METHOD FOR THE REDUCTION OF HALITE IN THE PREPARATION OF POTASSIUM SULPHATE FROM POTASSIUM-CONTAINING ORES AT HIGH AMBIENT TEMPERATURES

Abstract

There are provided methods for the production of potassium sulphate. The methods comprise contacting an aqueous potassium- and sulphate-containing composition with magnesium chloride (MgCl.sub.2), thereby obtaining a composition comprising kainite; optionally concentrating the kainite from the composition and reducing or removing halite therefrom; reacting the kainite with magnesium sulphate (MgSO.sub.4) and potassium sulphate (K.sub.2SO.sub.4) so as to convert the kainite into leonite (K.sub.2SO.sub.4.MgSO.sub.4.4H.sub.2O); optionally contacting the leonite with water to remove excess MgSO.sub.4; and contacting the leonite with water so as to leach the MgSO.sub.4, contained in the leonite, and to at least substantially selectively precipitate potassium sulphate (K.sub.2SO.sub.4), further involving a process brine sulphate control step, based on bloedite precipitation, to control the overall level of sulphate in the method and further comprising a step for the substantially complete reduction or removal of halite from the flotation concentrate, accompanied by an additional precipitation of kainite, thus also increasing the overall recovery of kainite in the process. The method according to the invention can be operated at higher temperatures, in particular at temperatures above 35 C. and does not require a cooling step at 20 to 25 C. The method produces potassium sulphate with a low amount of chloride.

Claims

1. A method for the production of potassium sulphate comprising the steps of: Ia) contacting an aqueous potassium- and sulphate-containing composition with magnesium chloride (MgCl.sub.2), thereby obtaining a composition that, upon evaporation of the water, produces solids comprising kainite (KCl.MgSO.sub.4.2.75 H.sub.2O); IIa) concentrating and separating the kainite from the composition, obtained in step Ia, by flotation, thereby producing a flotation concentrate comprising the kainite and a rest composition (flotation tailings); IIIa) reacting the kainite, obtained in step IIa, with water, optionally comprising magnesium sulphate (MgSO.sub.4) and potassium sulphate (K.sub.2SO.sub.4), at a temperature of about 35 C. to about 70 C., so as to convert the kainite into leonite (K.sub.2SO.sub.4.MgSO.sub.4.4H.sub.2O) and separating the leonite thereof, thereby producing a rest composition (conversion liquid); IVa) optionally, contacting the leonite, obtained in step IIIa, with water to remove remaining solid MgSO.sub.4 compounds; Va) contacting the leonite, obtained in step IIIa or IVa, with water so as to dissolve leonite and/or leach the MgSO.sub.4, contained in the leonite, and to at least substantially selectively crystallize potassium sulphate (K.sub.2SO.sub.4); and VIa) combining at least part of the balance composition (flotation tailings) from step IIa with at least part of the balance composition (conversion liquid) from step IIIa and optionally with water, to precipitate bloedite; characterized in that the method contains a further step IIa, situated between step IIa and IIIa, comprising contacting the flotation concentrate of step IIa with an aqueous solution of MgCl.sub.2.

2. A method for the production of potassium sulphate, comprising the steps of: Ib) contacting an aqueous potassium and sulphate-containing composition, further comprising sodium chloride, with magnesium chloride (MgCl.sub.2), thereby precipitating halite (NaCl) and obtaining a composition that, upon evaporation, produces solids comprising kainite (KCl.MgSO.sub.4.2.75 H.sub.2O); IIb) concentrating and separating the kainite from the composition, obtained in step Ib by flotation, thereby producing a flotation concentrate comprising the kainite and a rest composition (flotation tailings) and controlling the concentration of sodium chloride, present in the composition comprising kainite so as to maintain the concentration of sodium chloride below about 10% by weight on dry matter basis; IIIb) reacting the kainite, obtained in step IIb with water, optionally comprising magnesium sulphate (MgSO.sub.4) and potassium sulphate (K.sub.2SO.sub.4), at a temperature of about 35 C. to about 70 C., so as to convert the kainite into leonite (K.sub.2SO.sub.4.MgSO.sub.4.4H.sub.2O) and separating the leonite thereof, thereby producing a rest composition (conversion liquid); IVb) optionally, contacting the leonite, obtained in step IIIb, with water to remove any remaining solid MgSO.sub.4 compounds; and Vb) contacting the leonite, obtained in step IIIb or IVb, with water so as to dissolve leonite and/or leach the MgSO.sub.4, contained in the leonite, and to at least substantially selectively crystallize potassium sulphate (K.sub.2SO.sub.4); VIb) combining at least part of the balance composition (flotation tailings) from step IIb with at least part of the balance composition (conversion liquid) from step IIIb and optionally water, to precipitate bloedite; characterized in that the method contains a further step IIb, situated between step IIb and IIIb, comprising contacting the flotation concentrate of step IIb with an aqueous solution of MgCl.sub.2.

3. The method of claim 1, wherein said aqueous potassium- and sulphate-containing composition in step Ia or Ib is a solution mining brine.

4. The method of claim 3, wherein said method comprises contacting one or more potash-containing ores with water so as to obtain said aqueous potassium- and sulphate-containing composition.

5. The method according to claim 1, wherein said aqueous potassium- and sulphate-containing composition comprises about 5 to about 100 g/l of K.sup.+ ion, more in particular about 20 to about 50 g/l of K.sup.+ ion.

6. The method according to claim 1, wherein said aqueous potassium- and sulphate-containing composition comprises about 10 to about 150 g/l of SO.sub.4.sup.2 ion, more in particular about 40 to about 100 g/l of SO.sub.4.sup.2 ion.

7. The method according to claim 1, wherein said aqueous potassium- and sulphate-containing composition comprises about 1 to about 100 g/l of Mg.sup.2+ ion, more in particular about 20 to about 50 g/l of Mg.sup.2+ ion.

8. The method according to claim 1, wherein, in step Ia, contacting said aqueous potassium- and sulphate-containing composition with magnesium chloride is carried out by contacting said aqueous potassium- and sulphate-containing composition with an aqueous composition comprising said magnesium chloride.

9. The method according to claim 1, wherein said method comprises (in step IIa, step IIb, step IIa or step IIb) controlling the concentration of sodium chloride present in said composition comprising kainite so as to maintain said concentration of sodium chloride below about 10% by weight, preferably below about 5% by weight, more preferably below about 2.5% by weight, most preferably below 1% by weight on dry matter basis.

10. The method according to claim 9, wherein controlling the concentration of sodium chloride, present in said composition comprising kainite, is only carried out by means of a flotation technique.

11. The method according to claim 9, wherein controlling the concentration of sodium chloride, present in said composition comprising kainite is, further to a flotation step, carried out by the method step of IIa or IIb.

12. The method according to claim 9, wherein in step IIa or IIb, the aqueous solution of MgCl.sub.2 is at least part of the conversion liquid of step IIIa or IIIb, respectively, or at least part of the MgCl.sub.2 purge of step Ia or Ib, respectively.

13. The method according to claim 9, wherein said controlling of said concentration of sodium chloride, present in said composition comprising kainite, is effective for obtaining a concentration of kainite of above 50% by weight, preferable above 60% by weight, more preferably above 70% by weight, and most preferably above 80% by weight, on dry matter basis.

14. The method according to claim 1, wherein said composition comprising kainite is reacted with water, optionally comprising magnesium sulphate and potassium sulphate, at a temperature of about 45 C. to about 70 C.

15-16. (canceled)

17. The method according to claim 1, wherein said crystallized potassium sulphate obtained contains less than about 10% by weight of impurities, less than about 5% by weight of impurities, preferably less than about 2% by weight of impurities, less than about 1% by weight of impurities, or less than about 0.5% by weight of impurities.

18. The method according to claim 1, wherein contacting said leonite with water so as to leach said MgSO.sub.4 contained in said leonite and to at least substantially selectively precipitate said potassium sulphate (K.sub.2SO.sub.4) is effective for providing potassium sulphate that is crystallized and said method further comprises separating said crystallized potassium sulphate from a brine by means of a solid-liquid separation, wherein the brine may comprise potassium sulphate and magnesium sulphate.

19. The method according to claim 18, wherein said method further comprises recycling said brine and using said brine for reacting kainite with said brine that comprises magnesium sulphate and potassium sulphate to convert said kainite into leonite.

20. The method according to claim 1, wherein the crystallization and/or precipitation of said potassium sulphate is carried out at a temperature of about 45 C. to about 60 C., preferably about 48 C. to about 55 C.

21. The method according to claim 1, wherein the bloedite precipitation in the tailings leach is achieved using seeding, either initially, intermittently and/or continuously.

22. The method according to claim 1, wherein the bloedite precipitation in the tailings leach is used to control the overall sulphate level in the method according to claim 1.

Description

DETAILED DESCRIPTION OF THE DRAWINGS

[0102] In the following drawings, which represent by way of example only, various embodiments of the disclosure:

[0103] FIG. 1 shows a block diagram of an example of a process according to the present invention. The Roman numerals (I, II, etc.) refer to the process steps with the same Roman numerals.

[0104] According to one aspect of the invention, the brines (or salt compositions) that can be used in the methods of the present disclosure can be either naturally occurring, as in lakes, springs, or subsurface brine deposits, or produced by actively solution-mining deeper, more consolidated deposits. The brine can be concentrated in solar evaporation ponds by evaporation and the composition of the brine, as it progresses through a series of ponds, can be controlled by the use of recycled brine from subsequent steps in the process so as to produce salts comprising kainite, halite (NaCl), optionally carnallite (KMgCl.sub.3.6(H.sub.2O)) and hydrated magnesium sulphate salts, other than leonite or schoenite, such as MgSO.sub.4.6H.sub.2O in the solar ponds. For example, by management of the amount of bloedite precipitated in the tailings leach step, the chemistry of the solar ponds can be controlled so that harvested salts will not contain carnallite or magnesium sulphate hydrated salts.

[0105] Solar salts from the harvest ponds comprising kainite and halite can have a kainite concentration above about 50% by weight, for example above about 60% by weight, above about 70% by weight, or above about 80% by weight. For example, the concentration of kainite can be increased by means of flotation and/or leaching with suitable brine, where the species to be rejected are halite and hydrated magnesium sulphate salts, such that concentrated salts are obtained. The rejected species are further led to a tailings leach stage, where they can be removed from the process, or recycled to the ponds, either as a liquid or as a solid.

[0106] The concentrated salts can have a kainite concentration of above 65% or 70% by weight, in particular 80% by weight, or more, and they can then be reacted (conversion) at a temperature above about 35 C., or of about 35 C. to about 65 C., with recycled brine from subsequent steps in the process (also called SOP mother liquor) to convert the kainite into leonite. The use of this recycled brine (SOP mother liquor), which can contain a significant concentration of potassium sulphate, results in more leonite being produced than the potassium ion in the kainite feed alone would permit. For example, depending on the temperature of the conversion, other MgSO.sub.4 contaminants may be precipitated, as well as leonite, and the leonite resulting from this reaction, if necessary to achieve a purity which is suitable for a feed to a potassium sulphate crystallization circuit, may be leached with suitable brine (leonite leach) and subjected to known solid-liquid separation techniques. At temperatures above about 35 C. or above about 45 C., the formation of schoenite was not observed. The brine, resulting from the conversion (conversion brine) can be returned to the tailings leach.

[0107] The magnesium sulphate, contained in the leonite, can then be subjected to selective leaching with water (for example water added or added to water) and crystallization, for example, in a vessel or vessels designed to promote crystal growth, whereby substantially all of the magnesium sulphate and a portion of the potassium sulphate contained in the leonite are taken into solution (or leached), with the remaining portion of the potassium sulphate produced as crystalline material. This crystallization can be conducted at a temperature of about 45 C. to about 60 C. For example, and without wishing to be bound by such a theory, leonite can be dissolved substantially at the same time the K.sub.2SO.sub.4 crystallization occurs.

[0108] For example, clear brine from this step can be used in earlier steps of the process where additional leonite may be precipitated. For example, it can be used for reacting magnesium sulphate in the kainite conversion reaction step into leonite. The clear brine can have a magnesium to potassium weight ratio of about 0.4 to about 0.7 or of about 0.5 to about 0.6. Potassium sulphate, remaining in brine streams, eventually recycled to the solar evaporation ponds, can again be captured as solid kainite and recovered. The potassium sulphate solids can be withdrawn from the crystallization equipment and may or may not be leached with additional water before being subjected to known solid-liquid separation techniques, where they may or may not be washed with water.

[0109] The high purity potassium sulphate solids can then be dried, sized and either granulated to meet market specifications or sold as produced.

[0110] Brines containing ions of K, Mg, Na, Cl and SO.sub.4 can be concentrated by solar evaporation and by the use of recycle brines caused to precipitate salts comprising kainite, halite, carnallite and one or more hydrated magnesium sulphate salt.

[0111] The methods of the present disclosure can be directed to the production of high purity potassium sulphate, encompassing a maximized recovery of potassium sulphate in the crystallization step, by a process including conversion of kainite to high purity leonite in a system operating at high ambient temperature (for example temperatures above about 35 C.; temperatures of about 35 C. to about 65 C.; or about 35 C. to about 55 C.). At temperatures of about 45 C., formation of schoenite was not observed.

[0112] When tests were conducted to confirm conversion of kainite, containing appreciable amounts of halite and magnesium sulphate hydrates, to leonite in reaction with brine from the potassium sulphate crystallization step at a temperature at or above about 45 C., the resulting leonite was contaminated with bloedite (Na.sub.2Mg(SO.sub.4).4H.sub.2O) not removable by washing. It was subsequently discovered that this is related to a high concentration of sodium ions in solution which results in bloedite forming, not as a separate discrete species, but apparently as crystal lattice replacement within the leonite crystals (a solid solution of the two species). Without wishing to be bound by such a theory, this is likely the result of the similarity between leonite and bloedite crystal structure; they are analogs in that both are four water hydrates of a magnesium sulphate double salt, with very little difference in size between the potassium and sodium ions (1.33 and 0.96 Angstrom respectively). The inventors found that contamination of leonite with bloedite by this mechanism may be controlled by maintaining the concentration of sodium ion in the conversion reaction brine low, say, for example, below about 10% by weight, below about 4% by weight, below about 2% by weight, or below about 1.4% by weight, and controlling the degree of super saturation created in the reaction vessels.

[0113] Without wishing to be bound by such a theory, it is believed that this crystal lattice replacement phenomenon is analogous to the contamination of sodium carbonate decahydrate crystal by crystal lattice inclusions of sodium sulphate decahydrate, experienced by the inventors in previous work. For the sodium carbonatesodium sulphatewater system, the degree of contamination is directly proportional to the concentration of sulphate ion in the SOP mother liquor. There was also an apparent correlation observed with the degree of super saturation created in the crystallizerhigher super saturation level and more rapid crystal formation accompanied by more sulphate in the crystal latticealthough this was difficult to prove beyond question, as was an apparent correlation with temperature.

[0114] The presence of magnesium sulphate, not associated with the potassium sulphate ion, requires higher water to potassium sulphate ratio to dissolve all the magnesium sulphate contained in the leonite feed to the potassium sulphate crystallizer; this results in a higher percentage of the potassium sulphate contained in the leonite being taken into solution. Put in another way, the result is lower recovery of potassium as solid potassium sulphate and higher recycle brine flow because more water is used per unit of potassium sulphate produced, and larger evaporation ponds and plant are required for any given production capacity.

[0115] According to another aspect of the invention, the tailings leach can advantageously be used to control the sulphate level in the entire process as described above, through bloedite precipitation. This is due to the fact that the brine from the tailings leach tank is recycled to the ponds for potassium recovery. The net effect is that the brine that is returned to the pond system has a much lower (but controllable) sulphate concentration than without the innovative tailing leach step. In this new process, the tailings salts contain bloedite which provides a solid purge point in the process for excess sulphate. According to one embodiment, this could replace a liquid MgSO.sub.4 purge (FIG. 1: Purge brine) that is situated in the leonite leach step (step IV), increasing overall potassium recovery in the process.

The bloedite precipitation (Step VI) has an impact on several steps of the process.

[0116] a) Step I

[0117] The largest impact from the tailings leach step according to the invention, is on the ponds area. The tailings leach brine returns to the pond system for further evaporation and K-recovery. The composition of said recycle stream is directly affected by the Tailings Leach reaction, and thus, brine compositions in the pond system are also affected. Without the bloedite precipitation step, this stream contained a high concentration of sulphate, which led to the precipitation of undesirable hydrated magnesium sulphate salts (such as hexahydrite) in the pond system. According to the invention, within a reasonable reaction time in the tailings leach step, the brine returned to the pond system will contain much less magnesium sulphate, while containing more MgCl.sub.2. This will change the pond system chemistry to the point where no magnesium sulphate salts are expected to precipitate in the pond system. Incorporating the Tailings Leach step into the process according to the invention would thus reduce the total tons of material being harvested and transported to the plant. Furthermore, the fact that no magnesium sulphate salts are carried to the plant would completely eliminate the need for the leonite leach step (step IV).

[0118] Step II (Flotation)

[0119] The flotation operation of the wet process is also impacted by the tailings leach step according to the invention. Because the hydrated magnesium sulphate salts are no longer in the feed to the process step II, concerns about trying to keep them from floating with the kainite disappear. Optimization of the flotation circuit can focus entirely in getting rid of NaCl carried with the kainite. This should also improve the overall grade and recovery of the flotation concentrate produced in the process, which then is transferred to step III). Furthermore, the absence of MgSO.sub.4 salts in the ponds salts removes one possible variation in the composition. Lower variability in the solids feed will simplify control of the flotation equipment. The main impact on the flotation cells is the lower tonnage of salts processed, as no magnesium sulphate hexahydrate is harvested. This is due to the fact that less salt is harvested.

[0120] Step III (Conversion)

[0121] The conversion circuit is not impacted by the Tailings Leach step according to the invention. The operation remains the same, and the equipment required should not change. However, since there is absolutely no MgSO.sub.4 solids floating with kainite and being fed to the conversion reactors, the solids tonnage processed is lower.

[0122] Step IV (Leonite Leach)

[0123] As already indicated, the absence of MgSO.sub.4 solids in the process feed makes this step obsolete. This represents a direct elimination of mechanical equipment. According to one embodiment, common equipment linked to the leonite leach which can be eliminated are: an agitated tank, a distributor, pan filters, brine tank, pumps and conveyor. This is the single largest impact of the tailings leach step according to the invention.

[0124] Step V (Crystallization)

[0125] The crystallization section of the process is not affected directly by the tailings leach step according to the invention. The same amount of leonite has to be processed in order to reach the target SOP production, and the same product purity will be reached. However, the lower MgSO.sub.4 concentration in the brine, carried with the solid leonite to the crystallizer, is a benefit to the overall process. With the absence of MgSO.sub.4 solids in the harvested salts is an associated lower MgSO.sub.4 concentration in the process brines recycled through the process. This is particularly advantageous on the leonite being fed to the crystallizer, because all MgSO.sub.4 coming into the crystallizer (whether from the solids or the brine) will lower the recovery of the crystallization circuit. The brine carried with the leonite out of the conversion reactors contains less MgSO.sub.4 than without the claimed tailings leach step. This solid will be washed on the leonite pan filter, but a lower MgSO.sub.4 content in the brine will still reduce the MgSO.sub.4 fed to the crystallizer. Additionally, no pipeline to carry the MgSO.sub.4 purge from the plant to the pond area is required, resulting in more capital cost savings.

EXAMPLES

Example 1

[0126] The following example illustrates the method according to the invention. Optimization was not performed but the gist of the invention is shown hereunder. All process steps are performed in the laboratory on a laboratory scale.

[0127] Step I was not performed. The salt mixture used in the laboratory testing was made in the laboratory. The kainite salt was produced from a laboratory brine, made from commercially available halite and magnesium sulphates.

[0128] All testing was done in a bench scale range of 1-8 kg. However, the numbers in the tables below are adjusted to reflect a starting solid of 100 kg to Step II (kainite concentration).

[0129] Step II: Concentrating Kainite and Removal of Halite Through Flotation

[0130] A salt mixture of 57 weight % kainite, 18 weight % halite, 22 weight % magnesium sulphate and 6 weight % bishofite (MgCl.sub.2.6H.sub.2O) was slurried in a flotation brine (composition: NaCl, KCl, MgCl.sub.2, MgSO.sub.4.7H.sub.2O and water). A frother aid and a flotation aid was added and the frothy supernatant was collected, filtered to remove remaining brine and kept for further processing in Step III. The salt mixture was ground to a P.sub.80 of about 350 microns). Flotation was carried out at 45 C. Recovery of K was 90% (see Table 1).

TABLE-US-00001 TABLE 1 Concentrating kainite and removal of halite through flotation Slurry fed to flotation Flotation concentrate Salt mixture (solids in top fraction after (from Step I) Flotation Brine flotation) 100 kg 370 kg 64 kg K 9.0% 1.0% 13% Mg 8.6% 6.6% 10% S 10.5% 2.5% 12% Cl 21.3% 17.6% 18% Na 6.8% 1.8% 2% All % based on weight.

[0131] Step III: Conversion of Kainite into Leonite

[0132] The process was performed in semi-continuous mode to prevent problems with super-saturation and sudden precipitation. The solids top fraction from step II (Flotation concentrate) and SOP mother liquor brine from step V (synthetically made) was added in increments to a starting brine having the composition for an continuous process. The process was maintained at 45 C. and the retention time was 1 hour. The slurry was filtered and the solids (Leonite solids) were kept for further processing in Step IV. Leonite was added to seed the precipitation (Table 2).

TABLE-US-00002 TABLE 2 Conversion of kainite into leonite Flotation SOP Mother Starting Starting Leonite concentrate liquor Brine Leonite solids 64 kg 170 kg 177 kg 62 kg 180 kg K 13% 5.7% 1.9% 19.7% 18.7% Mg 10% 3.5% 5.5% 6.9% 7.5% S 12% 6.9% 4.5% 17.2% 16.7% Cl 18% 0.1% 10.4% 0% 2.9% Na 2% 0.05% 1.8% 0.03% 0.8% All % based on weight.

[0133] Step IV: Washing of Leonite

The solids from step III (Leonite solids) were re-slurried in leach brine to dilute entrained brine from the conversion reactor for 60 min (leach brine=SOP mother liquor almost saturated with MgSO.sub.4, similar to purge brine). It was then filtered and washed with brine from SOP crystallizer (SOP mother liquor). The filtered solids (Washed Leonite solids) were kept for further processing in Step V (Table 3).

TABLE-US-00003 TABLE 3 Washing of leonite SOP mother Washed Leonite solids Leach brine liquor Leonite solids Kilogram 180 kg 440 kg 127 kg 163 kg K 18.7% 2.8% 5.7% 20.2% Mg 7.5% 5.3% 3.5% 6.9% S 16.7% 8.3% 6.9% 18.4% Cl 2.9% 0.1% 0.1% 0.1% Na 0.8% 0.05% 0.05% 0.3% All % based on weight.

[0134] Step V: SOP Crystallization

This process was performed in a semi-continuous mode. The crystallizer was loaded with a starting brine made from 0.49 weight % of the water and 59 weight % of the solid (leonite). The remaining salts and water were added in increments, while clear liquid was removed to keep the amount constant. The procedure lasted approximately 6 hours. The slurry was then centrifuged and dried. The potassium sulphate produced had a K.sub.2O content over 50%, and a Cl content below 1%, which reflects the standard grade of chlorine free potassium sulphate (Table 4).

TABLE-US-00004 TABLE 4 SOP crystallization Filtrate Leonite solids Water (SOP Mother (total) (total) SOP solids Liquor) 147 kg 181 kg 24 kg 304 kg K 20.8% 41.9% 6.0% Mg 7.0% 0.4% 3.3% S 18.7% 17.5% 7.0% Cl 0% Na 0.2% 0.1% 0% All % based on weight.

[0135] Overall recovery is about 48% for this laboratory scale experiment. Although the recovery is somewhat low, the method can be optimized to achieve recoveries of 60% and more.

Experiment 2: (Step IIHalite Leach Step)

[0136] In the laboratory, three brine compositions were made as indicated in the Table 6 and agitated at 54 C. After two hours of mixing at a temperature of 54 C., some suspended solid remained and were filtered off on a hot Buchner funnel and vacuum filter system with 1.6 m filter paper. In a 3-liter agitated open top kettle heated with a heating mantle and controlled with a therm-o-watch to 54 C., 400 grams of flotation concentrate solids, 171 gram of a first brine (brine #1, flotation concentrate) and 300 gram of a second brine (brine #2, MgCl.sub.2) were mixed until the solution was a homogeneous slurry. Next, 753 gram of a third brine (brine #3, conversion liquid) was added to the slurry, comprising about 60 weight % of solids. The slurry was held at 54 C. and filtered brine samples (brine) were collected every 15 minutes for 90 minutes. After 90 minutes, the entire kettle content was dewatered by centrifugation and samples of concentrate and the final solids were collected and analysed.

[0137] The test was repeated in the same manner as above, except that a single representative brine was used instead of three separate ones. Comparable results were obtained. The results are shown in FIG. 2 and the Table 5 below.

TABLE-US-00005 TABLE 5 Halite Leach Time Sample No. Sample Type (min) % Na % K % Mg % Cl 261-69-01 Initial Flotation 3.60 12.1 8.58 18.8 concentrate solids 261-71-11 Brine #1 1.10 1.55 6.91 19.2 261-71-12 Brine #2 0.33 0.51 8.27 24.5 261-71-14 Brine #3 0.09 3.91 5.14 7.62 261-71-01 Brine 15 1.47 2.23 6.16 15.2 261-71-03 Brine 45 1.55 2.06 6.18 15.4 261-71-04 Brine 60 1.59 1.97 6.17 15.6 261-71-05 Brine 75 1.66 1.90 6.21 15.8 261-71-06 Brine 90 1.65 1.80 6.32 16.0 261-71-09 Concentrate 1.65 1.77 6.34 15.8 261-71-10 Final solids 0.21 15.0 9.77 13.5

[0138] From the Table 5 above, it can be seen that the halite (sodium in particular) is removed from the solids down to 0.21 weight %.

Experiment 3 (Step IIHalite Leach Step)

[0139] The experiment 2 above was repeated to study the amount of kainite precipitation. This was done by preparing a feed brine having the same compositions as expected in the process. Salts with a much higher halite content than expected in the process were then added to the system. The brine composition of the slurry was be studied throughout the test. This provided kinetics information as well as the maximum sodium concentration of the brine to be expected in the process.

[0140] The temperature of operations selected was 52 C. All brines were heated to this temperature and the vessel was maintained at this temperature throughout the test.

Experimental Details

1) Materials

[0141] a. 3000 g Flotation Concentrate with 30% solids, composed of: [0142] Solids Composition: 900 g total, 17% NaCl, 83% Kainite [0143] Brine Composition: 2,100 g total, 0.36% KCl, 30.33% MgCl.sub.2, 0.79% NaCl, 4.29% MgSO.sub.4, 64.23% H.sub.2O.

[0144] b. 480 g Wash Brine, composed of: [0145] 5.66% KCl, 1.4% NaCl, 9.71% MgCl.sub.2, 14.76% MgSO.sub.4, 68.47% H.sub.2O

[0146] c. 1,725 g Process Brine, composed of: [0147] 7.28% KCl, 5.56% MgCl.sub.2, 0.1% NaCl, 17.83% MgSO.sub.4, 69.23% H.sub.2O

2) Method

[0148] 1) Brines b and c were prepared in dedicated kettles, heated and agitated to 52 C. [0149] 2) Flotation concentrate a was prepared in a 4-liter agitated kettle, which as heated to 52 C. and maintained at that temperature throughout the test. Subsequently, brines b and c were added (some precipitation may occur). [0150] 3) The temperature in the kettle was stabilize at 52 C. for 30 minutes. A sample of about 100 ml of starting slurry (or brine) was taken with a syringe. [0151] 4) After sampling, specified quantities of solids were added to the agitated 4-liter kettle. The entire mass of solids was added in one step (time t=0). [0152] 5) At times t=5, 10, 15, 20, 30, 45, 60 minutes, a slurry sample was taken and analysed for solids and liquid composition. A sample of about 100 ml of slurry was taken with a syringe. [0153] 6) The remaining slurry was filtered and solids and liquid produced during test were analysed.

[0154] The results are shown below in Table 6.

TABLE-US-00006 TABLE 6 Conversion liquid Halite Halite Leach Flotation Wash Process Flotation concentrate solids Leach Solids to Brine + MgCl.sub.2 Brine Brine Halite Kainite Filtrate Conversion Amounts (g) 1778 398 1448 141 689 3642 1061 K (Wt %) 0.2 3.3 3.8 0 15.5 1.4 12.3 Mg (Wt %) 8.4 6.2 5.2 0 9.7 6.3 8.3 S (Wt %) 1.3 4.6 5.0 0 11.7 2.3 11.1 Na (Wt %) 0.5 0.7 0.1 37.1 0 1.3 1.8 Cl (Wt %) 23.2 12.2 7.7 60.5 16.4 17.1 15.0 SO.sub.4 (Wt %) 3.8 13.9 14.9 0 35.1 6.9 33.4

CONCLUSION

[0155] One may calculate that the solid input (Flotation concentrate solids) has an amount of 106.9 g of Potassium (689 g15.5%). After 60 minutes mixing with the Conversion Liquid, there is 133 g Potassium (1061 g12.3%) in the Solids to Conversion. Hence, the amount of Potassium in the Solid to Conversion has increased by about 20%.
The input had an amount of 17% Halite, or 6% of Na (141 g37.1%). After 60 minutes of mixing with the brines, the Na concentration decreased to 1.8% (or about 5% NaCl). In order to have a nearly complete removal of Na, preferably the amount of Na should be less than 6% in the Flotation Concentrate solids, as was also shown in Table 1.

[0156] While a description was made with particular reference to the specific embodiments, it will be understood that numerous modifications thereto will appear to those skilled in the art. The scope of the claims should not be limited by specific embodiments and examples provided in the present disclosure and accompanying drawings, but should be given the broadest interpretation consistent with the disclosure as a whole.