A collophanite reverse flotation process applied to yield improvement

Abstract

The application discloses a collophanite reverse flotation process applied to yield improvement, and the process is characterized in that flotation reagents are sequentially added into phosphorite slurry and then introduced into roughing operation, roughing is carried out, and rough concentrate and foam product I are obtained; the foam product I is introduced into one-stage scavenging after adding an adjusting agent, and scavenging concentrate I and foam product II are obtained; the scavenging concentrate is reselected after adding flotation reagents, and reselected concentrate and foam product III are obtained; the reselected concentrate and the rough concentrate are mixed and then introduced into cleaning after adding flotation reagents, and final concentrate and foam product IV are obtained, and the foam product IV is returned to the roughing operation; the foam product III is mixed with the foam product II, and two-stage scavenging is carried out, and scavenging concentrate II and foam product V are obtained, the foam product V is final tailings, and the scavenging concentrate II is returned to the one-stage scavenging operation. The process is stable, simple and efficient, has strong adaptability to ores, and has low phosphorus content of tailings and higher phosphorus yield.

Description

Technical Field

[0001] This invention belongs to the field of collophane beneficiation technology, specifically relating to a reverse flotation process for collophane used to improve yield. Background Technology

[0002] my country has abundant phosphate rock reserves, but due to the low reserves of high-grade phosphate rock, it suffers from the problem of "abundance without richness." Medium- and low-grade collophane accounts for over 90% of the country's phosphate rock resources, and this type of collophane has a high impurity content, requiring enrichment before it can be utilized at a high value. Currently, the mainstream enrichment method is flotation, with single-reverse flotation being the most widely used process. Its significant advantages are good process stability and low operating costs, but it suffers from significant phosphorus loss, and with the gradual deterioration of phosphate rock quality, it can no longer meet the needs of downstream production.

[0003] Chinese patent CN113182077A discloses a reverse flotation process for simultaneous removal of magnesium and aluminum from phosphate rock. This process employs a phosphate slurry-based simultaneous reverse flotation flow, where modifiers, fatty acid collectors, and amine collectors are added to the phosphate slurry for pre-flotation to obtain in-cell product I and froth product I. In-cell product I is then roughed to obtain in-cell product II and froth product II, where in-cell product II is the phosphate concentrate. Froth product II undergoes two scavenging processes to obtain phosphate tailings. This process is characterized by strong ore adaptability, low flotation reagent consumption, and low beneficiation cost. However, it suffers from significant phosphorus loss and the instability of the process due to foam stickiness during operation.

[0004] Chinese patent CN115283132A discloses a flotation method for low-grade mixed phosphate rock. This method involves adding sulfuric acid and phosphoric acid as pH adjusters and YP6-6 as a collector to the slurry. After one roughing, one cleaning, and one scavenging stage, cleaned and scavenged concentrates are obtained. The scavenged concentrate is then subjected to a second flotation to obtain a re-selected concentrate. The re-selected concentrate and cleaned concentrate are combined and subjected to cyclone classification and the addition of reverse flotation desilication reagents to obtain a desilication concentrate. By increasing the supply of cyclone classification, the useful minerals in the tailings are further extracted. Subsequently, combined with regrinding and re-selection in a frame mill, the P2O5 grade in the tailings can be reduced by more than 2%, and the phosphate concentrate recovery rate can be increased by 3%-5%. However, this method suffers from problems such as an excessively long process and high secondary grinding costs.

[0005] Chinese patent CN103949350A discloses a fine screening, regrinding, and classification flotation method for low-grade siliceous calcium phosphate ore. This method employs a combination of classification flotation and regrinding / re-concentration, where the raw ore pulp is screened and then subjected to separate flotation processes. Fine screening achieves narrow-grade material separation, significantly reducing the impact of fine particles on the flotation of coarse particles and improving the overall product recovery rate. However, the flotation process is too long and involves secondary grinding, resulting in high costs.

[0006] In his journal article "Re-concentration of Middlings from Dayukou Phosphate Ore" published in the Chemical Mining Technology Journal, Jiang Quanwei disclosed a middlings re-concentration process. By merging the tailings and scavenging concentrate as middlings for re-concentration, the final concentrate yield was increased by about 4%. However, the operation of returning the re-concentrated concentrate to the rougher caused the rougher concentration, fineness, and grade to be affected, with large fluctuations, which affected the stability of the process.

[0007] In his journal article "Experimental Study on Improving the Recovery Rate of Low-Grade Collophane Ore by Re-grinding Middlings" published in the Journal of Chemical Minerals and Processing, Jiang Zhensheng announced a flotation process of scavenging middlings and then re-grinding them. The concentrate yield was increased by 3.42% compared with the original process. However, the re-grinding process increased the grinding pressure of the ball mill, reduced the grinding capacity, and decreased the processing capacity.

[0008] Therefore, existing technologies can only choose one of high phosphorus yield, short processing flow, and low cost, which seriously restricts the development and application of phosphate rock flotation technology, and also has great limitations in the comprehensive utilization of phosphate rock. Summary of the Invention

[0009] To address the aforementioned technical problems, this invention provides a reverse flotation process for collophane ore to improve yield. This process is simple and efficient, adaptable to various ores, produces excellent concentrate indicators, has low phosphorus content in tailings, and achieves a higher phosphorus yield.

[0010] To achieve the above objectives, the present invention provides a reverse flotation process for collophane ore to improve yield, comprising the following steps:

[0011] (1) Grind the phosphate rock and add water to prepare phosphate rock slurry. Then, add modifier, inhibitor and fatty acid collector in sequence to adjust the slurry. After roughing, obtain rough concentrate and froth product I.

[0012] (2) Mix foam product I with a modifier and then perform a first-stage scavenging process to obtain scavenged concentrate I and foam product II;

[0013] (3) Add modifier, inhibitor and fatty acid collector to scavenging concentrate I in sequence, mix well and then re-select to obtain re-selected concentrate and foam product III;

[0014] (4) Mix the crude concentrate obtained in step (1) and the re-selected concentrate obtained in step (3), and add high-efficiency flocculant, fatty acid collector and amine collector in sequence. After fine selection, concentrate and foam product IV are obtained.

[0015] (5) Return the foam product IV to step (1) for coarse selection;

[0016] (6) After mixing the foam product II obtained in step (2) with the foam product III obtained in step (3), the foam product V is obtained by two-stage scavenging. The foam product V is the tailings.

[0017] (7) Return the scavenged concentrate II to step (2) for a scavenging process.

[0018] Preferably, the phosphate rock slurry in step (1) has a mass concentration of 30-40%, a P2O5 grade of 24-28%, a MgO grade of 2.5-4.0%, an Al2O3 grade of 2.0-2.5%, and a -0.074mm content of 55-80%.

[0019] Preferably, the modifiers mentioned in steps (1)-(3) are composed of sulfuric acid (98%) and phosphoric acid (85%) in a mass ratio of 1-7:9-3.

[0020] Preferably, the inhibitors in steps (1) and (3) are any one or more of citric acid, oxalic acid, sodium hypophosphite, sodium pyrophosphate, and sodium polyphosphate.

[0021] Preferably, the fatty acid collector in steps (1) and (3) is a fatty acid salt obtained by saponification of a mixture of palmitic acid, stearic acid, oleic acid, linoleic acid and palmitic acid in a mass ratio of 15-20:5-10:15-25:25-35:5-15.

[0022] Preferably, the high-efficiency flocculant in step (4) is any one or more of polyaluminum chloride, polyferric sulfate, alum, chitosan, high molecular weight polyacrylamide, low molecular weight polyacrylamide, polyferric silicate, and polyaluminum silicate.

[0023] Preferably, the fatty acid collector in step (4) is a mixed fatty acid salt obtained by saponification of palmitic acid, stearic acid, oleic acid, linoleic acid and palmitic acid in a mass ratio of 15-20:5-10:15-25:25-35:5-15.

[0024] Preferably, the amine collector in step (4) is dodecylamine hydrochloride.

[0025] Preferably, the dosage of the modifier in step (1) is 5-10 kg / t; the dosage of the inhibitor is 1-3 kg / t; and the dosage of the fatty acid collector is 0.8-1.5 kg / t.

[0026] The dosage of the modifier mentioned in step (2) is 0.5-1.5 kg / t;

[0027] The dosage of the modifier in step (3) is 2.5-4.5 kg / t, the dosage of the inhibitor is 1-3 kg / t, and the dosage of the fatty acid collector is 0.1-0.5 kg / t;

[0028] The dosage of the high-efficiency flocculant in step (4) is 0.05-0.2 kg / t, the dosage of the fatty acid collector is 0.5-1.0 kg / t, and the dosage of the amine collector is 0.05-0.2 kg / t.

[0029] The beneficial effects of this invention are as follows:

[0030] 1. This invention re-selects the scavenged concentrate I obtained after a first-stage scavenging process, discarding it and returning it to the roughing process. This avoids the influence of the scavenged concentrate I on the grade and concentration of the pulp entering the roughing operation, and effectively improves the separation effect of gangue minerals and useful minerals in the roughing operation.

[0031] 2. The obtained scavenging concentrate I is further processed to improve the phosphorus grade of the further processed concentrate, while reducing the phosphorus grade in the tailings; the further processed concentrate is then finely processed to further improve the quality of the phosphate rock and ensure that the concentrate indicators meet the requirements.

[0032] 3. Return the foam product IV to the roughing process for further separation. Without affecting the phosphorus grade and concentration in the roughing pulp, fully recover the phosphate rock, improve the phosphate rock recovery rate, and reduce losses.

[0033] 4. By adding an inhibitor composed of any one or more of citric acid, oxalic acid, sodium hypophosphite, sodium pyrophosphate, and sodium polyphosphate to the slurry and scavenging concentrate I, the separation effect can be significantly improved and the phosphorus content of the tailings can be further reduced.

[0034] 5. The fine separation process, through the combination of high-efficiency flocculants, fatty acid collectors, and amine collectors, can further reduce carbonate gangue minerals in the concentrate while removing some silicate gangue minerals, thus further improving the quality of the phosphate concentrate. This is mainly because the roughing and re-separation processes remove some fine-grained gangue minerals, mitigating the problem of strong foam toughness and poor flowability caused by the binding of amine collectors with fine-grained mineral particles. Furthermore, the high-efficiency flocculant can partially achieve selective flocculation of fine-grained silicate gangue minerals, further enhancing the foaming effect of the amine collector while strengthening the separation effect, thereby removing silicate gangue minerals and improving the ore's applicability. Attached Figure Description

[0035] Figure 1 This is a process flow diagram of the present invention.

[0036] Figure 2 This is a process flow diagram for Example 9. Detailed Implementation

[0037] The technical solution of the present invention will be further explained and described below with reference to the accompanying drawings and specific embodiments. It is worth noting that the following embodiments are only preferred embodiments of the present invention and should not be construed as limiting the present invention. The scope of protection of the present invention should be determined by the contents of the claims. Modifications and substitutions made by those skilled in the art to the technical solution of the present invention without creative effort all fall within the scope of protection of the present invention.

[0038] Medium- to low-grade collophane: P2O5 grade is 26%, MgO grade is 3.54%, and Al2O3 grade is 2.35%.

[0039] Example 1

[0040] (1) Take medium-low grade phosphate rock (ore processing capacity is 140t / h) and prepare phosphate rock slurry with a mass concentration of 30% by ball milling;

[0041] (2) Sulfuric acid (98%) and phosphoric acid (85%) were mixed in a mass ratio of 1:1 to prepare a modifier; palmitic acid, stearic acid, oleic acid, linoleic acid and palmitic acid were mixed in a mass ratio of 15:10:15:25:5 and subjected to saponification to obtain a mixed fatty acid sodium salt, which was used as a fatty acid collector; oxalic acid was used as an inhibitor.

[0042] (3) Add 6 kg / t of modifier, 1.5 kg / t of inhibitor and 1.2 kg / t of fatty acid collector to the phosphate rock slurry prepared in step (1) in sequence, and obtain rough concentrate and foam product I after roughing.

[0043] (4) Add 1.0 kg / t of modifier to the foam product I obtained in step (3), mix evenly, and then perform a first-stage scavenging to obtain scavenged concentrate I and foam product II;

[0044] (5) Add 3 kg / t of modifier, 1.5 kg / t of inhibitor and 0.4 kg / t of fatty acid collector to the scavenging concentrate I obtained in step (4) in sequence. After mixing, carry out the re-selection operation to obtain the re-selected concentrate and foam product III.

[0045] (6) After mixing the reconstituted concentrate obtained in step (5) with the rough concentrate obtained in step (3), add 0.1 kg / t of high-efficiency flocculant, 0.8 kg / t of fatty acid collector and 0.12 kg / t of amine collector in sequence. After mixing, perform fine selection to obtain concentrate and foam product IV. The high-efficiency flocculant is polyaluminum chloride, the amine collector is dodecylamine hydrochloride, and the foam product IV is returned to step (1) for rough selection.

[0046] (7) After mixing the foam product II obtained in step (4) with the foam product III obtained in step (5), a two-stage scavenging operation is performed to obtain scavenging concentrate II and foam product V. Foam product V is the final tailings, and scavenging concentrate II is returned to the first-stage scavenging operation.

[0047] Tests revealed that the concentrate obtained in step (6) had a P2O5 grade of 32.14%, an MgO grade of 0.78%, an Al2O3 grade of 1.98%, and a yield of 78.33%; the tailings obtained in step (7) had a phosphorus grade of 3.8%; and the phosphate rock recovery rate was 96.83%.

[0048] Example 2

[0049] The method and steps are the same as in Example 1, except that step (5) is canceled. The scavenged concentrate I obtained in step (4) is directly returned to step (3) for roughing again to obtain concentrate and tailings respectively.

[0050] Testing revealed that the concentrate contained 32.25% P2O5, 0.75% MgO, and 1.94% Al2O3, with a yield of 73.35%. The tailings contained 8.8% phosphorus, resulting in a phosphate rock recovery rate of 90.98%. Compared to Example 1, this example eliminated the re-selection process, leading to a significant decrease in phosphorus recovery from the resulting concentrate.

[0051] Example 3

[0052] The method and steps are the same as in Example 1, except that the re-selected concentrate obtained in step (5) is returned to step (3) for roughing operation to obtain concentrate and tailings respectively.

[0053] Testing revealed that the concentrate had a P2O5 grade of 32.65%, an MgO grade of 0.73%, an Al2O3 grade of 1.95%, and a yield of 75.21%. The tailings had a phosphorus grade of 5.83%, and the phosphate rock recovery rate was 94.44%. While the concentrate indicators were slightly better than in Example 1, the yield was lower, and the process stability was poor. The roughing operation was affected by the re-concentration concentrate pulp, resulting in significant fluctuations in the feed pulp concentration and phosphorus grade, leading to large variations in the roughing separation effect and making it impossible to obtain stable phosphate concentrate indicators.

[0054] Table 1. Comparison Results of Process Stability

[0055]

[0056] The comparison results show that although Example 3 can obtain qualified phosphate concentrate products, the yield fluctuates greatly.

[0057] Example 4

[0058] The method and steps are the same as in Example 1, except that the foam product III obtained from the re-selection operation in step (5) is returned to step (4) for another first-stage scavenging operation, instead of the second-stage scavenging operation in step (7), to obtain concentrate and tailings respectively.

[0059] Testing revealed that the concentrate had a P2O5 grade of 31.57%, an MgO grade of 0.88%, an Al2O3 grade of 2.11%, and a yield of 79.72%. The tailings had a phosphorus grade of 4.1%, and the phosphate rock recovery rate was 96.80%. While the concentrate yields were essentially the same, the indicators differed from Example 1. This was mainly because the first-stage scavenging operation was affected by the return of the froth product III from the re-selection process, leading to a decrease in the indicators of the scavenging concentrate I. This, in turn, affected the indicators of the re-selection concentrate, ultimately resulting in increased refining pressure and a further decrease in concentrate indicators.

[0060] Table 2 Comparison Results of Process Stability

[0061]

[0062]

[0063] Example 5

[0064] The method and steps are the same as in Example 1, except that the fine selection operation in step (6) is cancelled. The re-selected concentrate and the rough concentrate are mixed and used as the final concentrate to obtain concentrate and tailings respectively.

[0065] Testing revealed that the concentrate had a P2O5 grade of 31.35%, an MgO grade of 0.90%, and an Al2O3 grade of 2.41%, with a yield of 80.69%. The tailings had a phosphorus grade of 3.65%, and the phosphate rock recovery rate was 97.29%. Compared to Example 1, this example yielded a higher concentrate, but the concentrate had higher MgO and Al2O3 content and lower P2O5 content, resulting in unqualified concentrate indicators.

[0066] Example 6

[0067] The method and steps are the same as in Example 1. The foam product IV obtained from the fine selection operation in step (6) is returned to step (4) for a scavenging operation to obtain concentrate and tailings respectively.

[0068] Analysis revealed that the concentrate contained 32.59% P2O5, 0.71% MgO, and 1.95% Al2O3, with a yield of 74.55%. The tailings contained 6.7% phosphorus, and the phosphate rock recovery rate was 93.44%. Both the concentrate yield and phosphate rock recovery rate showed significant decreases, indicating substantial phosphorus loss. Therefore, this process is not suitable for reverse flotation of collophane.

[0069] Example 7

[0070] The method and steps are the same as in Example 1. In step (6), no high-efficiency flocculant is added in the fine selection operation, but only fatty acid collectors and amine collectors are added to obtain concentrate and tailings respectively.

[0071] Testing revealed that the concentrate contained 31.98% P2O5, 0.75% MgO, and 2.18% Al2O3, with a yield of 78.00%. The tailings contained 4.8% phosphorus, resulting in a phosphate rock recovery rate of 95.94%. Compared to Example 1, the concentrate had a higher Al2O3 content. Furthermore, practical application revealed that the selected foam product IV exhibited sticky foam with poor flowability, hindering normal process operation.

[0072] Example 8

[0073] The method and steps are the same as in Example 1. In step (6), no high-efficiency flocculant and amine collector are added, only fatty acid collector is added to obtain concentrate and tailings respectively.

[0074] Testing revealed that the concentrate contained 32.01% P2O5, 0.68% MgO, and 2.43% Al2O3, with a yield of 79.14%. The tailings contained 3.2% phosphorus, resulting in a phosphate rock recovery rate of 97.43%. Compared to Example 1, the concentrate yield was higher, but the Al2O3 content was also higher, leading to a lower concentrate quality.

[0075] Example 9

[0076] The method and steps are the same as in Example 1, except that step (5) is changed to a three-segment sweep selection, as follows:

[0077] (1)-(4): Same as in Example 1, wherein the crude concentrate obtained in step (3) is the final concentrate;

[0078] (5): Return the scavenged concentrate I obtained in step (4) to step (3) for roughing;

[0079] (6) Perform two-stage scavenging on foam product II to obtain scavenged concentrate II and foam product IV. Return scavenged concentrate II to step (5) for another one-stage scavenging.

[0080] (7) Perform three-stage scavenging on foam product IV to obtain scavenged concentrate III and foam product V, and return scavenged concentrate III to step (6) for two-stage scavenging. The resulting foam product V is the final tailings.

[0081] Testing revealed that the concentrate had a P2O5 grade of 32.00%, an MgO grade of 0.79%, an Al2O3 grade of 1.99%, and a yield of 78.42%. The tailings had a phosphorus grade of 4.2%, and the phosphate rock recovery rate was 96.52%. The concentrate yield was largely the same as in Example 1. However, returning scavenged concentrate I to the roughing operation resulted in significant fluctuations in the concentration and grade of the roughing pulp, and the equipment could not operate at its maximum efficiency. Although the raw ore processing capacity could be increased to 160 t / h, the concentrate yield decreased by approximately 2%, leading to greater losses.

[0082] Example 10

[0083] The method and steps are the same as in Example 1, except that the ore processing capacity is increased to 160 t / h, and concentrate and tailings are obtained respectively.

[0084] Testing revealed that the concentrate had a P2O5 grade of 31.93%, an MgO grade of 0.80%, an Al2O3 grade of 2.00%, a yield of 78.69%, and a phosphorus grade of 4.1% in the tailings, resulting in a phosphate rock recovery rate of 96.64%. Comparison showed no significant differences in the concentrate indicators, tailings phosphorus grade, and phosphate rock recovery rate compared to Example 1. This indicates that, compared to Example 9, the process described in Example 1, with unchanged equipment, can effectively increase the raw ore processing capacity and maximize the processing power of the equipment.

[0085] Example 11

[0086] The method and steps are the same as in Example 1, except that the inhibitor in step (2) is replaced by sodium polyphosphate instead of oxalic acid, with a dosage of 1 kg / t, to obtain concentrate and tailings respectively.

[0087] Testing revealed that the concentrate contained 32.01% P2O5, 0.79% MgO, and 1.97% Al2O3, with a yield of 77.33%. The tailings contained 5.5% phosphorus, resulting in a phosphate rock recovery rate of 95.20%. Sodium polyphosphate alone was less effective at inhibiting phosphate rock growth than oxalic acid.

[0088] Example 12

[0089] The method and steps are the same as in Example 1, except that the inhibitor in step (2) is replaced by oxalic acid with a mixture of citric acid, oxalic acid and sodium polyphosphate in a mass ratio of 1:3:2, and the dosage is 1.5 kg / t, to obtain concentrate and tailings respectively.

[0090] Testing revealed that the concentrate had a P2O5 grade of 32.08%, an MgO grade of 0.80%, an Al2O3 grade of 1.98%, and a yield of 78.61%. The tailings had a phosphorus grade of 3.65%, and the phosphate rock recovery rate was 97.00%. The compound inhibitor had a stronger inhibitory effect and a lower phosphorus content in the tailings.

[0091] Example 13

[0092] The method and steps are the same as in Example 1, except that the inhibitor in step (2) is replaced by a mixture of oxalic acid, sodium hypophosphite and sodium pyrophosphate in a mass ratio of 3:2:1, with a dosage of 2 kg / t, to obtain concentrate and tailings respectively.

[0093] Testing revealed that the concentrate contained 32.06% P2O5, 0.8% MgO, and 1.99% Al2O3, with a yield of 78.17%. The tailings contained 4.3% phosphorus, and the phosphate rock recovery rate was 96.39%. No significant improvement in the inhibition effect was observed.

[0094] Example 14

[0095] The method and steps are the same as in Example 1, except that the high-efficiency flocculant in step (6) is replaced with low molecular weight polyacrylamide (molecular weight 6 million), and the dosage is 0.05 kg / t, to obtain concentrate and tailings respectively.

[0096] Testing revealed that the concentrate had a P2O5 grade of 32.00%, an MgO grade of 0.78%, an Al2O3 grade of 2.11%, and a yield of 78.49%. The tailings had a phosphorus grade of 4.10%, and the phosphate rock recovery rate was 96.61%. The foam was sticky and difficult to break, and there was serious overflow in the trough.

[0097] Example 15

[0098] The method and steps are the same as in Example 1, except that the high-efficiency flocculant in step (6) is replaced with a mixture of polyaluminum chloride, chitosan, polysilicon aluminum and low molecular weight polyacrylamide in a mass ratio of 2:3:2:3, and the dosage is 0.05 kg / t, to obtain concentrate and tailings respectively.

[0099] Testing revealed that the concentrate had a P2O5 grade of 32.20%, an MgO grade of 0.77%, an Al2O3 grade of 1.96%, and a yield of 78.20%. The tailings had a phosphorus grade of 3.76%, and the phosphate rock recovery rate was 96.85%. The concentrate indicators were slightly better than those in Example 1, and the foam was more brittle, with no overflow observed.

[0100] Example 16

[0101] The method and steps are the same as in Example 1, except that the high-efficiency flocculant in step (6) is replaced with a mixture of polyaluminum chloride, polyferric sulfate, alum, and high molecular weight polyacrylamide (molecular weight 1300) in a mass ratio of 3:2:2:3, and the dosage is 0.05 kg / t, to obtain concentrate and tailings respectively.

[0102] Testing revealed that the concentrate had a P2O5 grade of 32.17%, an MgO grade of 0.78%, an Al2O3 grade of 1.98%, a yield of 78.08%, and a phosphorus grade of 4.02% in the tailings. The phosphate rock recovery rate was 96.61%, which is similar to that of Example 1.

[0103] Example 17

[0104] The method and steps are the same as in Example 1, except that the high-efficiency flocculant in step (6) is replaced with a mixture of polyferric sulfate, polyferric silicate and chitosan in a mass ratio of 2:1:2, and the dosage is 0.05 kg / t, to obtain concentrate and tailings respectively.

[0105] During the test, the foam became sticky and difficult to break, causing severe overflow and affecting the normal operation of the process.

Claims

1. A reverse flotation process for collophane ore to improve yield, characterized in that: Includes the following steps: (1) The phosphate rock is prepared by grinding and adding water to make phosphate rock slurry. Then, the modifier, inhibitor and fatty acid collector are added in sequence to adjust the slurry. The rough concentrate and froth product I are obtained by roughing. (2) Mix foam product I with a modifier and then perform a first-stage scavenging process to obtain scavenged concentrate I and foam product II; (3) Add modifier, inhibitor and fatty acid collector to scavenging concentrate I in sequence, mix well and then re-select to obtain re-selected concentrate and froth product III; (4) Mix the crude concentrate obtained in step (1) and the re-selected concentrate obtained in step (3), and add high-efficiency flocculant, fatty acid collector and amine collector in sequence. After fine selection, concentrate and foam product IV are obtained. (5) Return the foam product IV to step (1) for coarse selection; (6) After mixing the foam product II obtained in step (2) with the foam product III obtained in step (3), the foam product III is obtained by two-stage scavenging to obtain scavenged concentrate II and foam product V. Foam product V is the tailings. (7) Return the scavenged concentrate II to step (2) for a first-stage scavenging; The dosage of the high-efficiency flocculant in step (4) is 0.05-0.2 kg / t, the dosage of the fatty acid collector is 0.5-1.0 kg / t, and the dosage of the amine collector is 0.05-0.2 kg / t.

2. The reverse flotation process for improving the yield of collophane ore according to claim 1, characterized in that: The phosphate rock slurry in step (1) has a mass concentration of 30-40%, a P2O5 grade of 24-28%, a MgO grade of 2.5-4.0%, an Al2O3 grade of 2.0-2.5%, and a -0.074mm content of 55-80%.

3. The reverse flotation process for improving the yield of collophane ore according to claim 1, characterized in that: Step (1) (3) The modifier is composed of sulfuric acid and phosphoric acid in a mass ratio of 1-7:9-3.

4. The reverse flotation process for improving the yield of collophane ore according to claim 1, characterized in that: The inhibitors mentioned in steps (1) and (3) are any one or more of citric acid, oxalic acid, sodium hypophosphite, sodium pyrophosphate, and sodium polyphosphate.

5. The reverse flotation process for improving the yield of collophane ore according to claim 1, characterized in that: The fatty acid collectors mentioned in steps (1) and (3) are fatty acid salts obtained by saponification of a mixture of palmitic acid, stearic acid, oleic acid, linoleic acid, palmitic acid in a mass ratio of 15-20:5-10:15-25:25-35:5-15.

6. The reverse flotation process for improving the yield of phosphate rock according to claim 1, characterized in that: The high-efficiency flocculant mentioned in step (4) is any one or more of polyaluminum chloride, polyferric sulfate, alum, chitosan, high molecular weight polyacrylamide, low molecular weight polyacrylamide, polyferric silicate, and polyaluminum silicate.

7. The reverse flotation process for improving the yield of collophane ore according to claim 1, characterized in that: The fatty acid collector mentioned in step (4) is a mixed fatty acid salt obtained by saponification of palmitic acid, stearic acid, oleic acid, linoleic acid and palmitic acid in a mass ratio of 15-20:5-10:15-25:25-35:5-15.

8. The reverse flotation process for improving the yield of collophane ore according to claim 1, characterized in that: The amine collector in step (4) is dodecylamine hydrochloride.

9. The reverse flotation process for improving the yield of collophane ore according to claim 1, characterized in that: The dosage of the modifier mentioned in step (1) is 5-10 kg / t; the dosage of the inhibitor is 1-3 kg / t; and the dosage of the fatty acid collector is 0.8-1.5 kg / t. The dosage of the modifier mentioned in step (2) is 0.5-1.5 kg / t; The dosage of the modifier in step (3) is 2.5-4.5 kg / t, the dosage of the inhibitor is 1-3 kg / t, and the dosage of the fatty acid collector is 0.1-0.5 kg / t.

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