conventional collectors, providing core reagent support for achieving the final target.

[0020] As a preferred technical solution of the present invention, in step S6, the flotation separation operation includes a roughing operation, a cleaning operation, at least three scavenging operations, and at least two middlings re-dispersion operations performed sequentially; wherein, an anionic collector and phosphoric acid are added as modifiers in the cleaning operation, and phosphoric acid is added as modifiers in the third scavenging operation.

[0021] The above scheme incorporates multi-stage scavenging and middlings re-selection to ensure thorough separation and reduce phosphate mineral loss. Precise addition of modifiers (anionic collectors, phosphoric acid) specifically enhances the effects of fine separation, impurity removal, and scavenging recovery. The process design balances grade improvement with recovery assurance. Further removal of magnesium and aluminum impurities increases the purity of the phosphate concentrate, while simultaneously recovering phosphate minerals from scavenging and middlings, reducing tailings loss and ensuring both concentrate grade and recovery rate meet standards.

[0022] As a preferred technical solution of the present invention, the middlings and tailings of each flotation operation section are recycled according to the following principle: the tailings from the first stage of fine selection are returned to the first stage of roughing feed; the middlings from the first stage of scavenging are returned to the first stage of roughing feed; the middlings from the second stage of scavenging are returned to the first stage of scavenging feed; the middlings from the third stage of scavenging are returned to the classification equipment in step S2; the middlings from the fourth stage of middling re-selection are returned to the third stage of scavenging feed; and the middlings from the fifth stage of middling re-selection are returned to the fourth stage of middling re-selection feed.

[0023] The above scheme features a well-designed circulation path, achieving a closed-loop recovery process: tailings concentrate → roughing and scavenging stage 1 middlings → roughing and scavenging stage 2 middlings → scavenging stage 1 and scavenging stage 3 middlings → classification and middlings re-selection of middlings → corresponding preceding operations. This results in high resource utilization. It maximizes the recovery of phosphorus minerals from the slurry, reduces the P2O5 content in the tailings, improves the overall recovery rate, reduces resource waste, and stabilizes the operating conditions of the flotation process, preventing middlings accumulation from affecting the separation effect.

[0024] As a preferred technical solution of the present invention, the process parameters of each flotation stage are controlled as follows: the roughing stage residence time is 8-10 min, and the aeration pressure is 270 kPa; the cleaning stage residence time is 5-8 min, and the aeration pressure is 250-270 kPa; the first and second scavenging stages residence times are 4-6 min each, and the aeration pressure is 250-260 kPa; the third scavenging stage residence time is 4-6 min, and the aeration pressure is 240-260 kPa; the middlings reprocessing operation adopts bidirectional circulating flotation, with a single stage residence time of 2-5 min and an aerator inlet pressure of 250 kPa.

[0025] In the above scheme, the parameters for each stage (such as roughing residence time of 8-10 min and aeration pressure of 270 kPa; middlings re-selection residence time of 2-5 min) are matched to the separation requirements of the corresponding operation; the parameter range is controllable, facilitating operational adjustments in industrial production. This ensures the high efficiency and stability of the flotation process, avoids the decline in separation effect due to parameter fluctuations, and ensures the stability and reliability of the final concentrate indicators (P2O5 grade ≥30%, recovery rate ≥90%), providing technical support for industrial applications.

[0026] Compared with the prior art, the beneficial effects of the present invention are as follows: This invention ingeniously incorporates a specific ratio of gravity flotation concentrate into low- to medium-grade phosphate ore blends for roughing, cleaning, and scavenging, followed by a closed-circuit single-phase reverse flotation process. This enhances the phosphorus enrichment effect of single-phase reverse flotation with a single low- to medium-grade phosphate ore blend. Simultaneously, it removes the high levels of magnesium and aluminum gangue in the gravity flotation concentrate caused by the processing of fine particles. The overall separation recovery rate is high, and the tailings loss rate is low. This method has practical application value. Using this method, the P2O5 grade of the flotation concentrate is ≥30%, and the phosphate concentrate recovery rate is ≥90%. Attached Figure Description

[0027] Figure 1 This is a flowchart of a method for reducing magnesium and aluminum in medium- and low-grade collophane concentrate obtained by combined phosphorus enrichment and gravity flotation. Detailed Implementation

[0028] To better understand the present invention, the following embodiments further illustrate its content, but the scope of the present invention is not limited to the examples below. Any simple modifications, equivalent structural transformations, or alterations made based on the essential content of the present invention should be covered within the protection scope of the present invention.

[0029] The technical solution of the present invention will be further described and illustrated below through examples. All raw materials used in the examples are commercially available or prepared using conventional methods.

[0030] Example 1 The phosphate ore used in this example comes from a certain company, and the heavy flotation concentrate comes from the Houping phosphate mine beneficiation plant of Xingfa Group.

[0031] The ratio of 22% phosphate rock: 24% phosphate rock: 25% phosphate rock is 1:1:1, with the proportion of low-grade 22% phosphate rock exceeding 33%. Its impurities are mainly fine-grained magnesian gangue (calculated as MgO). The grinding fineness is -0.074 mm, and the percentage of mineral particles is 75 wt%. The indicators of the phosphate ore and the indicators of the gravity flotation concentrate with this ore blend ratio are shown in Table 1 below.

[0032] Table 1

[0033] like Figure 1 As shown in Table 1 above, the P2O5 grade of the raw ore is as follows. The raw ore is fed into a ball mill for grinding, and the grinding fineness is controlled to be -0.074 mm with a mass fraction of 72-85%. After grinding, the slurry enters a hydrocyclone for classification. The qualified slurry after classification is fed into the No. 1 mixing tank. 7 kg / t of phosphoric acid (38wt%) is added to the No. 1 mixing tank. The slurry after uniform mixing is added together with the phosphate concentrate from gravity flotation in a ratio of 4:1 to the No. 2 mixing tank. Flotation return water is added to the No. 2 mixing tank to adjust the slurry concentration to a flotation concentration of 30-35wt%. Then, 1.2 kg / t of anionic collector is added and thoroughly mixed.

[0034] The uniformly mixed slurry is fed into the roughing stage for flotation, with a residence time of 8-10 minutes and an aeration pressure of 270 kPa. The underflow slurry after roughing is then fed to the cleaning stage, where the residence time is 5-8 minutes and the aeration pressure is 250-270 kPa. 1-2 kg / t of anionic collector is added to the cleaning stage, and the cleaning underflow becomes the final concentrate, which is then thickened in the concentrate thickener. The froth from the roughing stage is fed to the scavenging stage, where the residence time is 4-6 minutes and the aeration pressure is 250-260 kPa. The froth from the scavenging stage is then fed into the scavenging stage II flotation cell. The underflow from the scavenging stage I is returned to the roughing stage, and the underflow from the scavenging stage II is returned to the scavenging stage, each undergoing bidirectional circulation flotation (scavenging stage II residence time 4-6 minutes, aeration pressure 250-260 kPa). The froth from the scavenging stage II is then fed to the scavenging stage III, where 2.0 kg / t of phosphoric acid is added. The residence time in the scavenging stage III is 4-6 minutes. The aeration pressure is 240-260 kPa. The three-stage scavenging froth return hydrocyclone is used for further classification. This step further separates the qualified phosphate rock particles and sends them into the flotation process, reducing phosphorus loss in the tailings. Simultaneously, two stages of middlings re-selection are added: the third stage middlings (concentration: 18-25%, fineness 70-85%, P2O5 20-22%, Mg 4-6%), and the fourth and fifth stages of middlings re-selection are bidirectional circulating flotation (each stage residence time is 5 min, aeration pressure at the aerator inlet is 250 kPa). The final tailings from the fifth stage are discharged to the tailings thickener, with a tailings P2O5 grade of 4.51%. After flotation using this method, a final concentrate with a P2O5 grade of 31.82% and an MgO grade of 0.92% is obtained, with a P2O5 recovery rate of 95.71%, as shown in Table 2 below.

[0035] Table 2

[0036] Comparative Example 1 For 22% phosphate rock: 24% phosphate rock: 25% phosphate rock = 1:1:1, flotation was performed separately, without the addition of gravity flotation concentrate. The flotation method was as in Example 1. The P2O5 grade of the raw ore was as shown in the table above. The raw ore was fed into a ball mill for grinding. The grinding fineness was controlled to -0.074 mm, and the mass fraction was 72-85%. After grinding, the slurry entered a hydrocyclone for classification. The qualified slurry after classification was fed into the No. 1 mixing tank. 7 kg / t of phosphoric acid (38wt%) was added to the No. 1 mixing tank. After stirring evenly, the slurry was added to the No. 2 mixing tank. Flotation return water was added to the No. 2 mixing tank to adjust the slurry concentration to a flotation concentration of 30-35wt%. Then, 1.2 kg / t of anionic collector was added and stirred evenly.

[0037] The uniformly mixed slurry is fed into the roughing stage for flotation, with a residence time of 8-10 minutes and an aeration pressure of 270 kPa. The underflow slurry after roughing is then fed to the cleaning stage, where the residence time is 5-8 minutes and the aeration pressure is 250-270 kPa. 1-2 kg / t of anionic collector is added to the cleaning stage, and the cleaning underflow becomes the final concentrate, which is then thickened in the concentrate thickener. The froth from the roughing stage is fed to the scavenging stage, where the residence time is 4-6 minutes and the aeration pressure is 250-260 kPa. The froth from the scavenging stage is then fed into the scavenging stage II flotation cell. The underflow from the scavenging stage I is returned to the roughing stage, and the underflow from the scavenging stage II is returned to the scavenging stage, each undergoing bidirectional circulation flotation (scavenging stage II residence time 4-6 minutes, aeration pressure 250-260 kPa). The froth from the scavenging stage II is then fed to the scavenging stage III, where 2.0 kg / t of phosphoric acid is added. The residence time in the scavenging stage III is 4-6 minutes. The aeration pressure is 240-260 kPa. The three-stage scavenging stage foam is returned to the hydrocyclone for further classification. This step further separates the qualified fine phosphate rock particles and sends them into the flotation process, reducing phosphorus loss in the tailings. At the same time, two stages of middlings re-selection are added. The middlings of the three-stage scavenging stage (concentration: 18%-25%, fineness 70%-85%, P2O5 20-22%, Mg 4%-6%), the middlings re-selection stage four and five are bidirectional circulating flotation (single stage residence time is 5 min, aeration pressure at the aerator inlet is 250 kPa). The final tailings are discharged from the middlings re-selection stage five to the tailings thickener. The final results are shown in Table 3 below.

[0038] Table 3

[0039] Comparing Tables 2 and 3, it can be found that the difference between the two flotation feeds is only in the phosphorus concentrate index. Specifically, the phosphorus concentrate of the blended feed with a ratio of 4:1 (greater flotation concentrate) is 1.14% higher than that of the blended feed with low-grade collophane, achieving a phosphorus-enriched effect compared to the blended feed with low-grade collophane alone.

[0040] Comparative Example 2 For the heavy flotation concentrate, separate flotation is performed using the method described in Example 1. The slurry is fed into a No. 1 mixing tank, and 7 kg / t of phosphoric acid (38 wt%) is added to the No. 1 mixing tank. After stirring evenly, the slurry is added to a No. 2 mixing tank, and flotation return water is added to a No. 22 mixing tank. The slurry concentration is adjusted to a flotation concentration of 30-35 wt%. Then, 1.2 kg / t of anionic collector is added and stirred thoroughly until evenly mixed.

[0041] The uniformly mixed slurry is fed into the roughing stage for flotation, with a residence time of 8-10 minutes and an aeration pressure of 270 kPa. The underflow slurry after roughing is then fed to the cleaning stage, where the residence time is 5-8 minutes and the aeration pressure is 250-270 kPa. 1-2 kg / t of anionic collector is added to the cleaning stage, and the cleaning underflow becomes the final concentrate, which is then thickened in the concentrate thickener. The froth from the roughing stage is fed to the scavenging stage, where the residence time is 4-6 minutes and the aeration pressure is 250-260 kPa. The froth from the scavenging stage is then fed into the scavenging stage II flotation cell. The underflow from the scavenging stage I is returned to the roughing stage, and the underflow from the scavenging stage II is returned to the scavenging stage, each undergoing bidirectional circulation flotation (scavenging stage II residence time 4-6 minutes, aeration pressure 250-260 kPa). The froth from the scavenging stage II is then fed to the scavenging stage III, where 2.0 kg / t of phosphoric acid is added. The residence time in the scavenging stage III is 4-6 minutes. The aeration pressure is 240-260 kPa. The three-stage scavenging stage foam is returned to the hydrocyclone for further classification. This step further separates the qualified fine phosphate rock particles and sends them into the flotation process, reducing phosphorus loss in the tailings. At the same time, two stages of middlings re-selection are added. The middlings of the three-stage scavenging stage (concentration: 18-25%, fineness: 70-85%, P2O5: 20-22%, Mg: 4-6%), the four-stage middlings re-selection stage and the five-stage middlings re-selection stage are bidirectional circulating flotation (single stage residence time is 5 min, aeration pressure at the aerator inlet is 250 kPa). The final tailings are discharged from the five-stage middlings re-selection stage to the tailings thickener. The final results are shown in Table 4 below.

[0042] Table 4

[0043] Comparing Tables 2 and 4, it can be found that although the magnesium reduction effect of the gravity flotation concentrate is obvious when it is used alone in this process, with high yield and recovery rate and small phosphorus loss, the aluminum content of the concentrate increases after re-selection, and the cost of re-flotation of the gravity flotation concentrate alone is too high, which does not meet the requirements of normal production.

[0044] Example 2 Based on production costs and performance indicators, the proportion of concentrate from gravity flotation is reduced, with the ratio of low-grade phosphate rock to gravity flotation being 8:1. The proportion of low-grade phosphate rock in the blend is not specified: 22% phosphate rock: 24% phosphate rock: 25% phosphate rock = 1:1:1.

[0045] Table 5

[0046] like Figure 1As shown in the table above, the P2O5 grade of the raw ore is as follows. The raw ore is fed into a ball mill for grinding, and the grinding fineness is controlled to be -0.074mm with a mass fraction of 72-85%. After grinding, the slurry enters a hydrocyclone for classification. The qualified slurry after classification is fed into the No. 1 mixing tank. 7 kg / t of phosphoric acid (38wt%) is added to the No. 1 mixing tank. The slurry after uniform mixing is added together with the phosphate concentrate from gravity flotation in a ratio of 8:1 to the No. 2 mixing tank. Flotation return water is added to the No. 2 mixing tank to adjust the slurry concentration to a flotation concentration of 30-35wt%. Then, 1.2 kg / t of anionic collector is added and thoroughly mixed.

[0047] The uniformly mixed slurry is fed into the roughing stage for flotation, with a residence time of 8-10 minutes and an aeration pressure of 270 kPa. The underflow slurry after roughing is then fed to the cleaning stage, where the residence time is 5-8 minutes and the aeration pressure is 250-270 kPa. 1-2 kg / t of anionic collector is added to the cleaning stage, and the cleaning underflow becomes the final concentrate, which is then thickened in the concentrate thickener. The froth from the roughing stage is fed to the scavenging stage, where the residence time is 4-6 minutes and the aeration pressure is 250-260 kPa. The froth from the scavenging stage enters the scavenging stage II flotation cell. The underflow from the scavenging stage I returns to the roughing stage, and the underflow from the scavenging stage II returns to the scavenging stage, each undergoing bidirectional circulation flotation (scavenging stage II residence time 4-6 minutes, aeration pressure 250-260 kPa). The froth from the scavenging stage II is then fed to the scavenging stage III, where 2.0 kg / t of phosphoric acid is added. The residence time in the scavenging stage III is 4-6 minutes. The aeration pressure is 240-260 kPa. The three-stage scavenging froth return hydrocyclone further classifies the phosphate rock particles, reducing phosphorus loss in the tailings. This step also adds two stages of middlings re-selection: scavenging stage three middlings (concentration 18%-25%, fineness 70%-85%, P2O5 20%-22%, Mg 4-6%), followed by two-way circulating flotation in stages four and five (each stage residence time is 5 min, aeration pressure at the aerator inlet is 250 kPa). The final tailings from stage five are discharged to the tailings thickener, with a tailings P2O5 grade of 4.82%. After flotation using this method, a final concentrate with a P2O5 grade of 31.58% and an MgO grade of 0.96% is obtained, achieving a P2O5 recovery rate of 94.84%.

[0048] Table 6

[0049] Comparing Table 6 and Table 2, after the proportion of concentrate in the combined gravity flotation and separation decreased, the phosphorus enrichment effect of the concentrate deteriorated, the phosphorus recovery rate decreased, and the phosphorus loss increased, resulting in the underutilization of resources.

[0050] Example 3 The phosphate ore used in this example comes from a certain company, and the heavy flotation concentrate comes from the Houping phosphate mine beneficiation plant of Xingfa Group.

[0051] The ratio of 22% phosphate rock: 24% phosphate rock: 25% phosphate rock is 1:1:1, with the proportion of low-grade 22% phosphate rock exceeding 33%. Its impurities are mainly fine-grained magnesian gangue (calculated as MgO). The grinding fineness is -0.074 mm, and the percentage of mineral particles is 75 wt%. The indicators of the phosphate ore and the indicators of the gravity flotation concentrate with this ore blend ratio are shown in Table 1 below.

[0052] Table 7

[0053] like Figure 1 As shown in Table 7 above, the P2O5 grade of the raw ore is as follows. The raw ore is fed into a ball mill for grinding, and the grinding fineness is controlled to be -0.074 mm with a mass fraction of 72-85%. After grinding, the slurry enters a hydrocyclone for classification. The qualified slurry after classification is fed into the No. 1 mixing tank. 7 kg / t of phosphoric acid (38wt%) is added to the No. 1 mixing tank. The slurry after uniform mixing is added together with the phosphate concentrate from gravity flotation in a 1:1 ratio to the No. 2 mixing tank. Flotation return water is added to the No. 2 mixing tank to adjust the slurry concentration to a flotation concentration of 30-35wt%. Then, 1.2 kg / t of anionic collector is added and thoroughly mixed.

[0054] The uniformly mixed slurry is fed into the roughing stage for flotation, with a residence time of 8-10 minutes and an aeration pressure of 270 kPa. The underflow slurry after roughing is then fed to the cleaning stage, where the residence time is 5-8 minutes and the aeration pressure is 250-270 kPa. 1-2 kg / t of anionic collector is added to the cleaning stage, and the cleaning underflow becomes the final concentrate, which is then thickened in the concentrate thickener. The froth from the roughing stage is fed to the scavenging stage, where the residence time is 4-6 minutes and the aeration pressure is 250-260 kPa. The froth from the scavenging stage is then fed into the scavenging stage II flotation cell. The underflow from the scavenging stage I is returned to the roughing stage, and the underflow from the scavenging stage II is returned to the scavenging stage, each undergoing bidirectional circulation flotation (scavenging stage II residence time 4-6 minutes, aeration pressure 250-260 kPa). The froth from the scavenging stage II is then fed to the scavenging stage III, where 2.0 kg / t of phosphoric acid is added. The residence time in the scavenging stage III is 4-6 minutes. The aeration pressure is 240-260 kPa. The three-stage scavenging stage returns the froth to the hydrocyclone for further classification. This step further separates the qualified phosphate rock particles and sends them into the flotation process, reducing phosphorus loss in the tailings. Simultaneously, it adds two stages of middlings re-selection: the third stage middlings (concentration: 18-25%, fineness 70-85%, P2O5 20-22%, Mg 4-6%), and the fourth and fifth stages of bidirectional circulating flotation (each stage residence time is 5 min, aeration pressure at the aerator inlet is 250 kPa). The final tailings from the fifth stage are discharged to the tailings thickener, with a tailings P2O5 grade of 4.82%. After flotation using this method, a final concentrate with a P2O5 grade of 32.10% and an MgO grade of 0.74% is obtained, with a P2O5 recovery rate of 97.10%, as shown in Table 8 below.

[0055] Table 8

[0056] Under this ratio, although the yield and phosphorus enrichment effect are high, the high proportion of heavy flotation concentrate leads to increased foaming on site, making actual control difficult and causing serious overflow in the flotation cell. In practical applications, the proportion of heavy flotation concentrate should not exceed 50%.

[0057] It should be understood that the above embodiments are for illustrative purposes only and are not intended to limit the scope of protection of the present invention. Furthermore, it should be understood that after reading the teachings of this invention, those skilled in the art can make various alterations or modifications to the invention, and these equivalent forms also fall within the scope defined by the appended claims.

Claims

1. A method for reducing magnesium and aluminum content in medium- and low-grade collophane concentrate obtained through combined phosphorus and gravity flotation processes, characterized in that... Includes the following steps: S1. Ore blending and grinding: Mix at least two types of medium and low grade phosphate rock of different grades in a predetermined ratio, and grind them together with water to obtain raw ore slurry. S2. Grading: The slurry obtained in step S1 is graded to separate the overflow slurry that meets the predetermined concentration and fineness requirements. S3, Mixed concentrate: The gravity flotation concentrate is mixed with the overflow slurry obtained in step S2 to obtain a mixed slurry; S4. pH adjustment: Add an acidic regulator to the mixed slurry obtained in step S3 to adjust the pH value of the slurry to the acidic range; S5. Slurry conditioning and reagent addition: Add flotation return water to the pH-adjusted slurry to adjust the slurry concentration to the predetermined flotation concentration, then add anionic collector and stir to condition the slurry; S6. Flotation Separation: The slurry after conditioning is subjected to flotation operations that include at least roughing, cleaning and scavenging stages, wherein a conditioning agent is added in the cleaning and / or scavenging stages to finally obtain phosphate concentrate.

2. The method according to claim 1, characterized in that, In step S1, the ore blending and grinding have at least one of the following characteristics a1)-a3): a1) The P2O5 grade in the mixed phosphate rock is 22-25%, and the MgO content is 3-6%; a2) The average grade of the medium and low grade phosphate rock is expressed as the mass fraction of phosphorus pentoxide, including three different grades of phosphate rock: 22% phosphate rock, 24% phosphate rock, and 25% phosphate rock, with a blending ratio of 1:1:1 for the three different grades of phosphate rock. a3) The water mentioned is the return water from the thickener production process, with a pH value of 3.5-4.

8.

3. The method according to claim 1, characterized in that, In step S2, the overflow slurry that meets the predetermined concentration and fineness requirements has a predetermined concentration of 30-42 wt% and a fineness of -0.074 mm of 72-85 wt%.

4. The method according to claim 1, characterized in that, In step S3, the mixed concentrate has at least one of the following characteristics b1)-b4): b1) The mass concentration of the gravity flotation concentrate is 55-65 wt%; b2) The gravity flotation concentrate and the overflow slurry are mixed at a volume ratio of 8:1 to 1:1; b3) Before mixing, adjust the concentration of the heavy flotation concentrate to match that of the overflow slurry; b4) The mixing time is 5-15 min.

5. The method according to claim 1, characterized in that, In step S4, the pH adjustment has at least one of the following features c1)-c3): c1) The dosage of the acidic regulator is 5-9 kg / t of feed; c2) The acid regulator is phosphoric acid; c3) The pH value of the adjusted slurry is 4.0-4.

8.

6. The method according to claim 1, characterized in that, In step S5, the slurry preparation and chemical addition have at least one of the following characteristics d1)-d4): d1) The predetermined flotation concentration is 30-35 wt%; d2) The dosage of the anionic collector is 1.1-1.6 kg / t of feed. d3) The anionic collector is a fatty acid saponified product.

7. The method according to claim 6, characterized in that, The fatty acid saponified product is prepared by constant-temperature saponification reaction of mixed fatty acids and caustic soda solution at a mass fraction ratio of 1:0.8-2.5 at 65-75 °C. The mixed fatty acids are composed of palmitic acid, linolenic acid, arachidic acid and oleic acid mixed in a mass fraction ratio of 22-28:8-12:5-9:51-57.

8. The method according to claim 1, characterized in that, In step S6, the flotation separation operation includes a first roughing stage, a first cleaning stage, at least three scavenging stages, and at least two middlings re-selection stages performed sequentially; wherein, anionic collector and phosphoric acid are added as modifiers in the first cleaning stage, and phosphoric acid is added as modifiers in the third scavenging stage.

9. The method according to claim 8, characterized in that, The middlings and tailings from each flotation stage are recycled according to the following principles: tailings from the first stage of fine separation are returned to the first stage of roughing feed; middlings from the first stage of scavenging are returned to the first stage of roughing feed; middlings from the second stage of scavenging are returned to the first stage of scavenging feed; middlings from the third stage of scavenging are returned to the classifying equipment in step S2; middlings from the fourth stage of middling re-selection are returned to the third stage of scavenging feed; and middlings from the fifth stage of middling re-selection are returned to the fourth stage of middling re-selection feed.

10. The method according to claim 9, characterized in that, The process parameters for each flotation stage are controlled as follows: the roughing stage has a residence time of 8-10 min and an aeration pressure of 270 kPa; the cleaning stage has a residence time of 5-8 min and an aeration pressure of 250-270 kPa; the first and second scavenging stages each have a residence time of 4-6 min and an aeration pressure of 250-260 kPa; the third scavenging stage has a residence time of 4-6 min and an aeration pressure of 240-260 kPa; the middlings reprocessing stage uses bidirectional circulating flotation with a single-stage residence time of 2-5 min and an aerator inlet pressure of 250 kPa.

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