High-silica collophanite flotation separation reagent system and application thereof

Abstract

The invention discloses a flotation separation reagent system for high-silica collophanite and application of the flotation separation reagent system for the high-silica collophanite, and the flotation separation reagent system for the high-silica collophanite comprises an inhibitor and a collecting agent, wherein the inhibitor comprises 2-3 parts of sulfuric acid, 1-2 parts of phosphoric acid, 1-2 parts of hydroxyethylidene diphosphate and 2-3 parts of sodium hexametaphosphate; the collecting agent comprises 4-5 parts of alkyl amino salt, 1-2 parts of octadecyl phosphate, 1-2 parts of laurate and 3-4 parts of vegetable fatty acid salt. Mainly solves the technical problems of large reagent usage amount, large environmental pollution and high cost in the high-silicon collophanite beneficiation process at present, and in addition, the invention further provides application of the high-silicon collophanite flotation separation reagent system in high-silicon collophanite flotation separation.

Description

Technical Field

[0001] The invention relates to the technical field of collophanite beneficiation, in particular to a high-silica collophanite flotation separation reagent system and application thereof. Background Art

[0002] As a major chemical raw material, phosphate rock is widely used in agriculture, food, medicine and other fields. It is closely related to people's daily life. At the same time, it is a non-renewable and non-recyclable resource. my country is a large agricultural country, and phosphate fertilizers are indispensable in agricultural production. my country's phosphate resources are of low grade, with a high content of siliceous gangue minerals, and mineral separation is difficult. Most ores need to go through complex processing processes to obtain high-grade qualified concentrate products.

[0003] The reserves of high-siliceous collophosphate in my country account for nearly 30%. The content of siliceous gangue minerals in high-siliceous collophosphate is high, and the development and utilization of the ore is difficult. The current mineral processing technology is mainly based on direct and reverse flotation technology or double reverse flotation technology, which has a series of problems such as long process, complex medium, large reagent consumption, and difficult tail water treatment, resulting in high mineral processing costs, low mineral processing efficiency, and high environmental pollution. The improved mineral processing technology can reduce the amount of phosphate ore entering the floatation process and improve resource utilization through the pre-throwing process, but the product after the pre-throwing still needs to use double reverse flotation to remove magnesium and silicon, which is easy to cause problems such as high cost of mineral processing reagents.

[0004] In summary, the existing technology of flotation mixed type collophanite beneficiation has the problems of high cost and large amount of reagents.

[0005] Therefore, we urgently need a high-silica collophosphate flotation separation reagent system that has less environmental pollution, lower cost and less reagent usage. Summary of the invention

[0006] The purpose of the present invention is to overcome the shortcomings of the prior art and provide a flotation separation reagent system for high-silica collophanite and its application, which mainly solves the technical problems of large amount of reagent used, great environmental pollution and high cost in the current high-silica collophanite beneficiation process.

[0007] The objective of the present invention is achieved through the following technical solutions:

[0008] Firstly, the present invention provides a flotation separation reagent system for high-silica collophanite, including a depressant and a collector;

[0009] The inhibitor comprises 2-3 parts of sulfuric acid, 1-2 parts of phosphoric acid, 1-2 parts of hydroxyethylene diphosphoric acid and 2-3 parts of sodium hexametaphosphate; the collector comprises 4-5 parts of alkyl amino acid salt, 1-2 parts of octadecyl phosphate, 1-2 parts of laurate and 3-4 parts of vegetable oil acid salt.

[0010] Furthermore, the preparation method of the alkyl amino acid salt comprises the following steps:

[0011] Methyl linoleate is added to dodecyl primary amine, and the mixture is stirred and then subjected to a first heating reaction to obtain a reaction solution; and NaOH is added to the reaction solution and subjected to a second heating reaction to obtain the alkylamino salt.

[0012] Furthermore, the weight ratio of the dodecyl primary amine to methyl linoleate is 1:(3-4).

[0013] Furthermore, the weight ratio of the NaOH to the reaction solution is (0.1-0.2):1.

[0014] Furthermore, the temperature of the first heating reaction is 100 to 120° C., and the time of the first heating reaction is 40 to 50 minutes.

[0015] Furthermore, the temperature of the second heating reaction is 60-80° C., and the time of the second reaction is 3-5 hours.

[0016] Furthermore, the preparation method of the collector comprises the following steps:

[0017] Octadecyl phosphate, lauric acid and vegetable oil are mixed to obtain a mixed solution; NaOH is added to the mixed solution to perform a third heating reaction to obtain a primary collector; and the alkyl amino acid salt is added to the primary collector to obtain the collector.

[0018] Furthermore, the vegetable oil comprises cottonseed oil and linoleic acid, and the mass ratio of the cottonseed oil to the linoleic acid is (1.5-2.5):1.

[0019] Furthermore, the weight ratio of the NaOH to the mixed solution is (0.1-0.2):1.

[0020] Furthermore, the temperature of the third heating reaction is 60-80° C., and the time is 3-5 hours.

[0021] Secondly, the present invention also provides an application of the high-silica collophanite flotation separation reagent system in the flotation separation of high-silica collophanite, comprising the following steps:

[0022] The raw ore is crushed, screened, photoelectrically scavenged and heavily scavenged in sequence to obtain a pre-enriched concentrate, and water-milled ore is added to the pre-enriched concentrate to obtain a flotation pulp; 2000-3000 g / t of the depressant and 400-800 g / t of the collector are added to the flotation pulp, and a roughing concentrate and a roughing tailing are obtained through stirring and aeration; 400-600 g / t of the depressant and 40-80 g / t of the collector are added to the roughing concentrate, and a final phosphate concentrate and a concentrated middling are obtained through stirring and aeration; and 800-1200 g / t of the depressant and 150-250 g / t of the collector are added to the roughing tailing, and a scavenging concentrate and a final tailing are obtained through stirring and aeration.

[0023] Furthermore, the selected middlings and scavenged concentrates can be returned to step S2, respectively, and steps S2 to S4 can be repeated.

[0024] According to the above technical solution, since the alkylamino salt is a drug molecule containing amino and carboxyl amphoteric groups, the amino group is a cation and can form electrostatic adsorption with the negatively charged mineral surface under suitable pH conditions. The carboxyl group is a functional group of anionic collectors and can form chemical adsorption with positively charged metal ions on the mineral surface. Through the adsorption of anions and cations of the hydrophilic group on the mineral surface, the collector molecules form an effective hydrophobic layer on the mineral surface, thereby enhancing the collector's ability to collect dolomite and siliceous minerals in the slurry, achieving improved selectivity and collection performance, and ensuring effective separation of apatite minerals, dolomite and siliceous gangue minerals; compared with conventional collectors, the collector in the above technical solution also adds dodecyl phosphate containing a phosphate group, and the selectivity of the collector is improved through the selective adsorption of the phosphate group and the metal ions on the surface of the gangue mineral. Cottonseed acid and linoleic acid are rich in a large number of unsaturated hydrocarbon groups, have a large critical micelle concentration, active chemical properties, and better polarity. The saponified sodium cottonseed acid and sodium linoleic acid molecules are easier to combine with ions on the surface of the target mineral due to the action of double bonds to form a hydrophobic layer.

[0025] The collector and the inhibitor in the present application can cooperate with each other to jointly improve the sorting effect of apatite minerals, thereby strengthening the flotation of fine-grained siliceous minerals in the flotation slurry, while reducing the usage of the two, and the components in the collector and the inhibitor are easy to degrade and recycle, thereby reducing pollution to the environment.

[0026] The beneficial effects of the present invention are:

[0027] The invention increases the flotation of fine-grained siliceous minerals in the flotation pulp by adding alkylamino salts to the collector, so that the collector molecules can be more effectively adsorbed on the surfaces of calcium-magnesium and siliceous minerals, thereby achieving the effect of effectively improving the sorting property of apatite minerals. BRIEF DESCRIPTION OF THE DRAWINGS

[0028] Figure 1 The present invention is a flow chart of a ore dressing process for step-by-step silicon reduction and impurity removal of high-siliceous collophosphate ore. DETAILED DESCRIPTION

[0029] The technical solution of the present invention is further described in detail below in conjunction with the accompanying drawings, but the protection scope of the present invention is not limited to the following.

[0030] Example 1

[0031] Prepare the alkyl amino acid salts as follows:

[0032] S11. 3 parts of methyl linoleate were added dropwise to 1 part of dodecyl primary amine, while stirring, and heated to 100° C., and reacted for 40 minutes to obtain a reaction solution;

[0033] S12. Add 1 part of NaOH to 10 parts of the reaction solution, heat to 60°C and react for 5 hours to obtain sodium alkyl amino acid, i.e., alkyl amino acid salt.

[0034] Example 2

[0035] Prepare the alkyl amino acid salts as follows:

[0036] S11. 3.5 parts of methyl linoleate were added dropwise to 1 part of dodecyl primary amine while stirring, and heated to 110° C. and reacted for 45 minutes to obtain a reaction solution;

[0037] S12. Add 1.5 parts of NaOH to 10 parts of the reaction solution, heat to 70°C and react for 54 hours to obtain sodium alkyl amino acid, i.e., alkyl amino acid salt.

[0038] Example 3

[0039] Prepare the alkyl amino acid salts as follows:

[0040] S11. 4 parts of methyl linoleate were added dropwise to 1 part of dodecyl primary amine while stirring, and heated to 120° C. and reacted for 50 minutes to obtain a reaction solution;

[0041] S12. Add 2 parts of NaOH to 10 parts of the reaction solution, heat to 80°C and react for 3 hours to obtain sodium alkyl amino acid, i.e., alkyl amino acid salt.

[0042] Example 4

[0043] Prepare vegetable oils as follows:

[0044] S21. Mix 3 parts of cottonseed oil and 2 parts of coconut oil to obtain the vegetable oil.

[0045] Example 5

[0046] Prepare vegetable oils as follows:

[0047] S21. Mix 3 parts of cottonseed oil and 2 parts of coconut oil to obtain the vegetable oil.

[0048] Example 6

[0049] Prepare vegetable oils as follows:

[0050] S21. Mix 5 parts of cottonseed oil and 2 parts of coconut oil to obtain the vegetable oil.

[0051] Example 7

[0052] Prepare the collector as follows:

[0053] S31. After mixing 1 part of octadecyl phosphate, 1 part of lauric acid and 4 parts of the vegetable oil prepared in Example 4, a mixed solution was obtained;

[0054] S32. Add 1 part of NaOH to 10 parts of the mixed solution, heat to 60° C. and react for 5 hours to obtain a primary collector;

[0055] S33. Add 4 portions of the alkyl amino acid salt prepared in Example 1 into the primary collector and stir to obtain the collector.

[0056] Example 8

[0057] Prepare the collector as follows:

[0058] S31. After mixing 1 part of octadecyl phosphate, 1 part of lauric acid and 3 parts of the vegetable oil prepared in Example 5, a mixed solution was obtained;

[0059] S32. Add 1.5 parts of NaOH to 10 parts of the mixed solution, heat to 70° C. and react for 4 hours to obtain a primary collector;

[0060] S33. Add 5 parts of the alkyl amino acid salt prepared in Example 2 to the primary collector and stir to obtain the collector.

[0061] Example 9

[0062] Prepare the collector as follows:

[0063] S31. After mixing 1 part of octadecyl phosphate, 1 part of lauric acid and 4 parts of the vegetable oil prepared in Example 6, a mixed solution was obtained;

[0064] S32. Add 2 parts of NaOH to 10 parts of the mixed solution, heat to 80 ° C and react for 3 hours to obtain a primary collector;

[0065] S33. Add 4 portions of the alkyl amino acid salt prepared in Example 3 into the primary collector and stir to obtain the collector.

[0066] Example 10

[0067] A high-siliceous collophosphate ore, taken from a concentrator in Leibo area, with a P2O5 grade of 20%. Its concentrating process is as follows: Figure 1 As shown, the following steps are included:

[0068] S41. Crushing the raw ore to obtain crushed ore with a particle size of ≤30 mm;

[0069] S42. The crushed ore is screened to obtain ores of -30mm+15mm and -30mm-15mm particle sizes, respectively;

[0070] S43. Photoelectric separation process is used for -30mm+15mm particle size ore to obtain photoelectric concentrate and photoelectric tailings;

[0071] S44. The photoelectric tailings obtained are subjected to photoelectric scavenging separation to obtain photoelectric scavenging concentrate and scavenging tailings respectively. The scavenging concentrate is returned to S43 operation, and the photoelectric scavenging tailings with P2O5 grade <10% are used as final tailings.

[0072] S45. The photoelectric concentrate obtained in S43 is further crushed and merged into the -15mm particle size obtained in S42. The obtained mixed -15mm particle size is screened and classified to obtain -15mm+0.5mm and -15mm-0.5mm particle sizes; among which, the -15mm+0.5mm particle size is separated by a heavy medium cyclone to obtain a heavy medium concentrate and a heavy medium tailings.

[0073] The heavy medium tailings obtained from S46.S45 are subjected to heavy medium scavenging to obtain heavy medium scavenging concentrate and scavenging tailings. The scavenging concentrate is returned to S45 operation as a roughing feed for the heavy medium cyclone, and the heavy medium scavenging tailings with a P2O5 grade of <10% is used as the final tailings.

[0074] The heavy medium concentrate obtained from the S47.S45 operation is further crushed and merged into the -0.5mm particle size obtained from S45. The obtained mixed -0.5mm particle size is screened and graded to obtain -0.5mm+0.038mm and -0.038mm particle sizes; among them, the -0.5mm+0.038mm particle size is re-selected by a spiral chute to obtain re-selection concentrate and re-selection tailings.

[0075] The gravity separation tailings obtained from S48 and S47 are scavenged by spiral chute to obtain gravity separation scavenging concentrate and scavenging tailings. The scavenging concentrate is returned to S47 operation as the roughing feed of spiral chute, and the gravity separation scavenging tailings with P2O5 grade <10% are used as the final tailings.

[0076] S49. Combine the -0.038mm particle size and the gravity separation concentrate obtained by S48 to obtain a pre-enriched concentrate; combine the photoelectric separation tailings obtained by S44, the heavy medium tailings obtained by S46 and the gravity separation tailings obtained by S48 to obtain a pre-enriched tailings.

[0077] S410. Grind the pre-enriched concentrate to obtain flotation pulp, and adopt a coarse-fine-sweep flotation operation to obtain the final phosphate concentrate and the final tailings. Specifically, the following steps are included:

[0078] A1. Add 2000 g / t of the original ore inhibitor and 800 g / t of the original ore collector prepared in Example 7 to the flotation slurry, and stir and aerate to obtain a rougher concentrate and a rougher tailings;

[0079] A2. Adding 500 g / t of the raw ore inhibitor and 200 g / t of the raw ore collector prepared in Example 7 to the rougher concentrate, stirring and aerating to obtain the final phosphate concentrate and selected ore;

[0080] A3. Adding 1000 g / t of the inhibitor of the raw ore and 400 g / t of the raw ore of the collector prepared in Example 7 to the rougher tailings, stirring and aerating to obtain a scavenged concentrate and a final tailings;

[0081] The selected middlings and scavenged concentrates can be returned to step A1 respectively, and steps A1 to A3 are repeated;

[0082] In parts by weight, the inhibitor includes 3 parts of sulfuric acid, 2 parts of phosphoric acid, 2 parts of hydroxyethylene diphosphoric acid and 3 parts of sodium hexametaphosphate.

[0083] Embodiment 11

[0084] A high-siliceous collophosphate ore, obtained from a concentrator in Mabian area, with a P2O5 grade of 19%. Its concentrating process is as follows: Figure 1 As shown, the following steps are included:

[0085] S41. Crushing the raw ore to obtain crushed ore with a particle size of ≤30 mm;

[0086] S42. The crushed ore is screened to obtain ores of -30mm+15mm and -30mm-15mm particle sizes, respectively;

[0087] S43. Photoelectric separation process is used for -30mm+15mm particle size ore to obtain photoelectric concentrate and photoelectric tailings;

[0088] S44. The photoelectric tailings obtained are subjected to photoelectric scavenging separation to obtain photoelectric scavenging concentrate and scavenging tailings respectively. The scavenging concentrate is returned to S43 operation, and the photoelectric scavenging tailings with P2O5 grade <10% are used as final tailings.

[0089] S45. The photoelectric concentrate obtained in S43 is further crushed and merged into the -15mm particle size obtained in S42. The obtained mixed -15mm particle size is screened and classified to obtain -15mm+0.5mm and -15mm-0.5mm particle sizes; among which, the -15mm+0.5mm particle size is separated by a heavy medium cyclone to obtain a heavy medium concentrate and a heavy medium tailings.

[0090] The heavy medium tailings obtained from S46.S45 are subjected to heavy medium scavenging to obtain heavy medium scavenging concentrate and scavenging tailings. The scavenging concentrate is returned to S45 operation as a roughing feed for the heavy medium cyclone, and the heavy medium scavenging tailings with a P2O5 grade of <10% is used as the final tailings.

[0091] The heavy medium concentrate obtained from the S47.S45 operation is further crushed and merged into the -0.5mm particle size obtained from S45. The obtained mixed -0.5mm particle size is screened and graded to obtain -0.5mm+0.038mm and -0.038mm particle sizes; among them, the -0.5mm+0.038mm particle size is re-selected by a spiral chute to obtain re-selection concentrate and re-selection tailings.

[0092] The gravity separation tailings obtained from S48 and S47 are scavenged by spiral chute to obtain gravity separation scavenging concentrate and scavenging tailings. The scavenging concentrate is returned to S47 operation as the roughing feed of spiral chute, and the gravity separation scavenging tailings with P2O5 grade <10% are used as the final tailings.

[0093] S49. Combine the -0.038mm particle size and the gravity separation concentrate obtained by S48 to obtain a pre-enriched concentrate; combine the photoelectric separation tailings obtained by S44, the heavy medium tailings obtained by S46 and the gravity separation tailings obtained by S48 to obtain a pre-enriched tailings.

[0094] S410. Grind the pre-enriched concentrate to obtain flotation pulp, and adopt a coarse-fine-sweep flotation operation to obtain the final phosphate concentrate and the final tailings. Specifically, the following steps are included:

[0095] A1. Add 2000 g / t of the original ore inhibitor and 800 g / t of the original ore collector prepared in Example 8 to the flotation slurry, and stir and aerate to obtain a rougher concentrate and a rougher tailings;

[0096] A2. Adding 500 g / t of the raw ore inhibitor and 200 g / t of the raw ore collector prepared in Example 8 to the rougher concentrate, stirring and aerating to obtain the final phosphate concentrate and selected ore;

[0097] A3. Adding 1000 g / t of the inhibitor of the raw ore and 400 g / t of the raw ore of Example 8 to the rougher tailings, stirring and aerating to obtain a scavenged concentrate and a final tailings;

[0098] The selected middlings and scavenged concentrates can be returned to step A1 respectively, and steps A1 to A3 are repeated;

[0099] In parts by weight, the inhibitor includes 2 parts of sulfuric acid, 1 part of phosphoric acid, 1 part of hydroxyethylene diphosphoric acid and 3 parts of sodium hexametaphosphate.

[0100] Example 12

[0101] A high-siliceous collophosphate ore, obtained from a concentrator in Jinyang, with a P2O5 grade of 18%. Figure 1 As shown, the following steps are included:

[0102] S41. Crushing the raw ore to obtain crushed ore with a particle size of ≤30 mm;

[0103] S42. The crushed ore is screened to obtain ores of -30mm+15mm and -30mm-15mm particle sizes, respectively;

[0104] S43. Photoelectric separation process is used for -30mm+15mm particle size ore to obtain photoelectric concentrate and photoelectric tailings;

[0105] S44. The photoelectric tailings obtained are subjected to photoelectric scavenging separation to obtain photoelectric scavenging concentrate and scavenging tailings respectively. The scavenging concentrate is returned to S43 operation, and the photoelectric scavenging tailings with P2O5 grade <10% are used as final tailings.

[0106] S45. The photoelectric concentrate obtained in S43 is further crushed and merged into the -15mm particle size obtained in S42. The obtained mixed -15mm particle size is screened and classified to obtain -15mm+0.5mm and -15mm-0.5mm particle sizes; among which, the -15mm+0.5mm particle size is separated by a heavy medium cyclone to obtain a heavy medium concentrate and a heavy medium tailings.

[0107] The heavy medium tailings obtained from S46.S45 are subjected to heavy medium scavenging to obtain heavy medium scavenging concentrate and scavenging tailings. The scavenging concentrate is returned to S45 operation as a roughing feed for the heavy medium cyclone, and the heavy medium scavenging tailings with a P2O5 grade of <10% is used as the final tailings.

[0108] The heavy medium concentrate obtained from the S47.S45 operation is further crushed and merged into the -0.5mm particle size obtained from S45. The obtained mixed -0.5mm particle size is screened and graded to obtain -0.5mm+0.038mm and -0.038mm particle sizes; among them, the -0.5mm+0.038mm particle size is re-selected by a spiral chute to obtain re-selection concentrate and re-selection tailings.

[0109] The gravity separation tailings obtained from S48 and S47 are scavenged by spiral chute to obtain gravity separation scavenging concentrate and scavenging tailings. The scavenging concentrate is returned to S47 operation as the roughing feed of spiral chute, and the gravity separation scavenging tailings with P2O5 grade <10% are used as the final tailings.

[0110] S49. Combine the -0.038mm particle size and the gravity separation concentrate obtained by S48 to obtain a pre-enriched concentrate; combine the photoelectric separation tailings obtained by S44, the heavy medium tailings obtained by S46 and the gravity separation tailings obtained by S48 to obtain a pre-enriched tailings.

[0111] S410. Grind the pre-enriched concentrate to obtain flotation pulp, and adopt a coarse-fine-sweep flotation operation to obtain the final phosphate concentrate and the final tailings. Specifically, the following steps are included:

[0112] A1. Add 2000 g / t of the original ore inhibitor and 800 g / t of the original ore collector prepared in Example 9 to the flotation slurry, and stir and aerate to obtain a rougher concentrate and a rougher tailings;

[0113] A2. Adding 500 g / t of the raw ore inhibitor and 200 g / t of the raw ore collector prepared in Example 9 to the rougher concentrate, stirring and aerating to obtain the final phosphate concentrate and selected ore;

[0114] A3. Adding 1000 g / t of the inhibitor of the raw ore and 400 g / t of the raw ore of the collector prepared in Example 9 to the rougher tailings, stirring and aerating to obtain a scavenged concentrate and a final tailings;

[0115] The selected middlings and scavenged concentrates can be returned to step A1 respectively, and steps A1 to A3 are repeated;

[0116] In parts by weight, the inhibitor includes 2 parts of sulfuric acid, 2 parts of phosphoric acid, 2 parts of hydroxyethylene diphosphoric acid and 2 parts of sodium hexametaphosphate.

[0117] Comparative Example 1

[0118] The various indicators of separating high-silica collophosphate in Example 10 of the present invention are compared with those in Control Example 1, wherein Control Example 1 uses the high-silica collophosphate in Example 10, and its beneficiation process is: the collector is replaced by a fatty acid salt mainly composed of sodium oleate in the prior art, and the inhibitor is replaced by sulfuric acid and phosphoric acid in a mass ratio of 5:1 in the prior art, and other conditions such as the dosage of the reagent, the grinding fineness and the adopted process flow are the same as those in Example 10 of the present invention. (This control example is compared with replacing the collector and the inhibitor with the prior art at the same time, to prove that the beneficiation process of the present invention has a better effect).

[0119] Comparative Example 2

[0120] The various indicators of separating high-silica collophanite in Example 10 of the present invention are compared with those in Control Example 2, wherein Control Example 2 uses the high-silica collophanite in Example 10, and its beneficiation process is: the inhibitor is replaced with sulfuric acid and phosphoric acid in a mass ratio of 5:1 in the prior art, and other conditions such as collector selection, reagent dosage, grinding fineness and adopted process flow are the same as those in Example 10 of the present invention. (This control example is compared with replacing the inhibitor with the prior art alone, to prove that the beneficiation process of the present invention has better effect).

[0121] Comparative Example 3

[0122] The various indicators of separating high-silica collophanite in Example 10 of the present invention are compared with those in Control Example 3, wherein Control Example 3 uses the high-silica collophanite in Example 10, and its beneficiation process is: the collector is replaced with a fatty acid salt mainly composed of sodium oleate in the prior art, and other conditions such as the dosage of the reagent, the grinding fineness and the adopted process flow are the same as those in Example 10 of the present invention. (This control example is compared with the case where the collector is replaced with the prior art at the same time, to prove that the beneficiation process of the present invention has a better effect).

[0123] The experimental results of Examples 10 to 12 and Comparative Examples 1 to 3 are compared as shown in Table 1 below (wherein, the raw ore grade in Comparative Examples 1 to 3 is slightly different from that in Example 10 because of human error in the experimental operation, which is a normal phenomenon):

[0124] Table 1 Comparison results of Examples 10 to 12 and Comparative Examples 1 to 3

[0125]

[0126]

[0127] As can be seen from the above table, compared with Example 10, the yield, grade and recovery rate of the final phosphate concentrate obtained in Reference Example 1 are significantly reduced, while the yield, grade and recovery rate of the obtained tailings are significantly increased; the yield of the final phosphate concentrate obtained in Reference Example 2 is slightly reduced, the grade and recovery rate are significantly reduced, while the yield of the obtained tailings is slightly increased, and the grade and recovery rate are significantly increased; the yield of the final phosphate concentrate obtained in Reference Example 3 is increased, and the grade is significantly reduced, while the yield of the obtained tailings is decreased, and the grade is significantly increased. It can be seen that Example 10 of the present invention can more fully separate the phosphate concentrate from the raw ore compared with the methods of Reference Examples 1 to 3; at the same time, although the yield of Reference Example 3 is increased, its grade is significantly reduced, which means that other impurities float together, proving that the collector performance of Reference Example 3 is poor compared with Example 10.

[0128] Therefore, the high-silica collophanite flotation separation reagent system and its application provided by the present invention achieve the effects of high ore dressing efficiency, small amount of ore entering the float, low energy consumption, low flotation reagent cost, small amount of flotation reagent used and environmental friendliness.

[0129] The above is only a preferred embodiment of the present invention. It should be understood that the present invention is not limited to the form disclosed herein, and should not be regarded as excluding other embodiments, but can be used in various other combinations, modifications and environments, and can be modified within the scope of the concept described herein through the above teachings or the technology or knowledge of the relevant field. The changes and modifications made by those skilled in the art do not deviate from the spirit and scope of the present invention, and should be within the scope of protection of the claims attached to the present invention.

Claims

1. A high-silica collophanite flotation separation reagent system, characterized in that: Includes inhibitors and collectors; The inhibitor comprises 2-3 parts of sulfuric acid, 1-2 parts of phosphoric acid, 1-2 parts of hydroxyethylene diphosphoric acid and 2-3 parts of sodium hexametaphosphate; the collector comprises 4-5 parts of alkylamino salt, 1-2 parts of octadecyl phosphate, 1-2 parts of laurate and 3-4 parts of vegetable oil acid salt.

2. The high-silica collophanite flotation separation reagent system according to claim 1 is characterized in that: The preparation method of the alkyl amino acid salt comprises the following steps: adding methyl linoleate to dodecyl primary amine, stirring and then performing a first heating reaction to obtain a reaction solution; and NaOH is added to the reaction solution to carry out a second heating reaction to obtain the alkylamino salt.

3. The high-silica collophanite flotation separation reagent system according to claim 2 is characterized in that: The weight ratio of the dodecyl primary amine to methyl linoleate is 1:(3-4); The weight ratio of the NaOH to the reaction solution is (0.1-0.2):

1.

4. The high-silica collophanite flotation separation reagent system according to claim 2, characterized in that: The temperature of the first heating reaction is 100 to 120° C., and the time of the first heating reaction is 40 to 50 minutes; The temperature of the second heating reaction is 60-80° C., and the time of the second reaction is 3-5 hours.

5. The high-silica collophanite flotation separation reagent system according to claim 1, characterized in that: The preparation method of the collector comprises the following steps: After mixing octadecyl phosphate, lauric acid and vegetable oil, a mixed solution is obtained; adding NaOH to the mixed solution to perform a third heating reaction to obtain a primary collector; as well as The alkyl amino acid salt is added into the primary collector and mixed to obtain the collector.

6. The high-silica collophanite flotation separation reagent system according to claim 5, characterized in that: The vegetable oil comprises cottonseed oil and linoleic acid, and the mass ratio of the cottonseed oil to the linoleic acid is (1.5-2.5):

1.

7. The high-silica collophanite flotation separation reagent system according to claim 5, characterized in that: The weight ratio of the NaOH to the mixed solution is (0.1-0.2):

1.

8. The high-silica collophanite flotation separation reagent system according to claim 5, characterized in that: The temperature of the third heating reaction is 60-80° C. and the time is 3-5 hours.

9. Application of the high-silica collophanite flotation separation reagent system according to any one of claims 1 to 8 in the flotation separation of high-silica collophanite, characterized in that: The steps include: The raw ore is crushed, screened, photoelectrically scavenged and heavy medium scavenged in sequence to obtain a pre-enriched concentrate, and water is added to the pre-enriched concentrate to obtain a flotation slurry; Adding 2000-3000 g / t of the depressant and 400-800 g / t of the collector to the flotation pulp, and obtaining rougher concentrate and rougher tailings through stirring and aeration; Adding 400-600 g / t of the depressant and 40-80 g / t of the collector to the rougher concentrate, and stirring and aerating to obtain the final phosphate concentrate and the selected middlings; as well as 800-1200 g / t of the inhibitor and 150-250 g / t of the collector are added to the rougher tailings, and scavenger concentrate and final tailings are obtained through stirring and aeration.

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