Classification flotation method for medium-low-grade silicon-calcium collophanite

Abstract

The invention belongs to the technical field of phosphorite beneficiation, and particularly discloses a graded flotation method of medium-low-grade silicon-calcium collophanite, which comprises the following steps: crushing collophanite, grinding and mixing to obtain phosphorite pulp, and reselecting by a spiral chute to obtain reselected concentrate and reselected tailings; the gravity concentrate is subjected to magnesium removal reverse flotation and desilicication reverse flotation in sequence, and concentrate and tailings are obtained; gravitational separation tailings are graded through grading equipment, and coarse-fraction products and fine-fraction products are obtained; carrying out desiliconization direct flotation and magnesium removal reverse flotation after grinding the coarse fraction product to obtain concentrate and tailings; thickening the fine fraction product, and sequentially carrying out magnesium removal reverse flotation and desilicication reverse flotation to obtain concentrate and tailings; and combining the concentrates to obtain the phosphate concentrate. According to the gravity separation tailings, coarse fractions which are not fully dissociated are separated out through grading equipment and a grading technology, ore grinding is conducted independently, the ore grinding efficiency can be improved, and minerals can be fully dissociated; the fine fraction in the heavy tailings is directly subjected to double reverse flotation after being thickened, and the flotation efficiency is high.

Description

Technical Field

[0001] The invention belongs to the technical field of phosphate ore dressing, and relates to a classification flotation method for medium- and low-grade calci-silico-phosphatic ore. Background Art

[0002] The average grade of siliceous calcareous collophanite is low, the impurity content is high, silicate and carbonate are common gangue minerals, and the collophanite presents a coarse grained embedded band structure, which brings difficulties to sorting. At present, the sorting of this type of ore mainly adopts heavy medium sorting or photoelectric sorting combined with single reverse flotation, or double reverse flotation, but in the above processes, there are problems such as large phosphorus loss and low recovery rate.

[0003] CN10945389A discloses a combined refloatation process of spiral chute for high-fold semi-colloid phosphate, which adopts combined refloatation and ore dressing, firstly adopts spiral chute concentrator for re-selection to obtain re-selection concentrate and re-selection tailings, and then performs double reverse flotation and forward-reverse flotation on the re-selection concentrate and re-selection tailings respectively. This method can avoid the negative impact of coarse and fine particle inclusions on flotation in flotation engineering, but the re-selection process only includes one spiral chute roughing operation, and the quality of re-selection products fluctuates greatly.

[0004] CN117258997A discloses a method for refloating and selecting collophanite in a spiral chute with multiple particle size products. The refloating process obtains three particle size products: coarse, medium and fine. The coarse and fine particle size products directly enter the reverse flotation process, and the medium particle size product is returned to the mill for re-grinding. The method can achieve secondary dissociation of the target mineral, but the ore belt of the spiral chute product is greatly affected by operating factors, and the quality of the three products is prone to fluctuation. In addition, the hardness of the medium ore product is relatively high, and it is directly returned to be mixed with the original ore for grinding, so the degree of secondary dissociation is limited.

[0005] Liu Yutong et al. disclosed the low-temperature flotation of low-grade phosphate ore in Yichang by modified fatty acids. Cottonseed oil fatty acid soap was used as raw material and polyoxyethylene ether non-ionic surfactant was compounded to prepare anionic collector MON-135. The anionic collector was able to effectively float colloidal phosphate ore at a temperature of 15°C. However, the grinding fineness required was -0.074mm and the content was above 93%, resulting in a high beneficiation cost. Summary of the invention

[0006] The invention provides a classification flotation method for medium- and low-grade calcareous siliceous collophanite, which has a reasonable process flow and stable products.

[0007] In order to achieve the above objectives, the present invention adopts the following technical solution: a classification flotation method for medium and low grade calcareous siliceous collophosphate, characterized in that it comprises the following steps: S1, crushing the collophosphate ore, grinding and slurrying to obtain phosphate slurry, and performing gravity separation through a spiral chute to obtain gravity separation concentrate and gravity separation tailings; S2, the gravity separation concentrate is subjected to magnesium removal reverse flotation to obtain magnesium removal concentrate and magnesium removal tailings 1, and the magnesium removal concentrate is further subjected to desiliconization reverse flotation to obtain concentrate 1 and desiliconization tailings 1; S3, the gravity separation tailings are classified by a classification device to obtain coarse-grained products and fine-grained products; the coarse-grained products are ground and then subjected to desiliconization direct flotation to obtain desiliconized tailings 2 and tank bottom materials, and the tank bottom minerals are then subjected to demagnesiumization reverse flotation to obtain concentrate 2 and demagnesiumized tailings 2; the fine-grained products are thickened and then subjected to demagnesiumization reverse flotation to obtain demagnesiumized concentrate and demagnesiumized tailings 3, and the demagnesiumized concentrate is then subjected to desiliconization reverse flotation to obtain concentrate 3 and desiliconized tailings 3; S4. The concentrates obtained in the above steps are combined to form phosphate concentrate, and the tailings obtained by combining are final phosphate tailings.

[0008] Optionally, the medium-low grade calcareous siliceous colloid ore is a high-magnesium-aluminum calcareous siliceous colloid ore, P 2 O 5 The grade is 19wt%-26wt%, the MgO content is 1wt%-6wt%, and the Al 2 O 3 The content is 5wt%-8wt%; grinding control -0.074mm accounts for 50wt%-75wt%; the concentration of phosphate slurry fed into the spiral chute after slurry adjustment is 15wt%-25wt%.

[0009] Optionally, during re-selection in S1, at least one of the following conditions is met: the diameter of the spiral chute is 500mm-800mm; the pitch-to-diameter ratio is 0.36-0.45; the number of turns is 4-6; and the dry ore processing capacity of a single spiral chute is 0.5-3t / h.

[0010] Optionally, when demagnesiumizing reverse flotation is performed in S2, the pulp concentration is 20%-30%, and acid is added as an inhibitor in an amount of 1kg / t-3kg / t; 1kg / t-5kg / t of demagnesiumizing collector is added for reverse flotation; and the amount of desiliconizing collector added during desiliconizing reverse flotation is 100g / t-500g / t.

[0011] Optionally, the grading equipment in S3 is a hydrocyclone or a vibrating screen; the underflow product obtained by hydrocyclone grading is the coarse-grained product, which contains more than 70% of +0.074mm; the overflow product obtained by hydrocyclone grading is the fine-grained product, which contains less than 8% of +0.074mm; if vibrating screen grading is used, the coarse-grained product corresponds to the product on the vibrating screen, which contains more than 90% of +0.074mm, and the fine-grained product corresponds to the product under the vibrating screen, which contains less than 6% of +0.074mm.

[0012] Optionally, after grinding the coarse-grained product, it is required that the grinding product with a particle size of less than 0.074mm accounts for more than 70%, the positive flotation pulp concentration is 25%-35%, the positive flotation adjuster is sodium carbonate, the dosage is 5kg / t-10kg / t, the positive flotation depressant is sodium silicate, the dosage is 1kg / t-5kg / t, the reverse flotation demagnesiumization pulp concentration of the bottom mineral is 20%-30%, the reverse flotation depressor is one or both of phosphoric acid or sulfuric acid, and the dosage of the demagnesiumization collector is 1kg / t-5kg / t.

[0013] Optionally, the fine-grained product is thickened by a thickener to a pulp concentration of 20-25%, and is fed into a reverse flotation process. The reverse flotation depressant is one or both of phosphoric acid and sulfuric acid, and the dosage is 2kg / t-4kg / t. The dosage of the magnesium removal collector is 1kg / t-5kg / t, and the dosage of the silicon removal collector is 100g / t-500g / t.

[0014] Optionally, the demagnesium collector is a fatty acid anion collector, which is mixed with linoleic acid, oleic acid, palmitoleic acid, myristic acid, linolenic acid and lauric acid in a certain proportion, wherein linoleic acid accounts for 20-25wt%, oleic acid accounts for 10-15wt%, palmitoleic acid accounts for 20-25wt%, linolenic acid accounts for 20-25wt%, lauric acid accounts for 15-25wt%, and the sum of the mass percentages of each component is 100%.

[0015] Optionally, the reverse flotation desiliconization collector is an organic ammonia cationic collector, including dodecylamine, N, N-dimethyldodecylamine, ether ammonia Y-10, ether polyamine K-2036 and polyoxyethylene hexadecyl ether mixed in a certain proportion, wherein dodecylamine accounts for 20wt%-30wt%, N, N-dimethyldodecylamine accounts for 15wt%-20wt%, ether ammonia Y-10 accounts for 15wt%-20wt%, ether polyamine K-2036 accounts for 25wt%-30wt%, and polyoxyethylene hexadecyl ether accounts for 10wt%-15wt%.

[0016] Optionally, the positive flotation collector in S3 is prepared by mixing oxidized paraffin soap, polyoxyethylene ether, tall oil, benzylphenol polyoxyethylene ether, and octylphenol polyoxyethylene ether in a mass percentage of 20%-30%: 20%-30%: 30%-40%: 10%-20%: 10%-20%, and heating and saponifying, and the sum of the mass percentages of each component is 100%.

[0017] The present invention has the following beneficial effects: (1) Through a spiral chute gravity separation process, the heavy concentrate with higher grade, coarser particle size and better monomer dissociation degree is first floated, avoiding the negative impact of coarse and fine particle inclusions on flotation.

[0018] (2) The coarse particles in the heavy tailings are separated and ground separately through grading equipment and processes, which solves the problem that some phosphate minerals cannot be effectively separated due to insufficient dissociation, thereby improving the phosphorus recovery rate. The fine particles in the heavy tailings are directly subjected to double reverse flotation after concentration, which has high flotation efficiency.

[0019] (3) After heavy tail coarse particle grinding, the main gangue minerals are silicate minerals. Therefore, positive flotation is first used to remove silicate minerals, and then reverse flotation is used to remove carbonate minerals from the bottom minerals. This is in line with the principle of "floating less and suppressing more", which is beneficial to reducing reagent costs and improving flotation efficiency.

[0020] (4) The modified fatty acid is used as a collector for positive flotation desiliconization, which can achieve flotation at room temperature and low temperature with good selectivity and flotation performance. BRIEF DESCRIPTION OF THE DRAWINGS

[0021] Figure 1 A process flow chart of the method is provided for the present invention. DETAILED DESCRIPTION

[0022] The embodiments of the present invention will be described in detail below with reference to examples, but those skilled in the art will appreciate that the following examples are only used to illustrate the present invention and should not be construed as limiting the scope of the present invention.

[0023] Example 1 A classification flotation method for treating medium and low grade calcareous silicate collophanite, such as Figure 1 As shown, the method specifically comprises the following steps: (1) The ore composition is P 2 O 5 The grade is 24wt%, the MgO content is 2.81wt%, and the Al 2 O 3 The content is 7.66wt%, grinding -0.074mm accounts for 65wt%, and after slurry adjustment, it is fed into a spiral chute (spiral chute diameter 700mm, 5 turns, pitch-diameter ratio 0.4, feed concentration 25wt%) for a roughing and gravity separation to obtain gravity separation concentrate (operating yield 60%) and gravity separation tailings (operating yield 40%). (2) The re-selected concentrate was fed into the reverse flotation process for reverse flotation of magnesium removal and silicon removal in sequence. The flotation pulp concentration was 25%, the amount of inhibitor phosphoric acid was 2 kg / t, the amount of magnesium removal collector (main components were linoleic acid, oleic acid, palmitoleic acid, myristic acid, linolenic acid, lauric acid were configured in the proportions of 25%: 15%: 20%: 23%: 17%, the same below) was 2.5 kg / t, and the amount of silicon removal collector (main components were dodecylamine, N, N-dimethyldodecylamine, etheramine Y-10, ether polyamine K-2036, polyoxyethylene hexadecyl ether were configured in the proportions of 20%: 25%: 35%: 10%: 10%, the same below) was 300 g / t, and concentrate 1, magnesium removal tailings 1, and silicon removal tailings 1 were obtained. (3) The gravity separation tailings are classified by a hydrocyclone with a feed concentration of 15% and a pressure of 0.1 MPa to obtain coarse-grained products (underflow, operating yield 20%) and fine-grained products (overflow, operating yield 80%). (4) The coarse-grained product is ground, with a grinding concentration of 65% and a grinding fineness of -200 mesh accounting for 72%. The ground product is fed into the direct flotation process to first undergo dealuminization direct flotation, with a flotation pulp concentration of 25%, a sodium carbonate dosage of 5 kg / t, and a sodium silicate dosage of 2 kg / t; the bottom minerals are then subjected to demagnesiumization reverse flotation, with a flotation pulp concentration of 20%, a phosphoric acid dosage of 4 kg / t, and a demagnesiumization collector of 2 kg / t, to obtain concentrate 2, desiliconized tailings 2, and demagnesiumized tailings 2; (5) The fine-grained product is thickened by a thickener to an underflow concentration of 20% and then pumped to a reverse flotation process for double reverse flotation of de-magnesiumization and de-aluminization. The amount of phosphoric acid used is 2 kg / t, the amount of de-magnesiumization collector used is 3 kg / t, and the amount of de-siliconization collector used is 450 g / t, to obtain concentrate 3, de-magnesiumization tailings 3, and de-siliconization tailings 3; The main component contents and yields of the above products are shown in Table 1 below.

[0024] Table 1

[0025] (6) The de-magnesiumized tailings 1, de-magnesiumized tailings 2, de-magnesiumized tailings 3 and the de-siliconized tailings 1, de-siliconized tailings 2, de-siliconized tailings 3 are combined to obtain the final tailings; the concentrates 1, concentrates 2 and concentrates 3 are combined to obtain the final phosphate concentrate, P 2 O 5 The grade is 31.16% and the phosphorus recovery rate is 85.23%.

[0026] Comparative Example 1 Compared with Example 1, the only difference is that the gravity separation tailings are not classified, and the heavy tailings samples are directly subjected to double reverse flotation. Other flotation reagents and process systems are completely consistent with the above embodiment. The specific steps are as follows: (1) The ore composition is P 2 O 5The grade is 24wt%, the MgO content is 2.81wt%, and the Al 2 O 3 The content is 7.66wt%, grinding -0.074mm accounts for 65wt%, and after slurry adjustment, it is fed into a spiral chute (spiral chute diameter 700mm, 5 turns, pitch-diameter ratio 0.4, feed concentration 25wt%) for a roughing and gravity separation to obtain gravity separation concentrate (operating yield 60%) and gravity separation tailings (operating yield 40%). (2) The gravity separation concentrate was fed into the reverse flotation process for reverse flotation of demagnesiumization and desiliconization in sequence, with the flotation pulp concentration of 25%, the amount of inhibitor phosphoric acid used being 2 kg / t, the amount of demagnesiumization collector used being 2.5 kg / t, and the amount of desiliconization collector used being 300 g / t, to obtain concentrate 1, demagnesiumization tailing 1, and desiliconization tailing 1; (3) The gravity separation tailings are fed into the direct flotation process to firstly carry out dealumination direct flotation, with a flotation pulp concentration of 25%, a sodium carbonate dosage of 5 kg / t, and a sodium silicate dosage of 2 kg / t; the bottom minerals are then subjected to demagnesiumization reverse flotation, with a flotation pulp concentration of 20%, a phosphoric acid dosage of 4 kg / t, and a demagnesiumization collector of 2 kg / t, to obtain concentrate 2, desiliconized tailings 2, and demagnesiumized tailings 2; The specific product indicators are shown in Table 2.

[0027] Table 2 The main component content and yield of each product of heavy tail direct flotation process

[0028] Comparing Table 2 with Table 1, the grade of concentrate 2 is obviously lower, and the grade of tailings 2 is also higher. The MgO and SiO contents corresponding to concentrate 2 in Table 2 are also higher, indicating that the separation degree of gangue minerals in direct flotation of heavy tailings is low, and the sorting effect is obviously worse.

[0029] Comparative Example 2 Compared with Example 1, the only difference is that the coarse particle size of the heavy tailings sample after gravity separation tailings classification is not ground, but directly subjected to double reverse flotation. Other flotation reagents and process systems are completely consistent with the above embodiment. The specific steps are as follows: (1) The ore composition is P 2 O 5 The grade is 24wt%, the MgO content is 2.81wt%, and the Al 2 O 3 The content is 7.66wt%, grinding -0.074mm accounts for 65wt%, and after slurry adjustment, it is fed into a spiral chute (spiral chute diameter 700mm, 5 turns, pitch-diameter ratio 0.4, feed concentration 25wt%) for a roughing and gravity separation to obtain gravity separation concentrate (operating yield 60%) and gravity separation tailings (operating yield 40%). (2) The gravity separation concentrate was fed into the reverse flotation process for demagnesiumization and desiliconization reverse flotation in sequence, the flotation pulp concentration was 25%, the inhibitor phosphoric acid dosage was 2 kg / t, the demagnesiumization collector dosage was 2.5 kg / t, and the dealuminization collector dosage was 300 g / t, to obtain concentrate 1, demagnesiumization tailings 1, and desiliconization tailings 1; (3) The gravity separation tailings are classified by a hydrocyclone with a feed concentration of 15% and a pressure of 0.1 MPa to obtain coarse-grained products (underflow, operating yield 20%) and fine-grained products (overflow, operating yield 80%). (4) The coarse-grained product is directly fed into the direct flotation process to firstly carry out dealumination direct flotation, with a flotation pulp concentration of 25%, a sodium carbonate dosage of 5 kg / t, and a sodium silicate dosage of 2 kg / t; the bottom minerals are then subjected to demagnesium reverse flotation, with a flotation pulp concentration of 20%, a phosphoric acid dosage of 4 kg / t, and a demagnesium collector) of 2 kg / t, to obtain concentrate 2, desiliconized tailings 2, and demagnesiumized tailings 2; (5) The fine-grained product is thickened by a thickener to an underflow concentration of 20% and then pumped to a reverse flotation process for double reverse flotation of de-magnesiumization and de-aluminization. The amount of phosphoric acid used is 2 kg / t, the amount of de-magnesiumization collector used is 3 kg / t, and the amount of de-siliconization collector used is 450 g / t, to obtain concentrate 3, de-magnesiumization tailings 3, and de-siliconization tailings 3; (6) The demagnesiumized tailings 1, demagnesiumized tailings 2, demagnesiumized tailings 3 and the desiliconized tailings 1, desiliconized tailings 2, desiliconized tailings 3 are combined to obtain the final tailings; the concentrates 1, concentrates 2 and concentrates 3 are combined to obtain the final phosphate concentrate 30.04%, and the phosphorus recovery rate is 82.21%.

[0030] The specific product indicators are shown in Table 3.

[0031] Table 3 The main component contents and yields of each product of the heavy tail classification coarse particle direct flotation process

[0032] By comparing the data in Table 1 and Table 3, it can be seen that the concentrate 2 grade obtained by direct flotation of the coarse-particle product after heavy-tail classification without grinding process is only 26.30%. The final phosphate concentrate grade dropped by 1 percentage point and the recovery rate dropped by about 3 percentage points, indicating that grinding the heavy-tail coarse-particle product is beneficial to improving the grade and recovery rate of the final concentrate.

[0033] Example 2 Compared with Example 1, the only difference is that the magnesium-removing collector is a mixture of linoleic acid, oleic acid, palmitoleic acid, myristic acid, linolenic acid, and lauric acid in the ratio of 20%: 10%: 25%: 25%: 20%, respectively, and the dosage of other agents is consistent with the process steps. The product indicators are shown in Table 4 below.

[0034] Table 4

[0035] Comparison of Table 1 and Table 4 shows that by appropriately reducing the amount of oleic acid, better magnesium removal indicators can also be achieved, and the indicators of phosphate concentrate products are basically the same.

[0036] Comparative Example 3 Compared with Example 1, the only difference is that the magnesium removal collector is a mixture of linoleic acid, oleic acid, palmitoleic acid, myristic acid, linolenic acid, and lauric acid in the ratio of 10%: 5%: 30%: 25%: 30%, respectively, and the dosage of other agents is consistent with the process steps. The product indicators are shown in Table 5 below.

[0037] Table 5

[0038] Comparing the data in Table 1 and Table 5, it can be seen that the amount of linoleic acid and oleic acid is reduced, the MgO content in the phosphate concentrate is increased, and the phosphorus grade is reduced, indicating that the flotation index of the flotation agent not prepared according to the formula provided by the present invention is poor.

[0039] Example 3 A classification flotation method for processing medium- and low-grade calcareous siliceous collophanite, the method specifically comprising the following steps: (1) The ore composition is P 2 O 5 The grade is 22.31wt%, the MgO content is 2.52wt%, and the Al 2 O 3 The content is 6.88wt%, grinding -0.074mm accounts for 70wt%, and after slurry adjustment, it is fed into a spiral chute (spiral chute diameter 650mm, 5 turns, pitch-diameter ratio 0.36, feed concentration 20wt%) for a roughing and gravity separation to obtain gravity separation concentrate (operating yield 65%) and gravity separation tailings (operating yield 35%). (2) The gravity separation concentrate was fed into the reverse flotation process for demagnesiumization and desiliconization reverse flotation in sequence, the flotation pulp concentration was 28%, the inhibitor phosphoric acid dosage was 2 kg / t, the demagnesiumization collector (same as in Example 1, the same below) dosage was 2 kg / t, and the dealuminization collector (same as in Example 1, the same below) dosage was 250 g / t, to obtain concentrate 1, demagnesiumization tailings 1, and desiliconization tailings 1; (3) The gravity separation tailings are classified by a hydrocyclone with a feed concentration of 10% and a pressure of 0.15 MPa, and coarse-grained products (underflow, operating yield 22%) and fine-grained products (overflow, operating yield 78%) are obtained; (4) The coarse-grained product is ground, with a grinding concentration of 65% and a grinding fineness of -200 mesh accounting for 75%. The ground product is fed into the direct flotation process to first undergo dealuminization direct flotation, with a flotation pulp concentration of 30%, a sodium carbonate dosage of 6 kg / t, and a sodium silicate dosage of 3 kg / t; the bottom minerals are then subjected to demagnesiumization reverse flotation, with a flotation pulp concentration of 28%, a phosphoric acid dosage of 4.5 kg / t, and a demagnesiumization collector of 2 kg / t, to obtain concentrate 2, desiliconized tailings 2, and demagnesiumized tailings 2; (5) The fine-grained product is thickened by a thickener to an underflow concentration of 20% and then pumped to a reverse flotation process for double reverse flotation of de-magnesiumization and de-aluminization. The amount of phosphoric acid used is 2.5 kg / t, the amount of de-magnesiumization collector used is 3 kg / t, and the amount of de-aluminization collector used is 400 kg / t, to obtain concentrate 3, de-magnesiumization tailings 3, and de-siliconization tailings 3; (6) The demagnesiumized tailings 1, the demagnesiumized tailings 2, the demagnesiumized tailings 3 and the desiliconized tailings 1, the desiliconized tailings 2, the desiliconized tailings 3 are combined to obtain the final tailings; and the concentrates 1, the concentrates 2 and the concentrates 3 are combined to obtain the final phosphate concentrate.

[0040] The above method obtains P in the comprehensive phosphate concentrate. 2 O 5 The grade is 31.5%, the mass content of MgO is 0.71%, and the Al 2 O 3 The mass content is 1.60%, comprehensive phosphate concentrate P 2 O 5 The recovery rate is 85.77%.

[0041] Example 4 A classification flotation method for processing medium- and low-grade calcareous siliceous collophanite, the method specifically comprising the following steps: (1) The ore composition is P 2 O 5 The grade is 20.06wt%, the MgO content is 1.84wt%, and the Al 2 O 3 The content is 7.68wt%, grinding -0.074mm accounts for 70wt%, and after slurry adjustment, it is fed into a spiral chute (spiral chute diameter 600mm, 5 turns, pitch-diameter ratio 0.42, feed concentration 25wt%) for a roughing and gravity separation to obtain gravity separation concentrate (operating yield 64%) and gravity separation tailings (operating yield 36%). (2) The gravity separation concentrate was fed into the reverse flotation process for reverse flotation of demagnesiumization and desiliconization in sequence, the flotation pulp concentration was 30%, the inhibitor phosphoric acid dosage was 2.5 kg / t, the demagnesiumization collector (formula was the same as in Example 1, the same below) dosage was 3 kg / t, and the dealuminization collector (formula was the same as in Example 1, the same below) dosage was 350 g / t, to obtain concentrate 1, demagnesiumization tailings 1, and desiliconization tailings 1; (3) The gravity separation tailings are graded using a high-frequency screen, with a feed concentration of 13% and a screening particle size of 0.1 mm, to obtain coarse-grained products (oversize, operating yield 15%) and fine-grained products (undersize, operating yield 85%); (4) The product on the screen is ground, with a grinding concentration of 60% and a grinding fineness of -200 mesh accounting for 76%. The ground product is fed into the direct flotation process for first dealuminization direct flotation, with a flotation pulp concentration of 30%, a sodium carbonate dosage of 5 kg / t, and a sodium silicate dosage of 3.5 kg / t; the bottom minerals are then subjected to demagnesiumization reverse flotation, with a flotation pulp concentration of 27%, a phosphoric acid dosage of 4.5 kg / t, and a demagnesiumization collector of 2.5 kg / t, to obtain concentrate 2, desiliconized tailings 2, and demagnesiumized tailings 2; (5) The undersize product is thickened by a thickener to an underflow concentration of 25% and then pumped to a reverse flotation process for double reverse flotation of de-magnesiumization and de-aluminization. The amount of phosphoric acid used is 2 kg / t, the amount of de-magnesiumization collector is 2.5 kg / t, and the amount of de-aluminization collector is 450 kg / t, to obtain concentrate 3, de-magnesiumization tailings 3, and de-siliconization tailings 3; (6) The demagnesiumized tailings 1, the demagnesiumized tailings 2, the demagnesiumized tailings 3 and the desiliconized tailings 1, the desiliconized tailings 2, the desiliconized tailings 3 are combined to obtain the final tailings; and the concentrates 1, the concentrates 2 and the concentrates 3 are combined to obtain the final phosphate concentrate.

[0042] The above method obtains P in the comprehensive phosphate concentrate. 2 O 5 The grade is 33%, the mass content of MgO is 0.78%, and the Al 2 O 3 The mass content is 1.75%, comprehensive phosphate concentrate P 2 O 5 The recovery rate is 82.96% The above embodiments describe the preferred implementation of the present invention, but the present invention is not limited thereto. Within the technical concept of the present invention, the technical solution of the present invention can be subjected to a variety of simple modifications, including the combination of various technical features in any other manner, and these simple modifications and combinations should also be regarded as the contents disclosed by the present invention and belong to the protection scope of the present invention. Therefore, the protection scope of the patent of the present invention shall be subject to the attached claims.

Claims

1. A method for the classification flotation of medium and low grade calcareous siliceous collophanite, characterized in that: The following steps are involved: S1, crushing the collophosphate ore, grinding and slurrying to obtain phosphate slurry, and performing gravity separation through a spiral chute to obtain gravity separation concentrate and gravity separation tailings; S2, the gravity separation concentrate is subjected to magnesium removal reverse flotation to obtain magnesium removal concentrate and magnesium removal tailings 1, and the magnesium removal concentrate is further subjected to desiliconization reverse flotation to obtain concentrate 1 and desiliconization tailings 1; S3, the gravity separation tailings are classified by a classification device to obtain coarse-grained products and fine-grained products; the coarse-grained products are ground and subjected to desiliconization direct flotation to obtain desiliconized tailings 2 and tank bottom materials, and the tank bottom minerals are subjected to demagnesium reverse flotation to obtain concentrate 2 and demagnesiumized tailings 2; The fine-grained product is then thickened and subjected to magnesium removal reverse flotation to obtain magnesium removal concentrate and magnesium removal tailings 3, and the magnesium removal concentrate is then subjected to desiliconization reverse flotation to obtain concentrate 3 and desiliconization tailings 3; S4. The concentrates obtained in the above steps are combined to form phosphate concentrate, and the tailings obtained by combining are final phosphate tailings.

2. The method according to claim 1, characterized in that: Among them, the medium and low grade siliceous-calcium colloidal phosphate ore is high-magnesium-aluminum siliceous-calcium colloidal phosphate ore, with a P2O5 grade of 19wt%-26wt%, a MgO content of 1wt%-6wt%, and an Al2O3 content of 5wt%-8wt%; the grinding control -0.074mm accounts for 50wt%-75wt%, and the concentration of the phosphate slurry fed into the spiral chute after slurry adjustment is 15wt%-25wt%.

3. The method according to claim 1, characterized in that During re-selection in S1, at least one of the following conditions shall be met: the diameter of the spiral chute is 500mm-800mm; the pitch-to-diameter ratio is 0.36-0.45; the number of circles is 4-6; the dry ore processing capacity of a single spiral chute is 0.5-3t / h.

4. The method according to claim 1, characterized in that: During the reverse flotation of demagnesium in S2, the pulp concentration is 20%-30%, and acid is added as an inhibitor in an amount of 1kg / t-3kg / t; 1kg / t-5kg / t of demagnesium collector is added for reverse flotation; during the reverse flotation of desiliconization, the amount of desiliconization collector added is 100g / t-500g / t.

5. The method according to claim 1, characterized in that: The grading equipment in S3 is a hydrocyclone or a vibrating screen; the underflow product obtained by hydrocyclone grading is the coarse-grained product, which contains more than 70% of +0.074mm; the overflow product obtained by hydrocyclone grading is the fine-grained product, which contains less than 8% of +0.074mm; if vibrating screen grading is used, the coarse-grained product corresponds to the product on the vibrating screen, which contains more than 90% of +0.074mm, and the fine-grained product corresponds to the product under the vibrating screen, which contains less than 6% of +0.074mm.

6. The method according to claim 1, characterized in that: After grinding the coarse-grained product, it is required that the grinding product with a particle size of less than 0.074mm accounts for more than 70%, the positive flotation pulp concentration is 25%-35%, the positive flotation adjuster is sodium carbonate, the dosage is 5kg / t-10kg / t, the positive flotation depressant is sodium silicate, the dosage is 1kg / t-5kg / t, the reverse flotation demagnesium pulp concentration of the bottom mineral is 20%-30%, the reverse flotation depressor is one or both of phosphoric acid or sulfuric acid, and the dosage of the demagnesium collector is 1kg / t-5kg / t.

7. The method according to claim 1, characterized in that: The fine-grained product is thickened by a thickener to a pulp concentration of 20-25%, and is fed into a reverse flotation process. The reverse flotation depressant is one or both of phosphoric acid and sulfuric acid, with a dosage of 2kg / t-4kg / t, a demagnesium collector dosage of 1kg / t-5kg / t, and a desiliconization collector dosage of 100g / t-500g / t.

8. The method according to claim 4 or 7, characterized in that: The magnesium removal collector is a fatty acid anion collector, which is mixed with linoleic acid, oleic acid, palmitoleic acid, myristic acid, linolenic acid and lauric acid in a certain proportion, wherein linoleic acid accounts for 20-25wt%, oleic acid accounts for 10-15wt%, palmitoleic acid accounts for 20-25wt%, linolenic acid accounts for 20-25wt%, lauric acid accounts for 15-25wt%, and the sum of the mass percentages of each component is 100%.

9. The method according to claim 4 or 7, characterized in that: The reverse flotation desiliconization collector is an organic ammonia cationic collector, which includes dodecylamine, N, N-dimethyldodecylamine, ether ammonia Y-10, ether polyamine K-2036 and polyoxyethylene hexadecyl ether mixed in a certain proportion, wherein dodecylamine accounts for 20wt%-30wt%, N, N-dimethyldodecylamine accounts for 15wt%-20wt%, ether ammonia Y-10 accounts for 15wt%-20wt%, ether polyamine K-2036 accounts for 25wt%-30wt%, and polyoxyethylene hexadecyl ether accounts for 10wt%-15wt%.

10. The method according to claim 1, characterized in that: S3 Zhongzheng flotation collector is made of oxidized paraffin soap, polyoxyethylene ether, tall oil, benzylphenol polyoxyethylene ether, octylphenol polyoxyethylene ether in a mass percentage of 20%-30%: 20%-30%: 30%-40%: 10%-20%: 10%-20%, heated and saponified, and the sum of the mass percentages of each component is 100%.

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