US3150957A - Treatment of phosphate rock - Google Patents

Description

WEAK

Sept. 29, 1964 J. E. SEYMOUR ETAL TREATMENT OF PHOSPHATE ROCK Filed June 7, 1961 LOW GAP/10c PHOSP/M 75 ROCK (an; $15.4 '1:

UN REAG'TED ROCK RES/DUE EXTRACT/0N UNDER 46/ T/O/V 05 c4 mm mm 2 0, WA H/N6 I TAIL/IVES FLOTA 7/0N H] GH GRA DE P/MSPHATE RO0K FL um P2 0,- EX 7R4 cr (51 um y) D/V/S/O/V INTO 7W0 PARTS MIXING TR/ LE SUPERPHOSPHATE .570 MOISTURE 4a 49% P 0,

CONCENTRATION fldLTEANATf srsp) ca m we STRON 6 PHOSPIIOAIC' 40/0 (42. a I1, 0,)

INVENTORS JAME E. SEYMOUR ROBERT WHITE ATTORNEY United States Patent 3 Claims. (QR. 71-37) This invention relates to a improved process for the treatment of phosphate rock and to an improved phosphate rock product resulting therefrom.

In most phosphate mining operations, the matrix containing the phosphate rock is excavated from the ground by large draglines and then slurried with water and pumped to a beneficiation area where the matrix is Washed and classified into various sized particles. Generally, the larger pebble size particles are suitable as such for starting material in the production of phosphoric acid or triple superphosphate. The medium size particles contain substantial proportions of silica and other impurities and must be subjected to an ore flotation step of one kind or another in order to concentrate the phosphatic values sufiiciently to be of use in the subsequent operations; Whereas the extremely fine particles such as clays (slimes) and colloidal phosphates and silica have such an adverse efl'ect on subsequent operations that it is found more expedient merely to discard them than to attempt to re cover the phosphate values therein. In conventional operations, unrecovered P 0 is lost, not only in the slirnes but also in the tailings from the ore flotation step, and also at various other points in the overall system. The amount of P 0 thus lost at these various stages totals more than 35% of the P 0 originally present in the matrix, and this is a source of concern to phosphate producers.

Another source of concern is that fact that large deposits of phosphate rock are of such low grade as to make it economically infeasible to mine them. At the present time there is little or no market for concentrated phosphate rock having a grade less than 68% B.P.L., the most desirable grades ranging from 72 to 77% B.P.L.

It is an object of the present invention to provide a process for producing high grade phosphate rock While eliminating a substantial part of the P 0 losses referred to above and at the same time omitting all or part of the preliminary washing, sizing and flotation steps referred to above. It is a further object of the invention to provide a process for upgrading low grade phosphate rock in a manner which is efficient and economical enough to make it feasible now to mine low grade ore. A still further object is to provide a concentrated phosphate rock material which is novel and possesses properties making it particularly suited for use in the production of phosphoric acid and triple superphosphate. Other specific objects and advantages will appear as the specification proceeds.

The present invention is based in part on the discovery that phosphate rock can be upgraded and/or its yield of P 0 improved if the rock is subjected to a combination of process steps-namely, an extraction step followed by a conventional flotation step. in the extraction step, the rock is mixed with a mineral acid, preferably phophoric acid, to cause partial digestion of the rock and thus extract monocalcium phosphate and other soluble and colloidal phosphates from the rock in a fluid form. The undigested rock residue is then subjected to a conventional flotation step to concentrate the phosphatic values and produce a rock high in B.P.L. percentage. It

Patented Sept. 29, 1964 has been discovered that the partial digestion which takes place in the extraction step conditions the unreacted residue and renders it much more amenable to the subsequent flotation process than is the case with normal untreated phosphate rock. This improved efi'iciency of the flotation step is responsible in large part for the economic success of the overall process. As will be described hereinafter, a process which includes this combination of steps may be used to improve the yield of P 0 from standard grade ore; to upgrade low grade material; and to eliminate some or all of the preliminary processing steps.

This invention in its broad aspect, therefore, comprises the step of mixing phosphatic ores or fractions thereof with a mineral acid or acid salt to extract a portion of the phosphatic values present in the starting material, thus placing the non-extracted residue in particularly suitable condition for the ensuing steps. The invention in one of its preferred embodiments comprises the steps of mixing the phosphatic starting material with the: acid to cause partial reaction and extraction or" phosphatic values present in the starting material, separating the fluid por tion of the reaction mixture from the non-solubilized residue, and concentrating the non-solubilized phosphatic values in said residue by a flotation process effective to separate the said phosphatic values from impurities.

The extraction step of the present invention is conveniently carried out by mixing the phosphate rock and the mineral acid, such as phosphoric acid, in a reaction vessel and agitating the mixture until the desired degree of extraction and rock conditioning has taken place. The reaction is subject to the influence of a number of variables, such as type, size and grade of rock, type and concentration of acid, ratio of acid to rock, time of re action, temperature, and the like. Each of these process elements may be varied to a substantial degree Without adversely affecting the operation of the process, provided of course that the other variables are correspondingly changed and controlled in accordance with the description hereinafter set forth. As will be brought out, the overall process of the invention is elfective in producing two products simultaneously-Le, (1) liquid phosphoric acid and (2) high grade rock, which may be used in the production of phosphoric acid or triple superphosphate. By controlling the process variables, it is possible in the practice of the invention to increase the amount of high grade phosphate rock produced while decreasing the amount of phosphoric acid produced, or vice versa, depending upon curren market demands for these two diiierent items of commerce.

In normal operations, it is desirable to control the conditions of reaction in such manner that at least 5% and not more than by weight of the P 0 originally present in the starting material be extracted during the extraction step. Prefer-ably, from 40 to 60% of the P 0 should be extracted.

The phosphate rock to be treated in the extraction step may be any suitable mineral material containing phosphatic values ordinarily used in the production of phosphoric acid, triple superphosphate, and the like. Phosphate rock containing B.P.L. percentages as low as 12, and as high as 75, have been used successfully in the process, and it is apparent that under certain conditions, the process would be applicable to even higher or lower grade rock. The geographical origin of the rock does not appear to be critical, the pebble rock of Florida being utilizable in the process, as well as the phosphate ores from Tennessee, Baja, California, Mexico, and other cations. The preparative history of the rock also is not particularly important, since the process may be used not only with the complete matrix as it is taken out of the ground but also with fractions of the matrix, such as coarse ground pebble, flotation feed (debris), flotation tailings, and the like. The process has particular application to the extensive deposits of the so-called low grade rock which has B.P.L. in the range of 55 to 68% and which are not considered economically recoverable by conventional methods for use as such in the production of phosphoric acid and triple superphosphate.

As far as the particle size of the phosphate rock to be used in the extraction step is concerned, there are no absolute limitations. Generally speaking, anything larger than 150 mesh is preferable. However, it has been shown that much finer materials are also applicable, as evidenced by the fact that the complete phosphate matrix, as mined from the ground, can be successfully treated in the process. The matrix, of course, contains a substantial amount of fines smaller than 150 mesh. At

the other end of the scale, it is not as desirable to use 7 phatic starting materials having a major portion of their particles in the size range l4+60 mesh.

The acid to be used in the extraction step may be any mineral acid which is effective in digesting phosphate rock to extract phosphatic values therefrom. Phosphoric acid is particularly suited for use in the process, and for this purpose, the phosphoric acid may be any convenient grade of acid made by the conventional wetprocess, the electric furnace process, or variations of these. The concentration of the phosphoric acid is one of the process variables which may be utilized in controlling the reaction to provide desired results. Concentrations of to (by weight) P 0 are preferred, but obviously, higher or lower concentrations may be used. In addition to phosphoric acid, other mineral acids, such as hydrochloric, nitric, hydrofluosilicic, and sulfuric, or the acid salts thereof, are also utilizable.

The temperature of the extraction process is not a critical condition. For the most part, ambient temperatures are satisfactory. Generally speaking, and within limits, cooler temperatures extract more phosphatic values than warmer temperatures, because of a dissociation factor involved in the use of the latter. However, this effect of temperature variation is not of substantial importance and can be controlled by regulating other variables in the process if compensation is desired.

The time of reaction is a variable which has considerable effect on the ratio of soluble phosphoric acid to high grade rock coming out at the end of the overall process. Generally speaking, the longer the reaction time, the greater the quantity of soluble phosphoric acid, and the lower the quantity of high grade rock produced by the overall process, and vice versa. Suitable reaction times vary from a matter of a few minutes up to three or four hours. For most applications, a reaction time of one hour should be sufficient.

The ratio of acid to rock in the extraction step is another variable which has a marked effect on the overall results of the process. Using phosphoric as the acid, it has been demonstrated that as much as 65% and as little as 9% (by weight) of the phosphatic values in the rock can be removed in the extraction step merely by varying the ratio of acid to rock in the starting material, all other variables being held constant. For most practical purposes in connection with phosphoric acid, a P 0 (acid) to P 0 (rock) ratio of about 16:1 is suitable. It is preferable to use a ratio of less than 2:1, because it appears that at these lower ratios, there is a selective step. The invention is particularly applicable to phos- H extraction of iron and aluminum phosphates, and accordingly, the result is an upgrading of the unreacted residue which is used in the subsequent flotation step for production of high grade rock. A lower practical limit on the ratio of acid to rock is ofiered by the fact that the lower ratios involve less liquid and, therefore, there may be difliculties in separating the fluid and solid components of the mixture unless water is added, which may not be desirable in all cases.

The mechanics of the extraction step are not particularly critical. The reaction can be conveniently carried out in a vessel or series of vessels, and apparatus such as a conventional screw type classifier is particularly satisfactory. It is desirable that the operation be continuous and that some kind of agitation during the extraction step be used in order to increase contact and to throw the reaction products away from the reactants. It is also possible to use a fluidized bed system for carrying out the reaction, passing and recycling the fluid acid through the bed of rock.

Following completion of the reaction in the extraction step, it is desired to separate the fluid portion of the mixture from the unreacted rock residue. The separation may be aided by a short preliminary settling step. Ordinarily, the length of time for settling is only a few moments, and this may be shortened by the addition of a iiocculating agent such as, for example, a polyelectrolyte, a silica gel (which may be formed in situ) or other similar material. in some cases, the settling step may be eliminated altogether if the subsequent separation step is of a type which corresponds to a classifying overflow system or the like.

Following the settling step, the reaction mixture is treated to separate the fluid portion from the non-solubilized rock residue. This may conveniently be done by merely decanting, such as in a backflow or overflow decantation. Other continuous separation systems are also possible. The fluid portion which is recovered in this fashion contains dissolved P 0 monocalcium phosphate, organic materials, iron and aluminum compounds, and other materials normally found in phosphoric acid produced by the wet process. In addition to the materials which are in the solution, the fluid portion usually contains a quantity of colloidal materials. These colloidal materials may be unreacted clays and iron or aluminum phosphates of silica hydrates, as well as quantities of monocalcium phosphate, organic materials, and the like, to the extent that such latter substances are present in proportions exceeding their solubilities. The separated fluid portion, therefore, comprises a mixture of a true solution with a quantity of colloidal material and is a rather viscous, slurry-like liquor which nevertheless is predominantly liquid and can be handled as such. The colloidal component of the slurry tends to settle slowly, and if desired, the slurry can, therefore, be divided into a clear fraction and an opaque fraction with perhaps an intermediate fraction also taken off. The extract or slurry, either in its complete form, or divided into fractions, has a considerable phosphatic content and is used to produce either triple superphosphate or higher grades of P 0 acid, as

will be described later in this specification.

The non-solubilized solid rock residue which is separated from the reaction mixture of the extraction step still contains substantial phosphatic values and is used to produce a marketable high grade phosphate rock, according to the following procedure:

The residue may first be washed to recover any soluble P 0 values which may have been carried along with the residue. The washing step is a conventional one, used at various other points in recovery plants, and involves merely mixing the residue with a quantity of water to remove the soluble P 0 and then separating the liquid from the solids. The separation can be accomplished by any suitable meanse.g., decantation, vacuum filtration, and the like. The efficiency of the washing step can be improved by the addition of well-known surface active agents. The wash water resulting from this step is a weak P solution which may be recycled and/or used at various other points in the conventional manufacture of phosphoric acid or phosphate fertilizers.

The washed rock residue is then subjected to a conventional flotation procedure to concentrate the phosphatic particles and separate them from silica and other impurities. The residue, having been subjected to special conditioning action of the phosphoric acid in the extraction step, is now in a form which is particularly susceptible to an eflicient flotation procedure. That is, the phosphoric acid, in extracting part of the phosphatic and other values from the rock, has left the rock residue in a highly porous, non-slimed condition, and in this condition, the rock residue particles are extremely more attachable to the flotatlon reagents than is the case with conventional phosphate rock particles. As a result, comparative tests show that close to 93% (by weight) of P 0 content of this conditioned rock can be recovered by the flotation process, as compared to a recovery of only about 82% (by weight). from a comparable but unconditioned phosphate rock.

The flotation process may be any suitable. conventional flotation procedure. Normally, in the practice of this invention, the procedure involves an anionic flotation as the first step, followed by an optional cationic flotation step. The anionic flotation involves the use of a fatty acid tall oil), fuel oil, high pH, and tie-oiling. The subsequent step, if practiced, involves the use of amine flotation reagents. Obviously, other sequences and other reagents can be used. These are standard, conventional procedures which are well-described in the literature (ezg. Wm. H. Waggaman, Phosphoric Acid, Phosphates and Phosphatic Fertilizers, 2nd ed., Reinhold, 1952, p. 57 et seq), and any of these which are eifective in the concentration of normal phosphate rock are also eifective with respect to the conditioned rock of this invention, only more so.

The phosphate rock, produced and concentrated according to the foregoing process, is a high grade phosphate rock material. it is a marketable item of commerce, useful in the production of phosphoric acid or in the production of triple superphosphate. Even starting with low grade rock with a B.P.L. of less than 60%, the rock product prepared by the process of this invention ends up with a BBL. of 72%, and this can be made higher if double flotation or other conventional techniques are used. Moreover, in addition to being equivalent to the conventional 72% BBL. rock of commerce as far as BPLL. content is concerned, the. rock product of the present invention possesses certain different. characteristics which make it even more desirable for its intended purpose. In the new product, a substantial part. of the phosphate content is in the form of. hydroxyl apatite. The fluorine content f the new rock is less than 2.5% by weight compared to 3.5% in natural rock. The new rock has a citrate soluble P 9 content of about 12 to 15% by weight, which is two to three times the amount of citrate soluble P 0 in natural rock. This increased solubility of the P 0 is important in connection with direct application of the rock to soil, the P 0 content of the new type being more readily available to the $011. The new rock contains less than 1.25% by weight of iron and alumina and is substantially free of limestone (carbonate of lime), as compared with natural rock which contains 2 to 3% iron and alumina and 5 to 7% limestone. Since these elements use up sulfuric acid when the rock is reacted with sulfuric acid in the production of phosphoric acid or triple snperphosphate, use of the new rock results in a substantial savings in theamonnt of sulfuric acid required.

Returning now to the point in the process where the fluid portion of the extraction mixture is separated from the unreacted rock residue, the said fluid portion has, as indicated previously, a considerable phosphatic content and is used to produce triple superphosphate or, alternatively, high grade P 0 acid. The production of the triple superphosphate is described as follows:

The fluid portion, or slurry, which is separated from the extraction mixture contains substantial portions of P 0 acid and monocalcium phosphate, together with other substances of minor concentration. The slurry as such may be added to rock dust of 70-75 B.P.L. grade to produce triple superphosphate. In the conventional production of triple superphosphate, P 0 acid itself is added to rock dust. In the new process, the slurry contains not only P 0 but also monocalcium phosphate. This means that the amount of rock dust used in the process thereby can be substantially reduced. Various alternative sequences of steps may be employed at this stage of the triple superphospha-te production:

(1) The mixture of slurry and rock dust may be treated according to a modified version of the process set forth in King Patent No. 1,780,620, wherein ground phosphate rock is mixed with phosphoric acid to form a slurry and the slurry heated and dried simultaneously to form a thick liquid which is discharged into a pit or other container. In the pit, the chemical reaction is completed to a substantial extent and the material sets to a solid mass which is then removed and dried.

(2) The slurry may be reacted with sulfuric acid to remove the calcium as insoluble calcium sulfate (gypsum), and the remaining P 0 liquid acid used in the conventional manner for production of triple superphosphate-e.g., concentration of the P 0 acid in evaporators and mixing of the concentrated acid with rock dust to produce triple superphosphate.

Instead of making triple superphosphate from the fluid portion, or slurry, which is separated from. the extraction mixture, one may alternatively treat the slurry to produce high grade soluble P 0 acid. This is done by reacting the slurry with sulfuric acid so that the calcium content of the slurry is precipitated as insoluble gypsum. The liquid portion is P 0 acid of higher grade, which can be used in the production of triple superphosphate; or can be used as the extracting acid in the process of the present invention; or can be sold as an item of commerce.

It has been found that, in producing P 0 acid from the slurry as in the preceding paragraph, an improved grade of gypsum can be produced if the slurry is allowed to settle and a top liquid layer substantially free of colloidal substances is reacted with the sulfuric acid. The resulting calcium sulfate (gypsum) is substantially free of organics and silica and is suitable for use as a source of high purity calcium sulfate.

The combination of the extraction step and the subsequent flotation step is a unique sequence of steps which provides the basis for improved phosphate rock operations of a very flexible nature. Some of the advantages flowing from the process of this invention. are stated as follows:

(1) Low grade rock, which heretofore had little or no market value, may 'now be treated economically by the new process to produce 72% or better BBL. It is now economically feasible to mine the low grade deposits.

(2) The 72% and better B.P.L. rock product which is obtained as the end product of the present invention is not only equivalent to natural 72% B.P.L. rock in B.P.L. value but possesses all the additional advantageous characteristics previously described in this specification.

(3) The process enables recovery of greater yields of P 0 values from any given grade or type of rock. The unique extraction step not only conditions the rock so as to improve the efliciency of the subsequent flotation step but also in so doing, it extracts soluble phosphatic values which are completely recoverable as high grade P 0 acids or as triple superphosphate ingredients. As compared to overall losses of P 0 amounting to more than 35% by weight in conventional phosphate rock recovery procedures from input matrix, the new process reduces losses down to at least 25% by weight, and in some cases 5 lower. Where the complete matrix is used as the starting materials, the overall P loss may be as low as 5 to by weight. In addition, many of the preliminary washing, classifying, and flotation steps of the conventional operation may be eliminated.

(4) The overall process of the present invention has a considerable degree of flexibility as far as market conditions are concerned. If the market demand is greater for the high grade phosphate rock than it is for the P 0 acid, then the conditions of the extraction step may be varied to reduce the amount of soluble P 0 values extracted from the rock and thus correspondingly increase the unreacted rock residue which are available for producing the desired high grade phosphate rock. If the market demand for the P 0 acid is greater then the reverse procedure is followed.

(5) The specific sequence of steps is eifective in directing fluorine values into the unreacted residue and then into the tailings of the flotation step. To a substantially greater extent than in conventional phosphate rock treatment, the fluorine values are carried out with the sand of the tailings, where they do not cause any further problems with respect to product contamination or product insolubilization.

Phosphate rock and phosphoric acid operations, based on the present invention, can take the form of several different systems or circuits. One example of this is illustrated by reference to FIG. 1 of the drawing, which in turn is explained in Example I below.

The following examples are intended to illustrate the invention in several of its aspects and are not intended to be construed as limiting thereof.

Example I This is an example showing the treatment of low grade phosphate rock (B.P.L. 58.4) to produce a marketable grade phosphate rock (B.P.L. 72), while at the same time utilizing part of the fluid extract to produce strong phosphoric acid (42.6% P 0 and the other part to produce triple superphosphate (48-49 P 0 This example should be read in connection with FIG. 1 of the drawing.

One hundred parts by weight of 58.4 B.P.L. phosphate rock of particle size -16+35 mesh was intimately mixed and reacted with 251.7 parts by weight of phosphoric acid containing 56.8 parts by weight of P 0 The reaction time was one hour, and reaction temperature was 200 F.

The phosphoric acid used in the reaction analyzed as follows:

#LA-l-5B: Percent P 0 26.7 Moisture 53.6 Fe O 0.436 A1 0 0.51 CaO 1.38 S0 1.235 F 1.91

The phosphate rock used in the reaction analyzed as follows:

Mesh, -16+35.

Subsequent separation of the reaction products by decantation and washing of the residual solid phase, followed by the alternate concentration step, gave slurry and residue products with the following analyses.

Slurry analysis #LA-1-5C): Percent Moisture 21.45 Total P 0 40.15 Citrate insoluble P 0 0.15 Available P205 Acid insoluble 1.10

F6 0 Al O 1.285 CaO 8.78 so. 1.66 F 2.70 Free acid as H to methyl orange 14.34

Residue rock analysis (#LA-1-5D): Percent Moisture 0.30 Total P205 Citrate insoluble P 0 18.32 Available P 0 5.68 Acid insoluble 35.00 F6203 0.40s A1 0 0.588 CaO 34.22 F 3.13

Alkaline as H SO value to methyl orange- 0.32

In the extraction step, approximately 41.9% by weight of the P 0 had been extracted from the phosphate rock.

The residue rock (#LA-l-SD) from the extraction step was washed and then subjected to a single anionic flotation step, involving the use of tall oil, fuel oil, high pH, and de-oiling. The flotation of the washed residue rock resulted in recovery of 92% of the P 0 flotation feed input as 72 B.P.L. The flotation step gave concentrate and tailings analyses as follows:

One hundred parts by weight of a portion of the slurry (#LA-l-SC) from the extraction step and 16.4 parts by weight of the beneficiated residue (concentrate #LA-l-SDA, having a 35 mesh size) were .intimately reacted at ambient temperature and allowed to stand. Analysis of the reaction mixture four days after manufacture was as follows:

Percent Moisture 16.59 Total P305 Available P 0 42.68 H O soluble P 0 41.40 Acid insoluble 2.62 F203 A1 0 0.822 CaO 15.16 80.; 1.7 2 F 3.00 CO 0.00

Reduction of free water to a 5.0% level by drying and reduction of citrate insoluble P by curing yielded a triple superphosphate product containing approximately 4849% available P 0 Two hundred parts by weight of a portion of the slurry (#LA1-5C) from the extraction step was intimately reacted with 32.4 parts by weight of 92.3% H 50 solution. The reaction mixture was filtered and washed immediately. The strong phosphoric acid produced analyzed 42.60% P 0 The ratio of the concentrated phosphoric acid plus Weak phosphoric acid (water wash of Cit-S04 cake) to the total P 0 value was 0.991.

Example H This is an example showing extraction of P 0 from 44.8 B.P.L. phosphate and subsequent beneficiation of residue by conventional anionic flotation procedure to obtain a phosphate rock having a B.P.L. of 72.

One hundred parts by weight of 44.8 B.P.L. phosphate rock of particle size 35+ 60 mesh was reacted with 206 parts by weight of crude wet process phosphoric acid containing 56.8 parts by Weight of P 0 Reaction time was one hour, and reaction temperature was 220 F. The ratio of P 0 (acid) to P 0 (rock) was 2.77.

Phosphate rock analysis (#LA12A): Percent lvloisture 0.218 Total P 0 20.50 Acid insoluble (silica) 37.40 F6203 1.085 A1 0 2.243 F 3.33

Subsequent separation of the reaction products by decantation and washing of the residual solid phase, followed, by the alternate concentration step, gave slurry and residue products with the following analyses.

Slurry analysis (#LA12C): Percent Moisture 17.53 Total P 0 44.65 Citrate insoluble P 0 0.17 Available P 0 44.48 Acid insoluble (silica) -a 090 P5203 A1 0 1.99 F 2.65 Free acidity as H PO 29.28

Residue (#LA-1-2D):

Moisture 0.36 Total P205 Citrate insoluble P 0 12.92 Available P 0 4.08 Acid insoluble (silica) 52.10 PC1203 0.286 A1 0 0.463 F 2.30

Approximately 41% of the total P 0 content of the phosphate rock was extracted.

Subsequent conventional single anionic flotation 9f the residue resulted in recovery in the concentrate of 85% of the total P 0 input in the flotation feed. Overall phosphate (P 0 recovery from the 44.8 B.P.L. phosphate rock Was 91.2% of the total P 0 input.

Example III Extraction of P 0 from 75.5 B.P.L. phosphate rock and subsequent beneficiation of residue by conventional double flotation methods.

One hundred parts by weight of 75.5 B.P.L. phosphate rock and 185 parts by weight of crude Wet process phosphoric acid containing 52.3 parts by Weight of P 0 were mixed and reacted for one hour at 78 F. The ratio of P 0 (acid) to P 0 (rock) was 1.51.

Phosphate rock (#LC 1-7A): Percent Moisture 1.1 1 Total P 0 34.55 Acid insoluble (silica) 6.21 Fe O 0.43 A1 0 0.74 CO 2.71 CaO 4 6.13 S0 0.66

Subsequent separation of the reaction products by decantation and washing the residual phosphatic solid phase gave slurry and residue products of the following analyses.

Slurry (#LC17C): Percent Moisture 46.39 Total P 0 29.35 Available P 0 29.25 Water soluble P 0 29.25 Acid insoluble (silica) 0.43 Free acidity asv B 1 0 23.2

Residue (#LC-1-7E):

Moisture 0.85 Total P205 Acid insoluble (silica) u 7.9

25.2% of the P 0 content of the phosphate rock was extracted.

Subsequent double flotation by conventional methods of residue yields the followingproducts.

Single flotation concentrate (#AA-LC-17EA): 35%, total P205.

The P 0 recovery (represented by the P 0 content of the single flotation concentrate) is equivalent to 83.3% of the total P 0 content of the flotation feed.

The overall P 0 recovery by the extraction and flotation method was 87.5% of the total P 0 content of the 75.5 B.P.L. rock phosphate used.

Example IV Extraction of P 0 from matrix with beneficiation by conventional flotation of residue.

One hundred parts by weight of Wet untreated phosphate matrix was reacted With 167.6 parts by weight of crude Wet process phosphoric acid containing 49.1 parts by weight of P 0 Reaction time was one hour, and reaction temperature was ambient temperature (70 F.). The ratio of P 0 (acid) to P 0 from matrix was 2.41.

Matrix analysis (#LC-l-SA): Percent Total P 0 19.70 Citrate insoluble P 0 17.55 Available P 0 2.15 Acid insoluble a- 3258 Subsequent separation of the reaction products by decantation and Washing of the residual phosphatic solid phase gave slurry and residue products with the following analyses.

Acid insoluble 46.22 Free acid as H PO 0.00

Approximately 34.5% of the matrix total P 0 content was extracted by this method.

Subsequent conventional single anionic phosphatic flotation of the residue yields a phosphatic concentrate of the following analysis.

Single flotation concentrate: 30.05% total P The concentrate P O yield represents 74.9% of the total P 0 content of the flotation feed.

Reverse flotation of the single flotation concentrate yields a concentrate of the following analysis.

#AALC18DC: Percent Total P 0 34.85 Combined tail (#AALC18DB):

Total P 0 1.6

The double flotation concentrate P 0 yield represents 87.1% of the total P 0 content of the flotation feed.

The overall P 0 recovery from the matrix is equivalent to 91.6% of the P 0 content of the matrix.

Example V This example describes a comparative test which demonstrates the improved flotation results which are achieved by subjecting phosphate rock to the preliminary extraction step prior to flotation.

The rock phosphate ore used in this example was 58.4 B.P.L. grade material and had the following analysis.

#LA-l-SA: Percent Moisture 0.342 Total P205 Acid insoluble 21.65 Fe O 1.195 5 5'53 2 I CaO 39.06 F 3.10

The above 58.4 B.P.L. phosphate rock was divided into two partsPart A and Part B.

Part A was beneficiated by single anionic flotation. The concentrate #AA-LA-SAA was 70.7 B.P.L. rock and contained 83.2% of the P 0 in the flotation feed input.

Part B was first extracted with phosphoric acid, as described in Example I, and the rock residue #LA-l-SD was beneficiated by single anionic flotation, using the same conditions which were used in connection with Part A. The concentrate #AALA15DA was 72% B.P.L. rock and contained 92% of the P 0 in the flotation input.

As a further comparison, the total overall recovery of P 0 from the Part B rock (including the P 0 in the beneficiated rock #AALA15DA and the P 0 in the liquid slurry coming out of the extraction step) was 95%, with the rock portion being 72 B.P.L. This cornpares .withan overall recovery of P 0 from Part A of only 83.2%, with a B.P.L. of 70.7.

f .While in the foregoing specification we have set forth 5 examples in which steps or processes are described in considerable detail, it will be understood that such details may be varied widely by those skilled in the art without departing from the spirit of our invention,

We claim:

1. In a process for treating low grade phosphate rock containing B.P.L. values of less than 68 percent to pro duce an improved type of phosphate rock having increased B.P.L. values above said 68 percent, the steps of mixing said low grade rock in particulate form with a rock digesting agent selected from the group consisting of phosphoric acid, sulfuric acid, hydrochloric acid, nitric acid and hydrofluosilicic acid, and the acid salts thereof, to cause partial digestion of the rock so as to extract monocalcium phosphate and other soluble and colloidal phosphates from the rock in fluid form and until 5 to 70 percent of the phosphatic values have been extracted, thus placing the undigested rock residue in a porous, nonslimed condition, separating the fluid portion containing the extracted phosphatic values from the non-solubilized rock residue, and thereafter treating said rock residue in a froth flotation step to concentrate the phosphatic values therein, and then recovering a phosphate rock having a BBL. value above 68 percent.

2. A process of treating low grade phosphate rock containing BEL. values of less than 68 percent to produce a phosphate rock product having increased B.P.L. values above 68 percent, comprising the steps of mixing said low grade rock with a quantity of phosphoric acid to cause partial digestion of the rock and thus extract monocalcium phosphate and other soluble and colloidal phosphates from the rock and until 40 to percent of the phosphatic values present in said low grade phosphate rock is extracted, whereby the non'extracted residue is placed in a porous, non-slimed condition, Washing said rock with water to remove phosphoric acid, subjecting said washed residue rock to froth flotation to concentrate the phosphatic particles therein, and then recovering a higher grade phosphate rock having a B.P.L. value above 68 percent.

3. In a process for treating low grade phosphate rock containing B.P.L. values of less than 68 percent to produce a phosphate rock product containing increased BBL. values above said 68 percent, the steps of mixing said low grade rock in particulate form, mainly in the size range 14+60 mesh, with phosphoric acid in the proportion of about two parts phosphoric acid to one part rock to extract from 40 to 60 percent of the phosphatic values present in said rock, thus placing the nonextracted residue in a porous condition, washing said rock with water to remove phosphoric acid, subjecting said washed residue rock to froth flotation to concentrate the phosphatic values therein, and then recovering a rock residue having a E.P.L. value above 68 percent.

References Cited in the file of this patent UNI ED STATES PATENTS 1,475,959 Meyers 'Dec. 4, 1923 1,547,732 Broadbridge et a1 July 28, 1925 FOREIGN PATENTS 852,538 Great Britain Oct. 26, 1960

Claims (1)

1. IN A PROCESS FOR TREATING LOW GRADE PHOSPHATE ROCK CONTAINING B.P.L. VALUES OF LESS THAN 68 PERCENT TO PRODUCE AN IMPROVED TYPE OF PHSOPHATE ROCK HAVING INCREASED B.P.L. VALUES ABOVE SAID 68 PERCENT, THE STEPS OF MIXING SAID LOW GRADE ROCK IN PARTICULATE FORM WITH A ROCKDIGESTING AGENT SELECTED FROM THE GROUP CONSISTING OF PHOSPHORIC ACID, SULFURIC ACID, HYDROCHLORIC ACID, NITRIC ACID AND HYDROFLUOSILICIC ACID, AND THE ACID SATLS THEREOF, TO CAUSE PARTIAL DIGESTION OF THE ROCK SO AS TO EXTRACT MONOCALCIUM PHOSPHATE AND OTHER SOLUBLE AND COLLOIDAL PHOSPHATES FROM THE ROCK IN FLUID FORM AND UNTIL 5 TO 70 PERCENT OF THE PHOSPHATIC VALUES HAVE BEEN EXTRACTED, THUS PLACING THE UNDIGESTED ROCK RESIDUE IN A POROUS, NONSLIMED CONDITION, SEPARATING THE FLUID PORTION CONTAINING THE EXTRACTED PHOSPHATIC VALUES FROM THE NON-SOLUBILIZED ROCK RESIDUE, AND THEREAFTER TREATING SAID ROCK RESIDUE IN A FROTH FLOTATION STEP TO CONCENTRATE THE PHOSPHATIC VALUES THEREIN, AND THEN RECOVERING A PHOSPHATE ROCK HAVING A B.P.L. VALUE ABOVE 68 PERCENT.

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