Collophanite re-flotation combined selection and middle grading quality-grading flotation method

Abstract

The invention belongs to the technical field of phosphorite beneficiation, and relates to a collophanite re-flotation combined selection and middle grading quality-grading flotation method. According to the technical scheme, the method comprises the steps that phosphoric ore pulp is fed into a spiral chute set to be reselected, reselected products including reselected concentrate and reselected tailings are obtained, then the reselected concentrate obtained through separation is fed into a flotation dealumination procedure, reselected concentrate and tailings are obtained, the tailings obtained after reselected flotation are fed into a grading and grading procedure, and tailings 1 and middlings are obtained. And then the middlings and the spiral chute heavy tailings are combined and enter a heavy tailing flotation dealumination procedure, and heavy tailing concentrates and tailings 2 are obtained. The process has the advantages that the mixed heavy, fine and heavy tailings can be effectively classified in the industrial production, so that the flotation comprehensive recovery rate in the industrial production is effectively improved, and the flotation auxiliary material cost is reduced.

Description

TECHNICAL FIELD

[0001] The application belongs to the technical field of phosphate ore dressing, and particularly relates to a method for adding an intermediate grading and quality separation flotation to collophanite gravity-flotation combined separation. BACKGROUND

[0002] In China, the reserves of medium and low grade phosphate ore are abundant, among which the reserves of siliceous and silicate type phosphate ore and mixed collophanite account for more than 75%. At present, collophanite is generally separated by flotation (such as single reverse flotation, direct-reverse flotation, double reverse flotation), gravity separation, and gravity-flotation combined separation. In recent years, the gravity-flotation combined separation has been greatly developed because it can overcome the problems of low yield and sticky foam in traditional direct-reverse and double reverse flotation. The gravity-flotation combined separation of collophanite uses a spiral chute to pretreat the phosphate ore, divides the raw ore into heavy concentrate with high grade and coarse particle size easy to float and heavy tailings with low grade and fine particle size, and then uses different flotation processes and control conditions for targeted flotation, so that a high comprehensive concentrate yield can be obtained. However, due to the limited processing capacity of a single spiral chute, thousands of spiral chutes need to be used in parallel in the industrial application of phosphate ore dressing to meet the processing capacity requirement. Due to the large number of spiral chutes, it is difficult to realize real-time adjustment of the strip interception of the spiral chutes in the actual production control process. Therefore, the separation quality of the heavy concentrate and the heavy tailings of the spiral chute will be affected by the fluctuations of factors such as flow rate and concentration in the production process, resulting in the inclusion of more low-grade fine particles in the heavy concentrate or the inclusion of more high-grade coarse particles in the heavy tailings, thereby causing the deterioration of the subsequent flotation effect.

[0003] The patent technology of "a high-multiple collophanite spiral chute gravity-flotation combined process (CN 109453891 B)" discloses a semi-collophanite spiral chute gravity-flotation combined process, and the patent technology of "a phosphate ore gravity-flotation combined beneficiation process (CN 110369120 A)" also discloses a phosphate ore gravity-flotation combined beneficiation process. Both of the above two processes include the separation of finely ground phosphate ore by a spiral chute, and the double reverse flotation and direct-reverse flotation of the obtained gravity separation concentrate and gravity separation tailings. However, the above processes cannot solve the problem that the large number of spiral chutes in the industrial application process, the fluctuations of factors such as flow rate and concentration cause the inclusion of more low-grade fine particles in the heavy concentrate or the inclusion of more high-grade coarse particles in the heavy tailings, thereby causing the deterioration of the subsequent flotation effect.

[0004] The patent technology "Method for heavy and floatation combined separation of collophanite by multi-grain product spiral chute (CN 117258997 A)" discloses a heavy and floatation combined separation method; the process uses a spiral chute to divide the raw ore into heavy concentrate, heavy tailings and an intermediate product, wherein the intermediate grain product is returned to a ball mill for regrinding, which can solve the problem of mutual inclusion of heavy concentrate and heavy tailings to a certain extent, but cannot solve the problem of deteriorated floatation effect, and the intermediate product is mainly low-grade coarse-grained feldspar minerals, which will inevitably produce more fine-grained feldspar minerals after returning to regrinding, which will adversely affect the subsequent fine-grained floatation.

[0005] The patent technology "Large-scale process method for collophanite spiral chute (CN 115921097 B)" discloses a large-scale industrial method for heavy and floatation combined separation of collophanite spiral chute, which can enlarge the diameter of the spiral chute to obtain a separation effect close to that of a small-diameter chute, but the diameter of the spiral chute has certain limitations, and a large number of spiral chutes are still needed for separation in large-scale industrial production, which cannot solve the problem of deterioration of the subsequent floatation effect caused by the large number of spiral chutes, flow rate, concentration and other factors. SUMMARY

[0006] The purpose of the present application is to provide a collophanite heavy and floatation combined separation method with an added intermediate classification and quality separation process, which can eliminate the influence of poor overall separation effect caused by the large number of spiral chutes, the inability to adjust the strip in real time, and the fluctuation of flow rate, concentration and other factors in the industrial application of heavy and floatation combined separation.

[0007] A collophanite heavy and floatation combined separation method with an added intermediate classification and quality separation process, comprising the following steps: S1, feeding the phosphate ore slurry into a spiral chute group for heavy separation to obtain heavy concentrate and heavy tailings; S2, feeding the heavy concentrate into a floatation aluminum removal process to obtain heavy concentrate concentrate and heavy concentrate floatation tailings; S3, feeding the heavy concentrate floatation tailings into a classification and quality separation process to obtain tailings 1 and middlings; S4, combining the middlings and the heavy tailings to enter a heavy tailings floatation aluminum removal process to obtain heavy tailings concentrate and tailings 2.

[0008] Further, the phosphate ore slurry has a concentration of 15-35%, a P2O5 grade of 18.0-27.5wt%, an Al2O3 content of 2.5-5.5wt%, and a particle size of less than 0.074mm accounting for 45-85%.

[0009] Furthermore, in the flotation dealuminization process described in S2, the added dealuminizing collector is one or more of dodecyl ether amine, benzyl dimethyl tetradecyl ammonium chloride, or trimethylchlorosilane; preferably, it comprises the following components by mass percentage: 25-45% dodecyl ether amine, 30-50% benzyl dimethyl tetradecyl ammonium chloride, and 20-35% trimethylchlorosilane.

[0010] Furthermore, in the flotation dealuminization process described in S2, the amount of flotation reagent used is 200-400 g / t.

[0011] Furthermore, the flotation dealuminization process described in S2 is reverse flotation dealuminization, with the flotation concentration for heavy fine dealuminization controlled at 20-30%.

[0012] Furthermore, the flotation dealuminization process described in S4 is reverse flotation dealuminization, with the flotation concentration for heavy tail dealuminization controlled at 5-15%.

[0013] Furthermore, the equipment used in the grading and sorting process is a high-frequency screen or a hydrocyclone.

[0014] Furthermore, the grading and sorting process ensures that the proportion of particles with a diameter less than 0.023 mm in the tailings 1 after heavy flotation is ≤20%.

[0015] Furthermore, in the flotation dealuminization process described in S4, the added dealuminizing collector is one or more of dodecyl ether amine, benzyl dimethyl tetradecyl ammonium chloride, or trimethylchlorosilane; preferably, it comprises the following components by mass percentage: 25-45% dodecyl ether amine, 30-50% benzyl dimethyl tetradecyl ammonium chloride, and 20-35% trimethylchlorosilane.

[0016] Furthermore, in the flotation dealuminization process described in S4, the amount of flotation reagent used is 50-200 g / t.

[0017] The beneficial effects of this invention are as follows: In practical industrial applications of combined gravity and flotation, gravity separation typically involves multiple spiral sluices connected in parallel. A separator distributes the slurry evenly to each spiral sluice. Due to the large number of spiral sluices, the cut-off strips are often fixed after adjustment according to production requirements. However, during production, flow rate, concentration, or blockage of the separator can cause some spiral sluice separators to be out of optimal position, leading to mixing of heavy concentrate and tailings, and deteriorating flotation performance. Therefore, the aforementioned technical solution can effectively reduce the impact on flotation performance and has the following main positive effects.

[0018] The gangue minerals in collophane are mainly silicate feldspar minerals and quartz, which are harder than the valuable mineral apatite and have a larger particle size after grinding. After gravity separation in a spiral sluice, most of the coarse-grained gangue minerals enter the heavy cleaner. When heavy tailings are mixed into the heavy cleaner, it will greatly affect the separation of valuable minerals and gangue minerals. The inclusion of high-grade fine particles in the tailings will also lead to a high tailings grade. Adding a classification process after heavy cleaning flotation reduces the proportion of particles smaller than 0.023mm in the tailings to below 20%, while increasing the tailings concentration. This can effectively improve the subsequent tailings scavenging and separation effect and increase the overall recovery rate.

[0019] The fine-grained product after heavy tailings classification and grading is characterized by low concentration, fine particle size, and high reagent residue. Combining it with heavy tailings for flotation meets the requirement of low concentration for fine-grained phosphate rock beneficiation. At the same time, the residual reagent can reduce the amount of reagent used in heavy tailings flotation, thus reducing the cost of auxiliary materials in production.

[0020] In this invention, the grade of heavy fine tailings (i.e., tailings 1 in the examples) can be reduced by 2-5 wt% compared to tailings obtained by conventional methods (i.e., heavy fine flotation tailings in Comparative Example 1); after classification and separation by hydrocyclone, due to the increased concentration of tailings 1, the amount of slurry entering the thickening stage is reduced, which can reduce the amount of flocculant used in the thickening stage.

[0021] Therefore, this invention effectively solves the problem of mixed heavy and light materials caused by the large number of spiral chutes and fluctuations in factors such as flow rate and concentration, leading to a deterioration in subsequent flotation performance. It can effectively improve the overall flotation recovery rate and reduce the cost of flotation auxiliary materials. Attached Figure Description

[0022] Figure 1 Process flow diagram of this invention. Detailed Implementation

[0023] The embodiments of the present invention will be described in detail below with reference to the examples. The following examples are only used to illustrate the present invention and should not be regarded as limiting the scope of the present invention.

[0024] In the following examples and comparative examples, the reverse flotation reagents used, by mass fraction, consist of 35% dodecyl etheramine, 40% benzyl dimethyl tetradecyl ammonium chloride, and 25% trimethylchlorosilane.

[0025] Example 1 Phosphate ore with a P2O5 grade of 25.5 wt% and an Al2O3 content of 3.9 wt% was ground to a particle size of -0.074 mm (65%). The slurry concentration was then adjusted to 20%, and the width of the concentrate strips in the spiral sluice unit was adjusted to 12 cm. The slurry was fed into the spiral sluice unit for separation, yielding a heavy concentrate with a yield of 62.3% and a grade of 27.1% and heavy tailings with a yield of 37.7% and a grade of 22.8%. The concentration of the heavy concentrate obtained from the separation was then controlled to 25%, and fed into the reverse flotation dealuminization process with a flotation reagent dosage of 300 g / t (based on dry ore quantity), yielding heavy concentrate and heavy concentrate flotation tailings. The grade of the heavy concentrate flotation tailings was 18.55%, and the particle size distribution is shown in Table 1-1. The heavy fine flotation tailings were fed into an FX150 hydrocyclone with an underflow diameter of 20 mm and a feed pressure controlled at 0.08 MPa for classification and separation, yielding tailings 1 and middlings. The particle sizes of tailings 1 and middlings are shown in Tables 1-2 and 1-3. Tailings 1 had a grade of 11.2%, and middlings had a grade of 20.3%. The middlings were then combined with the heavy tailings from the spiral sluice, and the flotation concentration was controlled at 12%. The mixture was then sent to the heavy tailings reverse flotation dealuminization process, using the same reagents as in the heavy fine flotation, at a dosage of 100 g / t (based on dry ore quantity), yielding heavy tailings concentrate and tailings 2.

[0026] The heavy concentrate obtained in this embodiment has a P2O5 grade of 30.9 wt% and an Al2O3 content of 1.78%; the heavy tailings concentrate has a P2O5 grade of 30.1 wt% and an Al2O3 content of 1.91%; and the overall P2O5 recovery rate is 75.37%.

[0027] Table 1-1 Grain size analysis of heavy flotation tailings

[0028] Table 1-2 Grain Size Analysis of Graded and Separated Tailings

[0029] Table 1-3 Grain Size Analysis of Minerals Based on Classification and Quality

[0030] In the table, +0.15mm represents a particle size greater than 0.15mm, and -0.023mm represents a particle size less than 0.023mm.

[0031] Example 2 Phosphate ore with a P2O5 grade of 23.3 wt% and an Al2O3 content of 4.5 wt% was ground to a particle size of -0.074 mm (65%). The slurry concentration was then adjusted to 20%, and the width of the concentrate strips in the spiral sluice unit was adjusted to 11 cm. The slurry was fed into the spiral sluice unit for separation, yielding a heavy concentrate with a yield of 58.8% and a grade of 25.8% and heavy tailings with a yield of 41.2% and a grade of 19.9%. The concentration of the heavy concentrate obtained from the separation was then controlled to 25%, and fed into the reverse flotation dealuminization process with a flotation reagent dosage of 350 g / t (based on dry ore quantity), yielding heavy concentrate and heavy concentrate flotation tailings. The tailings grade was 17.23%, and the particle size distribution is shown in Table 2-1. The tailings were fed into an FX150 hydrocyclone with an underflow diameter of 20 mm and a feed pressure controlled at 0.08 MPa for classification and separation, yielding tailings 1 and middlings. The particle sizes of tailings 1 and middlings are shown in Tables 2-2 and 2-3. Tailings 1 had a grade of 10.9%, and middlings had a grade of 19.5%. The middlings were then combined with the heavy tailings from the spiral sluice, and the flotation concentration was controlled at 12%. The same reagents as those used in the heavy cleaning flotation were used at a dosage of 150 g / t, and the mixture was then sent to the heavy tailings reverse flotation dealuminization process to obtain heavy tailings concentrate and tailings 2.

[0032] The heavy concentrate obtained in this embodiment has a P2O5 grade of 30.1 wt% and an Al2O3 content of 1.86%; the heavy tailings concentrate has a P2O5 grade of 29.8 wt% and an Al2O3 content of 1.98%; the overall P2O5 recovery rate is 74.68%. Compared with Example 1, due to the lower grade of the raw ore, the amount of reagent used has increased, but a good separation effect can still be achieved.

[0033] Table 2-1 Grain size analysis of gravity flotation tailings

[0034] Table 2-2 Grading and Sorting of Tailings 1. Particle Size Screening

[0035] Table 2-3 Grain Size Screening of Minerals Based on Grading and Quality

[0036] Example 3 Phosphate ore with a P2O5 grade of 25.5 wt% and an Al2O3 content of 3.9 wt% was ground to a particle size of -0.074 mm (75% of the total particle size). The slurry concentration was then adjusted to 20%, and the width of the concentrate strips collected in the spiral sluice unit was adjusted to 12 cm. The slurry was fed into the spiral sluice unit for separation, yielding a heavy concentrate with a yield of 56.3% and a grade of 26.8%, and heavy tailings with a yield of 43.7% and a grade of 23.8%. The separated heavy concentrate was then fed into the flotation dealuminization process. The heavy concentrate obtained from the separation was then controlled to a concentration of 25% and fed into the reverse flotation dealuminization process. The flotation reagent dosage was 300 g / t (based on dry ore quantity), yielding heavy concentrate and heavy concentrate flotation tailings with a grade of 19.04%. The particle size distribution is shown in Table 3-1. The tailings were then fed into an FX150 hydrocyclone with a bottom outlet diameter of 20 mm and a feed pressure controlled at 0.06 MPa for classification and separation, yielding tailings 1 and middlings. The particle sizes of tailings 1 and middlings are shown in Tables 3-2 and 3-3. Tailings 1 had a grade of 11.5%, and middlings had a grade of 21.3%. The middlings were then combined with the heavy tailings from the spiral sluice, with the flotation concentration controlled at 12%. The same reagents as those used in the heavy concentrate flotation were used at a dosage of 100 g / t, and the mixture was fed into the heavy tailings reverse flotation dealuminization process, yielding heavy tailings concentrate and tailings 2.

[0037] The heavy concentrate obtained in this embodiment has a P2O5 grade of 30.6 wt% and an Al2O3 content of 1.8%; the heavy tailings concentrate has a P2O5 grade of 30.3 wt% and an Al2O3 content of 1.88%; the overall P2O5 recovery rate is 74.97%. Compared with Example 1, it can be seen that due to the increased fineness of the raw ore, under the condition that the operating parameters of the spiral sluice remain unchanged, the heavy concentrate yield decreases by about 6%, and the grade also decreases. The fine particle content of the heavy concentrate flotation tailings also increases. The mixture of coarse and fine particles will have an adverse effect on subsequent scavenging. By controlling the operating parameters of the hydrocyclone, the heavy concentrate flotation tailings can be classified and separated, and the high-grade, fine-particle phosphate ore can be recovered and entered into the heavy tailings reverse flotation dealuminization flotation process, which can effectively improve the phosphate ore recovery rate and reduce the phosphate tailings grade. In the production environment, parameters such as grinding fineness often fluctuate, and the spiral sluice gravity separation is very sensitive to these fluctuations. By increasing the classification and separation, the requirements of the spiral sluice for these parameters can be reduced, and the separation effect can be improved.

[0038] Table 3-1 Grain size sieve analysis of gravity flotation tailings

[0039] Table 3-2 Grading and Sorting of Tailings 1 Particle Size Screening

[0040] Table 3-3 Grading and Quality Analysis of Medium Ore Particle Size Screening

[0041] Comparative Example 1 This example uses the same process as Example 1, except that the heavy concentrate flotation tailings classification and quality separation process is not added. The heavy concentrate is fed into the reverse flotation dealuminization process, with the concentration controlled at 25% and the reagent dosage at 300g / t (based on dry ore). An additional scavenging operation is added, without adding reagents or adjusting the flotation concentration, to obtain heavy concentrate and heavy concentrate flotation tailings, with a heavy concentrate flotation tailings grade of 14.11%. The heavy tailings are fed into the flotation process separately, with the flotation concentration controlled at 12%, and then enter the heavy tailings reverse flotation dealuminization process, increasing the reagent dosage to 150g / t to obtain heavy tailings concentrate.

[0042] The heavy concentrate obtained in this embodiment has a P2O5 grade of 30.5 wt% and an Al2O3 content of 1.81%; the heavy tailings concentrate has a P2O5 grade of 29.4 wt% and an Al2O3 content of 1.97%; the overall P2O5 recovery rate is 68.44%. Compared with Example 1, this example did not involve classification and separation. To reduce the tailings grade during heavy concentrate flotation, conventional methods require adding scavenging, which necessitates additional flotation machines, increases energy consumption, and raises costs. Furthermore, due to the coarse inclusions in the heavy concentrate flotation tailings, the scavenging separation effect is poor, making it difficult to reduce the flotation tailings grade, resulting in ineffective recovery of useful phosphorus and a low overall recovery rate. In Example 1, due to the addition of classification and separation, some phosphate rock with a high reagent content was fed into the heavy tailings reverse flotation dealuminization process, significantly reducing the amount of reagent used for heavy tailings reverse flotation dealuminization, reducing the consumption of production auxiliary materials, and saving costs.

Claims

1. A method for combining gravity flotation and separation of collophane ore with intermediate classification and quality-differentiated flotation, characterized in that, Includes the following steps: S1. The phosphate slurry is fed into a spiral chute group for gravity separation to obtain heavy concentrate and heavy tailings. S2. The heavy concentrate is fed into the flotation dealuminization process to obtain heavy concentrate and heavy concentrate flotation tailings; S3. Feed the heavy flotation tailings into the classification and sorting process to obtain tailings 1 and middlings; S4. Combine the middlings and heavy tailings and proceed to the heavy tailings flotation dealuminization process to obtain heavy tailings concentrate and tailings 2.

2. The method for combined gravity flotation and separation of collophane ore with intermediate classification and quality separation as described in claim 1, characterized in that, The phosphate rock slurry has a concentration of 15-35%, a P2O5 grade of 18.0-27.5 wt%, an Al2O3 content of 2.5-5.5 wt%, and a particle size of less than 0.074 mm accounting for 45-85%.

3. The method for combined gravity flotation and separation of collophane ore with intermediate classification and quality separation as described in claim 1, characterized in that, In the flotation dealuminization process described in S2, the added dealuminizing collector is one or more of dodecyl ether amine, benzyl dimethyl tetradecyl ammonium chloride, or trimethylchlorosilane; preferably, it comprises the following components by mass percentage: 25-45% dodecyl ether amine, 30-50% benzyl dimethyl tetradecyl ammonium chloride, and 20-35% trimethylchlorosilane.

4. The method for combined gravity flotation and separation of collophane ore with intermediate classification and quality separation as described in claim 1, characterized in that, In the flotation dealuminization process described in S2, the flotation reagent dosage is 200-400 g / t.

5. The method for combined gravity flotation and separation of collophane ore with intermediate classification and quality separation as described in claim 1, characterized in that, The flotation dealumination process described in S2 is reverse flotation dealumination, with the flotation concentration for heavy fine dealumination controlled at 20-30%.

6. The method for combined gravity flotation and separation of collophane ore with intermediate classification and quality separation according to claim 1, characterized in that, The flotation dealuminization process described in S4 is reverse flotation dealuminization, with the flotation concentration for heavy tail dealuminization controlled at 5-15%.

7. The method for combined gravity flotation and separation of collophane ore with intermediate classification and quality separation as described in claim 1, characterized in that, The equipment used in the grading and sorting process is a high-frequency screen or a hydrocyclone.

8. The method for combined gravity flotation and separation of phosphate rock according to claim 1, characterized in that, The grading and sorting process ensures that the proportion of particles with a diameter less than 0.023 mm in the tailings 1 after heavy flotation is ≤20%.

9. The method for combined gravity flotation and separation of collophane ore with intermediate classification and quality separation according to claim 1, characterized in that, In the flotation dealuminization process described in S4, the added dealuminizing collector is one or more of dodecyl ether amine, benzyl dimethyl tetradecyl ammonium chloride, or trimethylchlorosilane; preferably, it includes the following components by mass percentage: 25-45% dodecyl ether amine, 30-50% benzyl dimethyl tetradecyl ammonium chloride, and 20-35% trimethylchlorosilane.

10. The method for combined gravity flotation and separation of collophane ore with intermediate classification and quality separation according to claim 1, characterized in that, In the flotation dealuminization process described in S4, the flotation reagent dosage is 50-200 g / t.

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