Utilization method of collophanite gravity-flotation combined beneficiation backwater

Abstract

The invention discloses a collophanite gravity-flotation combined beneficiation backwater utilization method, which comprises the following steps: carrying out ore grinding and grading on collophanite, introducing into a reverse flotation magnesium removal process to obtain magnesium removal concentrate a and magnesium removal tailings b, carrying out gravity concentration on the magnesium removal concentrate a to obtain gravity concentration concentrate c and gravity concentration tailings d, and carrying out gravity concentration reverse flotation dealumination process on the gravity concentration concentrate c to obtain phosphate concentrate e and tailings f; and the direct utilization of the backwater refers to collecting the magnesium-removed tailings b, the phosphate concentrate e, the tailings f and the phosphate concentrate g, respectively thickening and filtering to obtain the backwater, feeding the backwater into a No.1 backwater pool, and directly returning the backwater to the processes of ore grinding, reverse flotation magnesium removal, reselection, reconcentration and reverse flotation dealumination for reuse. According to the method for utilizing the collophanite gravity-flotation combined mineral separation backwater, a process capable of reducing the treatment and utilization amount in the utilization of the separation backwater of the collophanite and impurities magnesium and aluminum contained in the collophanite is found, so that the backwater recycling in the whole process is realized, and the separation index of removing the impurities magnesium and aluminum from the collophanite is not influenced.

Description

Technical Field

[0001] The invention belongs to the technical field of phosphate ore dressing, and in particular relates to a method for utilizing return water of collophane ore refloatation combined ore dressing. Background Art

[0002] In the industrial production of collophanite, how to effectively realize the utilization of return water is closely related to the amount of new water used in the dressing plant and the quality of the mineral processing technical indicators, as well as the storage capacity pressure and environmental protection issues of the tailings pond. The separation process of collophanite is generally direct and reverse flotation and double reverse flotation. The wastewater of the reverse flotation concentrate in the direct and reverse flotation process has no adverse effect on the reverse flotation process, on the contrary, it can also reduce the use of mixed acid. However, if the wastewater of the reverse flotation concentrate and the wastewater of the tailings pond are directly returned to the direct flotation process, it will cause instability of the direct flotation process, deterioration of the mineral processing indicators, and increase in the consumption of mineral processing reagents. In the double reverse flotation process, the return water from the de-magnesium section and the dealuminization section is inconsistent in the residual reagents, pH value, COD content, suspended particles, etc. in the water. If it is not separately treated and reused, it will directly lead to problems such as sticky flotation foam, poor stability of the flotation process, and reduced flotation concentrate indicators. The return water from double reverse flotation needs to be fully treated before it can be reused. There are many types of return water treatment reagents, and the return water reuse plan is complicated and lengthy.

[0003] CN111167610A discloses a reverse flotation method for calcium-silicon. The technical scheme is that after the raw ore is ground, a flotation process is adopted in which reverse flotation is first used for magnesium removal and then for silicon removal by forward flotation. The reverse flotation return water enters the reverse flotation return water tank and then returns to the reverse flotation for magnesium removal. The forward flotation return water enters the forward flotation return water tank and then returns to the forward flotation for silicon removal. The surplus forward flotation return water is fed into the reverse flotation return water tank for reverse flotation. The invention realizes the full reuse of wastewater to a certain extent, but the sorting process is too long, the types of return water treatment agents used are many, and the return water utilization scheme is complicated.

[0004] CN112237993B discloses a method for recycling wastewater from the double reverse flotation process of collophanite. The technical scheme is as follows: after removing silicate minerals from the collophanite, tailings and coarse concentrate are obtained. The tailings enter a tailings thickener for thickening and filtration, and a reagent is added to the tailings thickener to treat the return water before reuse; the coarse concentrate enters an intermediate thickener for thickening and then enters thickening and filtration, and a reagent is added to the intermediate thickener to treat the return water before reuse; the filter cake of the coarse concentrate is subjected to reverse flotation again after pulping to remove carbonate minerals, so as to obtain tailings and concentrate slurry, and the tailings are thickened and filtered by a tailings thickener; the concentrate slurry is thickened and filtered by a concentrate thickener, and a reagent is added to the concentrate thickener to treat the return water before reuse. This method has created a scientific and economical industrial implementation plan for the wastewater quality recycling of double reverse flotation of collophanite. However, the three types of return water in this plan have inconsistent properties. Not only do they need to be treated separately for recycling, but they also need to be treated completely before they can be reused. There are many types of return water treatment agents and the return water reuse plan is complicated. Summary of the invention

[0005] In view of the problems existing in the prior art, the purpose of the present invention is to provide a method for utilizing the return water of the refloatation combined beneficiation of collophanite. The method for utilizing the return water of the refloatation combined beneficiation of collophanite described in the present invention is to find a process that can reduce the amount of return water that needs to be processed in the utilization of the return water for separating the collophanite from the impurities magnesium and aluminum contained therein, so as to realize the full-process return water reuse without affecting the separation index of the collophanite for removing the impurities magnesium and aluminum, greatly simplify the return water treatment scheme, reduce the return water treatment utilization cost, and the return water treatment agent is montmorillonite, which has a wide source and is cheap.

[0006] To achieve the above object, the present invention provides a method for utilizing the return water of the refloating combined beneficiation of collophanite, the method comprising the steps of refloating combined beneficiation of collophanite and utilizing the return water, specifically comprising: A. Refloatation combined beneficiation of collophanite: after grinding and classification, the collophanite is introduced into the reverse flotation de-magnesiumization process, and the adjusting agent I and the de-magnesiumization collector are added in sequence to obtain the de-magnesiumized concentrate a and the de-magnesiumized tailings b; the de-magnesiumized concentrate a enters the gravity separation to obtain the gravity separation concentrate c and the gravity separation tailings d; the gravity separation concentrate c is introduced into the gravity concentration reverse flotation dealumination process, and the dealumination collector is added to obtain the phosphate concentrate e and tailings f; the gravity separation tailings d is introduced into the heavy tailing reverse flotation dealumination process, and the adjusting agent II and the dealumination collector are added in sequence to obtain the phosphate concentrate g and tailings h.

[0007] B. Backwater reuse: (1) Direct reuse of return water: The return water obtained by collecting the de-magnesiumized tailings b, phosphate concentrate e, tailings f and phosphate concentrate g and concentrating and filtering them respectively enters the No. 1 return water tank and then directly returns to the grinding, reverse flotation de-magnesiumization process, gravity separation, gravity concentration reverse flotation dealuminization process for reuse; (2) Return water treatment and reuse: The return water obtained by collecting the tailings h and thickening and filtering enters the No. 2 return water tank, and then is treated with montmorillonite and returned to the heavy tailings reverse flotation dealuminization process for reuse.

[0008] In the above scheme, adjusting agent I is added during the reverse flotation demagnesium process, and the adjusting agent I is a mixture of phosphoric acid and raffinate acid in a mass ratio of 3:1-1:1. Phosphoric acid and raffinate acid are industrial-grade products. The P2O5 content in phosphoric acid is 20 wt%-42 wt%, and the P2O5 content in raffinate acid is 20 wt%-38 wt%.

[0009] In the above scheme, the addition amount of adjusting agent I is 6-10kg / t phosphate slurry dry basis.

[0010] In the above scheme, an adjusting agent II is added in the heavy tail reverse flotation dealumination process, and the adjusting agent II is composed of the following components in percentage by mass: 60-80% aluminum sulfate, 5-15% polyethylene glycol, and 5-15% cationic starch flocculant, and the sum of the percentages by mass of the components is 100%.

[0011] In the above scheme, the preparation process of the cationic starch flocculant is as follows: Step 1. After the corn starch is gelatinized at high temperature and cooled to room temperature, nitrogen is introduced, and then an initiator ammonium cerium nitrate and DMC are added dropwise to react to obtain grafted starch; Step 2. Add 3-chloro-2-hydroxypropyltrimethylammonium chloride (CTA) dropwise into a NaOH solution in an ice water bath, and stir to prepare an etherifying agent GTA; Step 3. Add the grafted starch obtained in step 1 to the ethanol solution, add the etherifying agent GTA obtained in step 2 dropwise at 50-70°C with stirring, then slowly add NaOH as a catalyst, and react at a constant temperature for 3-6 h to obtain the cationic starch flocculant.

[0012] In step 1, the mass ratio of corn starch to dimethyl carbonate DMC is 1:0.8-1:1.2, and the added amount of ammonium cerium nitrate is 0.06-0.12% of the mass of corn starch.

[0013] In step 2, the molar ratio of NaOH (5-10 mol·L-1) solution to 3-chloro-2-hydroxypropyltrimethylammonium chloride is 2.5-3:1.

[0014] The mass ratio of grafted starch to etherifying agent GTA in step 3 is 1:0.6-1:1.2. In the above scheme, the amount of adjusting agent II added is 2-4 kg / t dry basis of phosphate slurry.

[0015] In the above scheme, the montmorillonite is a mixture of calcium montmorillonite and sodium montmorillonite in a mass percentage ratio of 3:1-1:1, and the addition amount is 1-3kg / ton of return water.

[0016] In the above scheme, the direct reuse of return water refers to the return water that does not need to be treated during the beneficiation process of collophanite before it can be used as mineral processing circulating water; the treated return water reuse refers to the return water that needs to be treated during the beneficiation process of collophanite before it can be used as mineral processing circulating water; the mass ratio of the direct reuse of return water to the treated return water reuse is 8:1-10:1.

[0017] Compared with the prior art, the present invention has the following beneficial effects: (1) The present invention adopts the combined re-flotation beneficiation technology to realize the separate flotation of high-grade coarse-grained gravity separation concentrate and low-grade fine-grained gravity separation tailings, thereby reducing the adverse effects of the inclusion of the two materials in the flotation pulp on flotation. During the flotation process, the low-grade fine-grained particles and the flotation reagent often form a three-phase stable state of gas, solid and liquid, which causes the foam produced during the flotation process to have high viscosity and be difficult to break, thereby affecting the flotation index and foam fluidity. The requirements for return water for the flotation of high-grade coarse-grained minerals are much lower than those for low-grade fine-grained minerals. Therefore, in the present invention, except for the dense and filtered water of the tailings obtained by the dealumination process of the low-grade fine-grained gravity separation tailings, which need to be treated before entering the mineral processing system for recycling due to their high content of residual reagents, COD content, and suspended particles, the other return water can be uniformly collected and directly reused. It effectively reduces the amount of return water that needs to be processed during the separation of collophanite and its impurities of magnesium and aluminum, realizes the full-process return water reuse without affecting the separation indicators of collophanite for removing impurities of magnesium and aluminum, greatly simplifies the return water treatment plan, and reduces the return water treatment and utilization cost.

[0018] (2) The present invention achieves dealumination flotation of fine-grained particles through the adjusting agent II, reduces the viscosity of the flotation foam, and greatly improves the fluidity of the foam. The mechanism of action of the adjusting agent in the present invention to adjust the foam can be roughly summarized as follows: first, the cationic starch flocculant compresses the electrical double layer of the fine-grained phosphate mineral particles, effectively reduces the repulsive force between the mineral particles, and can capture organic suspended impurities in the water and treat organic and inorganic substances with floating points in the water; second, after aluminum sulfate is added to the slurry, aluminum hydroxide floccules can be generated to adsorb a large number of fine-grained mineral particles; third, polyethylene glycol can assist in defoaming. Due to the synergistic effect of the cationic starch flocculant, aluminum sulfate and polyethylene glycol, the electronegativity of the fine-grained particles in the slurry system is greatly weakened, which weakens the effect of the fine-grained particles and the amine cationic collector, resulting in a decrease in the number of electronegative particles in the foam, a worsening of the foam toughness, and an improvement in fluidity.

[0019] (3) After the return water in the present invention is treated with montmorillonite, the residual reagents, COD content and suspended particle content in the return water can be further reduced, which can effectively eliminate the adverse effects of the return water on the mineral processing indicators and obtain high-quality phosphate concentrate. In addition, montmorillonite is widely available and inexpensive.

[0020] (4) The method of the present invention is simple, requires few types of reagents, and has low cost. It can realize the full reuse of the return water in the whole process, and does not affect the separation index of the impurities magnesium and aluminum in the collophanite, and has good social and economic benefits. BRIEF DESCRIPTION OF THE DRAWINGS

[0021] Figure 1 The present invention discloses a process flow chart of a method for utilizing return water from the refloatation and combined beneficiation of collophanite. DETAILED DESCRIPTION

[0022] In order to better understand the present invention, the content of the present invention is further explained below in conjunction with the examples, but the content of the present invention is not limited to the following examples. Any simple modification and equivalent structural transformation or modification made according to the essential content of the present invention should be included in the protection scope of the present invention.

[0023] Unless otherwise specified in the following examples, the reagents used are commercially available chemical reagents or industrial products.

[0024] Raffinate acid: P2O5 content 31.5wt%, MgO content 6.42wt%, Al2O3 content 5.8wt%, Fe2O3 content 2.2wt%, CaO content <0.05wt%, K2O content <0.05wt%, Na2O content <1.0wt%, solid content 6.4wt%, pH value less than 1.1; density of phosphoric acid raffinate clear liquid 2.0g / cm 3 .

[0025] Phosphoric acid is an industrial grade product, and the P2O5 content in phosphoric acid is 36 wt %.

[0026] Adjuster I is a mixture of phosphoric acid and raffinate acid in a mass ratio of 2:1.

[0027] The adjusting agent II is composed of the following components in the following mass percentages: 70% aluminum sulfate, 15% polyethylene glycol, and 15% cationic starch flocculant.

[0028] In the above scheme, the preparation process of the cationic starch flocculant is as follows: step 1. After the corn starch is gelatinized at high temperature and cooled to room temperature, nitrogen is introduced, and then initiators ammonium cerium nitrate and dimethyl carbonate DMC are added dropwise to react to obtain grafted starch; the mass ratio of corn starch to dimethyl carbonate DMC is 1:1.1, and the amount of ammonium cerium nitrate added is 0.1% of the mass of corn starch.

[0029] Step 2. Add NaOH (10 mol·L -1 ) solution was added dropwise in an ice-water bath to 3-chloro-2-hydroxypropyltrimethylammonium chloride (CTA), and stirred to obtain the etherifying agent GTA; NaOH (10 mol·L -1) The molar ratio of the solution to 3-chloro-2-hydroxypropyltrimethylammonium chloride is 2.6:1.

[0030] Step 3. Add ethanol solution as solvent to the grafted starch obtained in step 1, add the etherifying agent GTA obtained in step 2 dropwise at 50-70°C with stirring, then slowly add NaOH solution as catalyst, and react at constant temperature for 6 hours to obtain a cationic starch flocculant. The mass ratio of grafted starch to etherifying agent GTA is 1:1.2.

[0031] The magnesium removal collector is fatty acid soap.

[0032] The dealuminized collector is: dodecylamine.

[0033] The montmorillonite is a mixture of calcium-montmorillonite and sodium-montmorillonite in a mass percentage ratio of 2:1.

[0034] Example 1 A method for utilizing the return water of the combined refloatation and beneficiation of collophanite. The P2O5 grade of the collophanite in this embodiment is 24wt%, and it is mainly composed of three types of minerals: phosphate, dolomite and silicate. Figure 1 As shown, the method for utilizing the return water of the collophanite refloatation combined ore dressing comprises the steps of collophanite refloatation combined ore dressing and return water utilization, specifically comprising: A. Refloatation combined beneficiation of collophanite: After grinding and classification, the collophanite is introduced into the reverse flotation demagnesization process, and the adjusting agent I is added in an amount of 8kg / t phosphate slurry dry basis, and the demagnesization collector is added in an amount of 0.8kg / t phosphate slurry dry basis to obtain demagnesized concentrate a and demagnesized tailings b; the demagnesized concentrate a enters the re-selection to obtain the re-selection concentrate c and the re-selection tailings d. The fineness of the re-selection concentrate c is -0.074mm, accounting for 65wt%, and the P2O5 grade is 28.5wt%. The fineness of the re-selection tailings d is -0.074mm, accounting for 85wt%. , P2O5 grade is 22wt%; the gravity separation concentrate c is introduced into the heavy concentrate reverse flotation dealumination process, and a dealumination collector is added, and the amount of dealumination collector added is 0.3kg / t phosphate slurry dry basis to obtain phosphate concentrate e and tailings f; the gravity separation tailings d is introduced into the heavy tailing reverse flotation dealumination process, and adjusting agent II is added in an amount of 4kg / t phosphate slurry dry basis and dealumination collector is added in an amount of 0.2kg / t phosphate slurry dry basis to obtain phosphate concentrate g and tailings h.

[0035] B. Backwater reuse: (1) Direct reuse of return water: The return water obtained by collecting the de-magnesiumized tailings b, phosphate concentrate e, tailings f and phosphate concentrate g and concentrating and filtering them respectively enters the No. 1 return water tank and then directly returns to the grinding, reverse flotation de-magnesiumization process, gravity separation, gravity concentration reverse flotation dealuminization process for reuse; (2) Treatment and reuse of return water: After the return water obtained by collecting tailings h and thickening and filtering enters the No. 2 return water pool, montmorillonite is added. The amount of montmorillonite added is 1 kg / ton of return water. After treatment, it is returned to the heavy tail reverse flotation dealuminization process for reuse.

[0036] The obtained phosphate concentrate e has a phosphorus grade of 33.6% and a yield of 51.6%; the obtained phosphate concentrate g has a phosphorus grade of 30.5% and a yield of 14.6%.

[0037] The process steps are the same as those in Example 1, except that the polyethylene glycol is not added to the adjusting agent II. The achieved effects are as follows: the phosphorus grade of the phosphate concentrate e is 33.4%, and the yield is 50.7%; the phosphorus grade of the phosphate concentrate g is 28.16%, and the yield is 15.6%.

[0038] The process steps are the same as those in Example 1, except that the cationic starch flocculant is not added to the adjusting agent II. The achieved effects are as follows: the phosphorus grade of the phosphate concentrate e is 33%, and the yield is 48.3%; the phosphorus grade of the phosphate concentrate g is 26.3%, and the yield is 18.4%.

[0039] The process steps are the same as those in Example 1, except that the cationic starch flocculant added to the adjusting agent II does not contain DMC in the preparation. The achieved results are as follows: the phosphorus grade of phosphate concentrate e is 33.2%, and the yield is 50.7%; the phosphorus grade of phosphate concentrate g is 27.0%, and the yield is 16.2%.

[0040] The process steps are the same as those in Example 1, except that the cationic starch flocculant in the adjusting agent II is replaced by polyacrylamide. The achieved effects are as follows: the phosphorus grade of phosphate concentrate e is 32.8%, and the yield is 49.5%; the phosphorus grade of phosphate concentrate g is 29.1%, and the yield is 10.6%.

[0041] Example 2 A method for utilizing the return water of the combined refloatation and beneficiation of collophanite. The P2O5 grade of the collophanite in this embodiment is 26.5wt%, and it is mainly composed of three types of minerals: phosphate, dolomite and silicate. Figure 1 As shown, the method for utilizing the return water of the collophanite refloatation combined ore dressing comprises the steps of collophanite refloatation combined ore dressing and return water utilization, specifically comprising: A. Refloatation combined beneficiation of collophanite: After grinding and classification, the collophanite is introduced into the reverse flotation demagnesization process, and the adjusting agent I is added in an amount of 8kg / t phosphate slurry dry basis, and the demagnesization collector is added in an amount of 0.8kg / t phosphate slurry dry basis to obtain demagnesized concentrate a and demagnesized tailings b; the demagnesized concentrate a enters the re-selection to obtain the re-selection concentrate c and the re-selection tailings d. The fineness of the re-selection concentrate c is -0.074mm, accounting for 62wt%, and the P2O5 grade is 29.5wt%. The fineness of the re-selection tailings d is -0.074mm, accounting for 85wt%. , P2O5 grade is 24wt%; the gravity separation concentrate c is introduced into the heavy concentrate reverse flotation dealumination process, and a dealumination collector is added, and the amount of dealumination collector added is 0.3kg / t phosphate slurry dry basis to obtain phosphate concentrate e and tailings f; the gravity separation tailings d is introduced into the heavy tailing reverse flotation dealumination process, and adjusting agent II is added in an amount of 4kg / t phosphate slurry dry basis and dealumination collector is added in an amount of 0.2kg / t phosphate slurry dry basis to obtain phosphate concentrate g and tailings h.

[0042] B. Backwater reuse: (1) Direct reuse of return water: The return water obtained by collecting the de-magnesiumized tailings b, phosphate concentrate e, tailings f and phosphate concentrate g and concentrating and filtering them respectively enters the No. 1 return water tank and then directly returns to the grinding, reverse flotation de-magnesiumization process, gravity separation, gravity concentration reverse flotation dealuminization process for reuse; (2) Treatment and reuse of return water: After the return water obtained by collecting tailings h and thickening and filtering enters the No. 2 return water pool, montmorillonite is added. The amount of montmorillonite added is 1 kg / ton of return water. After treatment, it is returned to the heavy tail reverse flotation dealuminization process for reuse.

[0043] The obtained phosphate concentrate e has a phosphorus grade of 33.8% and a yield of 53.7%; the obtained phosphate concentrate g has a phosphorus grade of 31.2% and a yield of 15.4%.

[0044] Example 2-1 A method for utilizing the return water of the refloatation and combined beneficiation of collophanite. This example is a blank example of Example 2. The P2O5 grade in the collophanite is 26.5wt%, and it is mainly composed of three types of minerals: phosphate, dolomite and silicate. The specific technical implementation plan is: A. Double reverse flotation of collophanite: the collophanite is introduced into the reverse flotation de-magnesiumization process after grinding and classification, and the adjusting agent I is added in an amount of 8 kg / t phosphate slurry dry basis and the de-magnesiumization collector is added in an amount of 0.8 kg / t phosphate slurry dry basis to obtain de-magnesiumized concentrate a and de-magnesiumized tailings b; the de-magnesiumized concentrate a enters the reverse flotation dealumination process, and the adjusting agent II is added in an amount of 4 kg / t phosphate slurry dry basis and the dealumination collector is added in an amount of 0.2 kg / t phosphate slurry dry basis to obtain phosphate concentrate c and tailings d.

[0045] B. Backwater reuse: Direct reuse of return water: The return water obtained by collecting the de-magnesiumized tailings b, phosphate concentrate c and tailings d and thickening and filtering them respectively enters the return water pool and then directly returns to the grinding, reverse flotation de-magnesiumization process and reverse flotation dealuminization process for reuse; The phosphorus grade of the phosphate concentrate obtained was 32.04% and the yield was 48.4%. During the test, the flotation foam was very sticky.

[0046] Example 2-2 A method for utilizing the return water of the refloatation and combined beneficiation of collophanite. This example is a blank example of Example 2. The P2O5 grade in the collophanite is 26.5wt%, and it is mainly composed of three types of minerals: phosphate, dolomite and silicate. The specific technical implementation plan is: A. Double reverse flotation of collophanite: the phosphate pulp obtained after grinding and classification of collophanite is introduced into the reverse flotation de-magnesiumization process, and adjusting agent I is added in an amount of 8 kg / t phosphate pulp dry basis and de-magnesiumization collector is added in an amount of 0.8 kg / t phosphate pulp dry basis to obtain de-magnesiumized concentrate a and de-magnesiumized tailings b; the de-magnesiumized concentrate a enters the reverse flotation dealumination process, and adjusting agent II is added in an amount of 4 kg / t phosphate pulp dry basis and dealumination collector is added in an amount of 0.2 kg / t phosphate pulp dry basis to obtain phosphate concentrate c and tailings d.

[0047] B. Backwater reuse: Direct reuse of return water: After collecting the magnesium removal tailings b, phosphate concentrate c and tailings d, the return water obtained by thickening and filtering enters the return water pool, montmorillonite is added, and the amount of montmorillonite added is 1kg / ton of return water. After treatment, it is returned to the grinding, reverse flotation magnesium removal process, and reverse flotation dealuminization process for reuse; The phosphorus grade of the phosphate concentrate obtained was 32.64%, and the yield was 53.2%. During the test, the flotation foam was sticky. Compared with Example 2-1, the stickiness of the foam was improved to a certain extent.

[0048] Example 3 A method for utilizing the return water of the combined refloatation and beneficiation of collophanite. The P2O5 grade of the collophanite in this embodiment is 26.5wt%, and it is mainly composed of three types of minerals: phosphate, dolomite and silicate. Figure 1 As shown, the method for utilizing the return water of the collophanite refloatation combined ore dressing comprises the steps of collophanite refloatation combined ore dressing and return water utilization, specifically comprising: A. Refloatation and combined separation of collophanite: After grinding and classification, the collophanite is introduced into the reverse flotation demagnesization process, and the adjusting agent I is added in an amount of 8kg / t phosphate slurry dry basis, and the demagnesization collector is added in an amount of 0.8kg / t phosphate slurry dry basis to obtain demagnesized concentrate a and demagnesized tailings b; the demagnesized concentrate a enters the re-separation to obtain the re-separation concentrate c and the re-separation tailings d. The fineness of the re-separation concentrate c is -0.074mm, accounting for 62wt%, and the P2O5 grade is 29.5wt%. The fineness of the re-separation tailings d is -0.074mm, accounting for 85wt%, and the P2O5 grade is 24wt%. ; The gravity separation concentrate c is introduced into the heavy concentrate reverse flotation dealumination process, and a dealumination collector is added, and the amount of dealumination collector added is 0.3kg / t phosphate slurry dry basis to obtain phosphate concentrate e and tailings f; the gravity separation tailings d is introduced into the heavy tailing reverse flotation dealumination process, and a dealumination collector is added, and the amount of dealumination collector added is 0.2kg / t phosphate slurry dry basis to obtain phosphate concentrate g and tailings h.

[0049] B. Backwater reuse: (1) Direct reuse of return water: The return water obtained by collecting the de-magnesiumized tailings b, phosphate concentrate e, tailings f and phosphate concentrate g and concentrating and filtering them respectively enters the No. 1 return water tank and then directly returns to the grinding, reverse flotation de-magnesiumization process, gravity separation, gravity concentration reverse flotation dealuminization process for reuse; (2) Treatment and reuse of return water: After the return water obtained by collecting tailings h and thickening and filtering enters the No. 2 return water pool, montmorillonite is added. The amount of montmorillonite added is 1 kg / ton of return water. After treatment, it is returned to the heavy tail reverse flotation dealuminization process for reuse.

[0050] The obtained phosphate concentrate e had a phosphorus grade of 33.6% and a yield of 53.6%; the obtained phosphate concentrate g had a phosphorus grade of 28.0% and a yield of 13.1%. During the test, the flotation index of the heavy tail reverse flotation dealuminization process was not ideal, and the foam volume was large and sticky.

[0051] Example 4 A method for utilizing the return water of the combined refloatation and beneficiation of collophanite. The P2O5 grade of the collophanite in this embodiment is 26.5wt%, and it is mainly composed of three types of minerals: phosphate, dolomite and silicate. Figure 1 As shown, the method for utilizing the return water of the collophanite refloatation combined ore dressing comprises the steps of collophanite refloatation combined ore dressing and return water utilization, specifically comprising: A. Refloatation and combined separation of collophanite: After grinding and classification, the collophanite is introduced into the reverse flotation demagnesization process, and the adjusting agent I is added in an amount of 8kg / t phosphate slurry dry basis, and the demagnesization collector is added in an amount of 0.8kg / t phosphate slurry dry basis to obtain demagnesized concentrate a and demagnesized tailings b; the demagnesized concentrate a enters the re-separation to obtain the re-separation concentrate c and the re-separation tailings d. The fineness of the re-separation concentrate c is -0.074mm, accounting for 62wt%, and the P2O5 grade is 29.5wt%. The fineness of the re-separation tailings d is -0.074mm, accounting for 85wt%. , P2O5 grade is 24wt%; the gravity separation concentrate c is introduced into the heavy concentrate reverse flotation dealumination process, and a dealumination collector is added, and the amount of dealumination collector added is 0.3kg / t phosphate slurry dry basis to obtain phosphate concentrate e and tailings f; the gravity separation tailings d is introduced into the heavy tailing reverse flotation dealumination process, and a dealumination collector is added, and the amount of dealumination collector added is 0.2kg / t phosphate slurry dry basis to obtain phosphate concentrate g and tailings h.

[0052] B. Backwater reuse: After collecting the de-magnesiumized tailings b, phosphate concentrate e, tailings f and phosphate concentrate g, tailings h and thickening and filtering the returned water into the return water pool, montmorillonite is added, and the amount of montmorillonite added is 1 kg / ton of returned water. After treatment, it is returned to the heavy tailings reverse flotation dealuminization process for reuse.

[0053] The obtained phosphate concentrate e has a phosphorus grade of 34% and a yield of 552%; the obtained phosphate concentrate g has a phosphorus grade of 30.5% and a yield of 15.4%. The flotation foam flows smoothly and the flotation indicators are ideal, but the return water treatment volume is large and the treatment cost is high.

[0054] Compared with the prior art, this specific implementation has the following positive effects: Compared with the prior art, the present invention has the following beneficial effects: 1) The present invention adopts the re-flotation combined beneficiation technology to realize the separate flotation of high-grade coarse-grained gravity separation concentrate and low-grade fine-grained gravity separation tailings, reducing the adverse effects of the inclusion of the two materials in the flotation pulp on flotation. In the flotation process, the low-grade fine-grained particles and the flotation reagent often form a three-phase stable state of gas, solid and liquid, which causes the foam produced in the flotation process to have high viscosity and is not easy to break, thereby affecting the flotation index and foam fluidity, and the flotation of high-grade coarse-grained minerals has much lower requirements for return water than low-grade fine-grained minerals. Therefore, in the present invention, except for the dense and filtered water of the tailings obtained by the dealumination process of the low-grade fine-grained gravity separation tailings, which needs to be treated before entering the mineral processing system for recycling due to its high content of residual reagents, COD content, and suspended particles, the other return water can be uniformly collected and directly reused. It effectively reduces the amount of return water that needs to be processed during the separation of collophanite and its impurities of magnesium and aluminum, realizes the full-process return water reuse without affecting the separation indicators of collophanite for removing impurities of magnesium and aluminum, greatly simplifies the return water treatment plan, and reduces the return water treatment and utilization cost.

[0055] 2) The present invention realizes the dealumination flotation of fine-grained particles through the adjusting agent II, reduces the viscosity of the flotation foam, and greatly improves the fluidity of the foam. The mechanism of action of the adjusting agent in the present invention to adjust the foam can be roughly summarized as follows: first, the cationic starch flocculant compresses the electrical double layer of the fine-grained phosphate mineral particles, effectively reduces the repulsive force between the mineral particles, and can capture organic suspended impurities in the water, and treat organic and inorganic substances with floating points in the water; second, after aluminum sulfate is added to the slurry, aluminum hydroxide floccules can be generated to adsorb a large number of fine-grained mineral particles; third, polyethylene glycol can assist in defoaming. Due to the synergistic effect of the cationic starch flocculant, aluminum sulfate and polyethylene glycol, the electronegativity of the fine-grained particles in the slurry system is greatly weakened, and the effect of the fine-grained particles and the amine cationic collector is weakened, so that the electronegative particles in the foam are reduced, the foam toughness is worse, and the fluidity is improved.

[0056] 3) After the return water in the present invention is treated with montmorillonite, the residual reagents, COD content and suspended particle content in the return water can be further reduced, which can effectively eliminate the adverse effects of the return water on the mineral processing indicators and obtain high-quality phosphate concentrate. Montmorillonite is widely available and inexpensive.

[0057] 4) The method of the present invention is simple, requires few types of reagents, and has low cost. It can realize the full reuse of the return water in the whole process, and does not affect the separation index of the collophanite to remove the impurities of magnesium and aluminum, and has good social and economic benefits.

Claims

1. A method for utilizing the return water of collophanite refloatation combined ore dressing, characterized in that: The steps include: A. Combined beneficiation of collophanite by re-floatation: After grinding and classification, the collophanite is introduced into the reverse flotation de-magnesiumization process to obtain de-magnesiumized concentrate a and de-magnesiumized tailings b; the de-magnesiumized concentrate a enters the re-separation to obtain the re-separation concentrate c and the re-separation tailings d; the re-separation concentrate c is introduced into the re-concentration reverse flotation de-aluminization process to obtain the phosphate concentrate e and tailings f; the re-separation tailings d is introduced into the re-tailing reverse flotation de-aluminization process to obtain the phosphate concentrate g and tailings h; B. Backwater reuse: (1) Direct reuse of return water: The return water obtained by collecting the magnesium removal tailings b, phosphate concentrate e, tailings f and phosphate concentrate g and concentrating and filtering them respectively enters the No. 1 return water tank and is directly reused in the grinding, reverse flotation magnesium removal process, gravity separation, and gravity concentration reverse flotation aluminum removal process; (2) Return water treatment and reuse: The return water obtained by collecting the tailings h and thickening and filtering enters the No. 2 return water tank, and then is treated with montmorillonite and returned to the heavy tailings reverse flotation dealuminization process for reuse.

2. The method for utilizing the return water of the collophanite refloatation combined beneficiation according to claim 1, characterized in that: During the reverse flotation demagnesium process, adjusting agent I is added, wherein the adjusting agent I is a mixture of phosphoric acid and raffinate acid in a mass ratio of 3:1-1:1, the phosphoric acid and raffinate acid are industrial-grade products, the P2O5 content in the phosphoric acid is 20 wt%-42 wt%, and the P2O5 content in the raffinate acid is 20 wt%-38 wt%.

3. The method for utilizing the return water of the collophanite refloatation combined ore dressing according to claim 1, characterized in that: The addition amount of adjusting agent I is 6-10kg / t phosphate slurry dry basis.

4. The method for utilizing the return water of the collophanite refloatation combined beneficiation according to claim 1, characterized in that: In the heavy tail reverse flotation dealumination process, an adjusting agent II is added, and the adjusting agent II is composed of the following components in mass percentage: 60-80% aluminum sulfate, 5-15% polyethylene glycol, and 5-15% cationic starch flocculant, and the sum of the mass percentages of the components is 100%.

5. The method for utilizing the return water of the collophanite refloatation combined beneficiation according to claim 4, characterized in that: The preparation process of the cationic starch flocculant is as follows: Step 1. After the corn starch is gelatinized at high temperature and cooled to room temperature, nitrogen is introduced, and then an initiator ammonium cerium nitrate and DMC are added dropwise to react to obtain grafted starch; Step 2. Add 3-chloro-2-hydroxypropyltrimethylammonium chloride (CTA) dropwise into a NaOH solution in an ice water bath, and stir to prepare an etherifying agent GTA; Step 3. Add the grafted starch obtained in step 1 to the ethanol solution, add the etherifying agent GTA obtained in step 2 dropwise at 50-70°C with stirring, then slowly add NaOH as a catalyst, and react at a constant temperature for 3-6 h to obtain the cationic starch flocculant.

6. The method for utilizing the return water of the collophanite refloatation combined ore dressing according to claim 5, characterized in that: In step 1, the mass ratio of corn starch to dimethyl carbonate DMC is 1:0.8-1:1.2, and the amount of cerium ammonium nitrate added is 0.06-0.12% of the mass of corn starch; In step 2, the molar ratio of 5-10 mol·L-1 NaOH solution to 3-chloro-2-hydroxypropyltrimethylammonium chloride is 2.5-3:1; In step 3, the mass ratio of grafted starch to etherifying agent GTA is 1:0.6-1:1.

2.

7. The method for utilizing the return water of the collophanite refloatation combined beneficiation according to claim 4, characterized in that: The amount of adjusting agent II added is 2-4 kg / t phosphate slurry dry basis.

8. The method for utilizing the return water of collophanite refloatation combined ore dressing according to claim 1, characterized in that: The montmorillonite is a mixture of calcium montmorillonite and sodium montmorillonite in a mass percentage ratio of 3:1-1:1, and the addition amount thereof is 1-3 kg / ton of return water.

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