US3462017A

US3462017A - Phosphate flotation process

Info

Publication number: US3462017A
Application number: US663491A
Authority: US
Inventors: Charles Herbert George Bushell; Horst Eberhard Hirsch
Original assignee: Teck Metals Ltd
Current assignee: Teck Metals Ltd
Priority date: 1966-12-29
Filing date: 1967-08-28
Publication date: 1969-08-19
Anticipated expiration: 1986-08-19

Description

United States Patent O 3,462,017 PHOSPHATE FLOTATION PROCESS Charles Herbert George Bushell, Montrose, British Columbia, and Horst Eberhard Hirsch, Trail, British Columbia, Canada, assignors to Cominco Ltd., Montreal, Quebec, Canada, a corporation of Canada No Drawing. Filed Aug. 28, 1967, Ser. No. 663,491 Claims priority, applicgtigon ganada, Dec. 29, 1966,

Int. Cl. B03d 1/02 U.S. Cl. 209-166 Claims ABSTRACT OF THE DISCLOSURE This invention relates to the process for beneficiating phosphate-bearing materials and is particularly directed to a flotation process for the concentration of phosphate minerals in phosphate rock high in silica, calcium carbonate and magnesium carbonate.

It is known to upgrade phosphate values in a phosphate containing ore by means of flotation. One generally accepted process comprises suspending finely ground phosphate rock in a slightly alkaline aqueous solution containing a carboxylic acid or other anionic flotation reagent and subjecting the suspension to a first froth flotation step in which the phosphate values are concentrated and collected in the froth and a substantial portion of the siliceous gangue material is depressed and separated as the underflow tailing product. The concentrate from this first flotation step is then treated with sulphuric acid to remove the negative-ion collector reagent, filtered and washed with water to remove the residual sulphuric acid. The concentrate is then suspended in a neutral or substantially neutral aqueous solution containing an amine or other cationic flotation reagent and subjected to a froth flotation step in which the residual silica is collected in the froth and an enriched phosphate-containing concentrate is recovered as the underflow product.

A disadvantage of this process is the necessity for maintaining two separate and distinct flotation circuits, the first anionic and the second cationic, with the attendant problem of removing the flotation reagents from the concentrate from the anionic circuit before commencing treatment in the cationic circuit.

Another known process is that disclosed in U.S. Patent No. 3,113,838 issued Dec. 10, 1963. This patent describes a flotation process which employs a fatty acid in the presence of soluble alkaline phosphate produced by dissolving phosphate salt or by attacking a portion of the phosphate rock with a mineral acid such as sulphuric acid until phosphoric acid is freed, to transform calcium phosphate into a soluble phosphate.

This acid addition is impractical when treating phosphate rock containing substantial amounts of silica as Well as calcium and magnesium carbonates. In the presence of soluble phosphates there is no effective separation of phosphate minerals from silica and silicates. Also, treatment of a carbonate-containing ore with a strong mineral acid. such as sulphuric acid, would require the addition of 3,462,017 Patented Aug. 19, 1969 ice an excessive amount of acid due to the preferential attack on the carbonates, and would unfavorably alter the characteristics of the feed to the flotation operation.

An application of the process described in the foregoing patent to phosphate rock known as collophane high in silica and calcium and magnesium carbonates resulted in substantially no improvement in the phosphate grade. In a typical example using a feed containing 19.5% P 0 and 1.6% MgO, the concentrate assay was 21.8% P 0 and 0.7% MgO. While there was some reduction in the MgO grade, there was no significant improvement in P 0 grade. It is thus evident that the process described in Patent No. 3,113,838 is not suitable for beneficiating phosphate rock containing appreciable quantities of silica.

It is important for the production of a phosphate concentrate suitable for use in the wet phosphoric acid process that carbonate-containing minerals, especially magnesium-containing carbonates, be maintained at a minimum. The presence of carbonates not only causes undesirable foaming in the reactor chambers and undue consumption of acids, but also the magnesium content in the carbonates forms a magnesium sulphate which is soluble in phosphoric acid producing a magnesium salt upon subsequent ammoniation of the phosphoric acid which lowers the value of the ammonium phosphate fertilizer product. Also, the presence of magnesium in the phosphoric acid raises the viscosity of the acid.

We have found that phosphate rock high in silica and magnesium carbonate can be beneficiated by subjecting the comminuted rock to a two-stage anionic flotation utilizing the same collecting reagent or reagents in each stage wherein the bulk of the silica is removed as an underflow and the phosphate bearing minerals and calcium and magnesium carbonates report as float in the first stage, and the phosphate-bearing minerals are depressed selectively for removal as an underflow concentrate and the calcium and magnesium carbonates continuously report as float in the second stage, by the addition of a soluble alkali phosphate to the flotation circuit prior to the second stage flotation. We have discovered, surprisingly, that the presence in the second stage flotation of phosphate ions such as produced by alkali phosphates selected from the group of ammonium phosphate, sodium phosphate and potassium phosphate selectively de-activates the residual anionic collector on the surface of the phosphate particles and depresses the said phosphate particles while the calcium and magnesium carbonates continue to report in the float.

It is a principal object of the present invention, therefore, to provide a method of beneficiating comminuted phosphate rock high in silica and magnesium-containing carbonates to produce a concentrate rich in phosphate minerals substantially free from the aforesaid undesirable gangue minerals.

It is another object of the present invention to provide a process for beneficiating phosphate rock by a two-stage anionic flotation process utilizing common anionic collectors.

And another object of the invention is the provision of a process which permits utilization, and recovery for reuse, of a readily available conditioning agent.

These and other objects of the invention and the manner in which they can be attained will become apparent from the following detailed description of the embodiments and examples of the process.

According to the process of the invention, prior to beneficiation, phosphate rocks containing minerals such as collophane, are first comminuted to a size range to substantially liberate the phosphate-bearing minerals from the gangue minerals. The degree of crushing and grinding applied to the rock will depend, of course, on the texture of the rock and association of the phosphate values with the gangue minerals. The crushed and ground rock is then classified to yield a sized product which is suspended in water or recycled solution and conditioned by the addition of carboxylic acid, sodium silicate as required and an alkali for a pH of 9 to 10.5. The conditioned material is subjected to a first stage anionic flotation using fatty acid collector reagents such as oleic acid, or other carboxylic acids, including tall oils and the like to provide an enriched phosphate concentrate containing calcium and magnesium carbonates as a float product and a silica tailing as an underflow depressed product. The primary concentration of phosphate minerals from siliceous gangue minerals as described above is well known in the art.

The phosphate-enriched float concentrate is then thickened and passed into a conditioner having an aqueous solution therein of soluble phosphate salts maintained at a suitable pH by the controlled addition of acid or base. We have found that optimum results are attained with a pH within the range of from about 4.5 to about 7.0, preferably within the range of from about 5.5 to about 6.5. The optimum pH and alkali phosphate concentration are primarily determined by the nature of the ore.

Alkali phosphates present in amount of from about 0.02 gram per litre (g.p.l.) to about 20 g.p.l. are generally effective in conditioning and depressing the phosphate hearing minerals. The preferred range of phosphate salt concentration is from about 0.2 g.p.l. to about 10 g.p.l., however, concentrations below about 0.2 g.p.l. result in depression of certain phosphate minerals only and are not universally effective for all phosphate rock. Concentrations above about the 10 g.p.l. level, although effective up to g.p.l. for both sodium phosphate and ammonium phosphate, do not result in any significant improvement in phosphate mineral recovery. The phosphate salts can be selected from the group of salts consisting of ammonium phosphate, sodium phosphate, potassium phosphate and the like water-soluble phosphates having cations which will not interfere with the collector reagent such as by reacting with the collector to form insoluble salts. Primary, secondary and tertiary phosphate salts can be employed.

It is not fully understood why the solution pH and phosphate ion concentration are critical in depressing the phosphate particles while permitting the carbonates to continue to report in the float of the second stage flotation under the influence of the same fatty acid collectors employed to collect both phosphate and carbonate particles in the first stage flotation. While it will be understood that the present invention is free from hypothetical consideraitons, it is believed the addition of soluble phosphate salts provides PO ions which, while within the preferred pH range, produce an optimum concentration of H P ions according to the following reaction:

The H PO ions are believed to de-activate eflectively the residual anionic collector on the surface of the phosphate particles while not affecting the activated surfaces of the calcium carbonate and magnesium carbonate particles.

The conditioned slurry, to which may be added a flotation reagent such as oleic acid, is then passed to a second stage anionic flotation for production of an overflow product of gangue material, especially calcium carbonate and magnesium carbonate minerals, and an underflow product of phosphate-enriched material which is thickened and filtered for the production of a phosphate concentrate. The float product from the second stage flotation normally is also thickened and the solutions recovered from the thickener and filter units consolidated for discharge to waste or for recycle to the conditioning step for utilization of the conditioning agents contained therein.

The following examples show the effectiveness of the present integrated process in the treatment of phosphate rock known as collophane indigenous to the Western United States.

EXAMPLE 1 In this example of the process of our invention a phosphate rock was ground to 60-80% minus 200 mesh and suspended in an aqueous solution having a pH of 9 to 10.5 and containing the following reagents:

Lbs. per ton of ore Fatty acid 1.5 Sodium hydroxide 3.0 Sodium silicate 1.0

The liquids-solids ratio of this suspension was adjusted to 30 to 35% solids by weight and the suspension was fed to a flotation machine.

The concentrate from the first stage separation was pumped to a thickener and the thickener underflow containing approximately solids was pumped to a conditioning tank and the following reagents were added:

Ammonium phosphate (11-400 fertilizer grade monoammonium phosphate (NH )H PO 12 lbs. per ton of first stage concentrate producing an ammonium phosphate concentration of 3 g.p.l. and solution pH of 5.8 to 6.0 after the suspension had been diluted to 33% solids by weight.

Tall oil (fatty acid), 2.28 lbs. per ton of first stage concentrate.

This suspension was then diluted with water to 30 to 35% solids by weight and fed to a flotation machine.

Table 1 below gives the analysis and percent distribution of the P 0 MgO, CaO, Fe, A1 0 and SiO contained in the feed, concentrate and tailing produced in the first stage separation, indicating ignition losses (I.L.).

Table 2 gives the corresponding analysis and percent distribution of the feed to the second stage separation and of the concentrate and tailing produced.

Table 3 provides a tabulation of overall recoveries of the process.

In this and subsequent examples, first stage concentrate can be diluted with return solution from the second stage separation to achieve a corresponding decrease in phosphate salt added to maintain the desirable phosphate concentration in solution.

Composition (percent) Distribution (percent) P 05 MgO CaO Fe A1 0; 810: I.L. Wt. P205 MgO CaO Fe A120 SiOz I.L. Feed 19. 1 1. 2 29. 8 1. 6 6. 4 31. 0 4. 4 100 100 100 100 100 100 Concentrate (Overflow) 28. 5 1. 68 43. 2 0.9 2. 1 11. 9 7. 0 58. 2 86. 9 81. 7 84. 4 32. 5 19. 1 22. 4 g Ta11mgs (Underflow) 6. 0 0. 53 11. 1 3. 6 12. 4 57. 6 0.8 41. 8 13. 1 18. 3 15. 6 67. 6 80. 9 77. 6 7. 5

TABLE 2.SECOND STAGE Composition (percent) Distribution (percent) P 0 MgO CaO Fe A S10; LL. Wt. P20 MgO CaO Fe A1 0 S102 I.L. Feed 28. 5 1. 68 43. 2 0. 9 2. 1 11. 9 7. 0 100 100 100 100 100 100 100 100 Tallmgs (Overflow) 9. 0 7. 5 38. 0 1. 5 2. 2 9. 3 18. 9 14. 6 9. 7 64. 8 12.8 24. 0 15. 7 11. 4 39. 5 Concentrate (Underflow). 30. 1 0. 7 44. 1 0. 8 2. 0 12.3 5. 0 85. 4 90. 3 35. 2 87. 2 76. 0 84. 3 88. 6 60. 5

TABLE 3.PERCENT OVERALL RECOVERIES P205 78.5 MgO 28.8 CaO 73.6 Fe 24.7 A1203 16.1 s10, 19:8 LL. 56.0 Wt. 49.7

It will be noted from the above tables that the overall recovery of the P was 78.5% and the overall reduction of MgO was (100%-28.8%)=71.2%.

EXAMPLE 2 In this example of the process a collophane rock comminuted as described above was subjected to anionic flotation according to the conditions set forth in Example 1 and transferred to a conditioning tank to which the following reagents were added:

Ammonium phosphate, 40 lbs. per ton of first stage concentrate to produce an ammonium phosphate concentration of 1 0 g.p.l. and solution pH of 5.6 after diluting the suspension to 33% solids by weight.

Alkylarylsulphonate, cumulate lbs. per ton in each of four floats: Float 1-spontaneous, Float 20.023, Float 30.07, Float 4-0.16.

Tables 4, 5 and 6 below provide data for the first stage separation, second stage separation and overall recoveries of the P 0 and MgO.

TABLE 4.-FIRs'1 STAGE Composition (percent) TABLE 7.FIRST STAGE Composition Distribution (percent) (percent) P206 MgO Wt. P 0 MgO Fee 20. 0 0. 94 100 100 100 1. 33 57. 6 87. 5 81. 5 0. 41 42. 4 12. 5 18. 5

TABLE 8.-SECOND STAGE Distribution Composition (percent) (floats (percent) accumulative) P 0 MgO Wt. P 0 Mg() 30. 4 1. 33 100 100 100 8. 6 13. 0 6. 3 1. 8 61. 6 17. 9 5. 2 10. 0 4. 0 76. 0 F at 3 20. 5 3.0 15. 0 7. 3 87. 3 Concentrate (Underflow) 33. 1 0. 2 85. 0 92. 7 12. 7

TABLE 9.PERCENT OVERALL RECOVERY P 0 81.1 MgO 10.4 Wt. 49.0

It will be observed from Table 6 that the overall recovery of P 0 was 81.1% and the overall reduction of MgO was (10010.4%)=89.6%.

The present invention provides a number of important advantages. Phosphate rock high in silica and calcium and magnesium carbonates can be beneficiated by removal Distribution (percent) P205 MgO LL. Wt. P205 MgO LL.

Feed 20.1 0. 7 3. 6 100 100 100 100 Concentrate (Overflow) 30. 2 0. 86 5. 3 57. 7 86.7 70. 9 85.0 6. 3 0. 48 1. 3 42. 3 13. 3 29. 1 15. 0

TABLE 5-SECOND STA GE Distribution (percent) Composition (percent) (floats accumulative) P105 MgO I.L. Wt. P205 MgO LL TABLE 6 PERCENT OVERALL RECOVERY of the silica and carbonates, especially the magnesiump o 8 ,7 50 containing ca b nates, in an integrated two-stage anionic uflnu 7 fl tation circuit by the simple expedient of conditioning IL the phosphate concentrate from the first stage in an w I 5 aq eous solution containing soluble alkali phosphates to de-activate anionic collectors residual on the phosphate It will be noted that from Table 6 the overall recovery of the P 0 was 81.7% and the overall reduction of MgO was (100% 29.7% )=70.3%.

EXAMPLE 3 In this example a collophane rock was comminuted for liberation of phosphate-bearing minerals and provided with anionic flotation according to the conditions set forth in Example 1. The concentrate from the first stage flotation was transferred to a conditioning tank in which the following reagents were added:

Sodium phosphate, 40 lbs. per ton of first stage concentrate to produce a sodium phosphate concentration of 10 g.p.l. and solution pH of 6.2 after diluting the suspension to 33% solids by weight.

Fatty acid, cumulative lbs. per ton in each of three floats: Float l-spontaneous, Float 2-0.25, Float 3-- 0.90.

Tables 7, 8 and 9 below provide data for the first stage separation, second stage separation and overall recoveries of the P 0 and MgO.

minerals. The reagents from the first stage can often provide the necessary concentration of conditioning and collector reagents necesssary to effect the subsequent flotation operation of the gangue minerals in the second stage.

What we claim as new and desire to protect by Letters Patent of the United States is:

1. A process for beneficiating phosphate rock high in silica and magnesium carbonate and containing calcium carbonate comminuted for the substantial liberation of phosphate minerals from gangue by a two-stage froth flotation which comprises the steps of subjecting said comminuted phosphate rock to a first stage anionic flotation for separately withdrawing a phosphate enriched float product containing calcium carbonate and magnesium carbonate and depressing silica for discharge as underflow tailings, conditioning said float product in an aqueous solution of an alkali phosphate for de-activating the phosphate minerals, and subjecting said conditions float product to a second stage anionic flotation for separately withdrawing a float product high in calcium carbonate and magnesium carbonate and an underflow concentrate high in phosphate minerals.

2. In a process as claimed in claim 1, said alkali phosphate being present in amount of from about 0.02 gram per litre to about 20 grams per litre.

3. In a process as claimed in claim 2, said alkali phosphate being selected from the group consisting of ammonium phosphate, sodium phosphate and potassium phosphate.

4. In a process as claimed in claim 1, said alkali phosphate being present in amount of from about 0.2 gram per litre to about 10 grams per litre.

5. In a process as claimed in claim 1, maintaining the pH of the aqueous solution in the second stage flotation in the range of from about 4.5 to about 7.0.

6. In a process as claimed in claim 2, maintaining the pH of the aqueous solution in the second stage flotation in the range of from about 4.5 to about 7.0.

7. In a process as claimed in claim 3, maintaining the pH of the aqueous solution in the second stage flotation in the range of from about 4.5 to about 7.0.

8. In a process as claimed in claim 1, maintaining the pH of the aqueous solution in the second stage flotation in the range of from about 5.5 to about 6.5.

9. In a process as claimed in claim 2, maintaining the pH of the aqueous solution in the second stage flotation in the range of from about 5.5 to about 6.5.

10. In a process as claimed in claim 3, maintaining the pH of the aqueous solution in the second stage flotation in the range of from about 5.5 to about 6.5.

References Cited UNITED STATES PATENTS 2,442,455 6/1948 Booth 209166 2,959,281 11/1960 Faucher 209-166 X 3,113,838 12/1963 Perri 209166 X 3,259,242 7/ 1966 Snow 209-466 HARRY B. THORNTON, Primary Examiner ROBERT HALPER, Assistant Examiner