CA2552104C - Process for recovery of sulphate of potash - Google Patents

Abstract

A novel integrated process for the recovery of sulphate of potash (SOP) from sulphate rich bittern is disclosed. The process requires only bittern and lime as raw materials. Kainite type mixed salt is obtained by fractional crystallization of the bittern. Kainite is converted to schoenite with simultaneous removal of NaCl by processing it with water and end liquor obtained from reaction of schoenite with MOP for its conversion to SOP. The end liquor from kainite to schoenite conversion (SEL) is used for the recovery of MOP. SEL is desulphated and supplemented with MgC12using end bittern generated in the process of making carnallite. The carnallite is decomposed to get crude potash which in turn processed to get MOP. The carnallite decomposed liquor produced in the decomposition of carnallite is reacted with hydrated lime for preparing CaC12 solution and high purity Mg(OH)2 having low boron content. The CaC12 solution is used for desulphatation of SEL producing high purity gypsum as a byproduct. It is shown that the liquid streams containing potash are recycled in the process, the recovery of potash in the form of SOP is quantitative.

Description

PROCESS FOR RECOVERY OF SULPHATE OF POTASH

Field of the invention The present invention provides an integrated process for the recovery of sulphate of potash (SOP) from sulphate-rich bittern. The process requires only bittern and lime as raw materials and affords, besides SOP, low boron containing Mg(OH)2 , gypsum and salt, as co-products, all of which are obtained in pure form.

Background of the invention SOP is a dual fertilizer containing 50% K20 and 18% S. It has the lowest salt index and is virtually free of chloride, which makes it a superior fertilizer to muriate of potash (MOP). On the other hand, MOP is easy to produce, especially, when brine/bittern is low in sulphate content such as in the Dead Sea and this accounts for its lower price compared to SOP. Countries such as India, which do not have low sulphate bittern, but which have adequate bittern of sea and sub-soil origin, would be greatly benefited if SOP can be produced economically from such bittern sources.
Besides its application as a fertilizer, potassium sulphate has numerous industrial applications as well.
Mg(OH)2 is commercially used in pulp and paper industries and also as antacid and fire retardant. Waste water and acidic effluent treatment represent additional high growtli areas for its application. Mg(OH)2 is also used for production of magnesia (MgO), magnesium carbonate and other magnesium chemicals. Mg(OH)2 that is low in B2O3 impurity is especially suitable for production of refractory grade MgO.
High quality gypsum (CaSO4.2H2O) finds applications in the white cement industry and for manufacture of high strength a and 0 Plaster of Paris. Sodium chloride that contains small quantities of potassium chloride finds application in the edible salt industry.
Reference is made to the well-lcnown Mannheim process involving reaction of MOP with sulphuric acid. The major problem with the process is that it is energy intensive and poses a problem of HCl management when no application of commensurate volume for HCl is available in the vicinity. J. A. Fernandez Lozano and A. Wint, ("Production of potassium sulphate by an ammoniation process", Chemical Engineer, 349, pp 688-690, October 1979) disclose a process of SOP manufacture from MOP through reaction with gypsum in presence of ammonia. The principle of the process is double decomposition reaction between gypsum and potassium chloride in presence of ammonia at 0 C. The main disadvantage of the process is that it is energy intensive and necessitates careful.design of the reactor for safe operation.

H. Scherzberg et al. ('Messo pilots new potassium sulphate process', Phosphorous & Potassium, 178, March-April 1992, p-20) describe the successful trials on a process involving reaction of MOP with sodium sulphate to produce the double salt glaserite (3KaSO4.NaZSO4). The glaserite is in turn reacted with MOP to produce SOP. The main disadvantage of the process is that it would be unsuitable for those who do not have access to such raw materials. Moreover, the process involves several complex unit operations including the need for chilling. Such processes have their limitation on large scale.
H. Scherzberg and R. Schmitz ('Duisberg's alternative to Mannheim', Phosphorous & Potassium, 178, March-April 1992, p-20), describe an integrated process for production of SOP from KCl and MgSO4 or Na2SO4. The main drawback of the process is that the amount of NaCl in raw materials has a critical effect on the process and, as such, is less applicable to crude mixed salt as obtained from sea bittern.
Another disadvantage is that the process involves heating and cooling which makes it energy intensive. Yet another disadvantage is that the by-product obtained is MgC12 in concentrated solution form which has a limited market and lower appeal compared to low B2O3 containing Mg(OH)2 solid produced as part of the integrated process of the present invention.

G. D. Bhatt et al. ('Mixed Salt from Sea Bittern', Salt Research & Industry, 2, .20 126-128, 1969) describe a process for the manufacture of mixed salt, i.e., comprising of a mixture of NaCI and kainite (KC1.MgSO4,3H2O), from sea bittern through solar evaporation and fractional crystallisation.
Patel et al. (Salt Research & Industry, Vol.6, No. 14, 1969) disclose a process for the preparation of syngenite from mixed salt in pure form. K. P. Patel, R.
P. Vyas and K. Seshadri (`Potassium Sulphate from Syngenite', Salt Research &
Industry, Vol.6, No. 2, April 1969) disclose a process for preparation of SOP by leaching syngenite (K2S04.CaSO4.H2O) with hot water and then recovering it by solar evaporation. The main drawback of the process is that it is energy intensive.
Moreover, production of syngenite from mixed salt is itself an involved affair.
K. Sehsadri et al ("Manufacture of Potassium chloride and byproducts ftom Sea Bittern" Salt Research and Industry, Apr=il-July 1970, Vol. 7, page 39-44) disclose a process wherein mixed salt (NaCI and kainite) obtained from bittern is dispersed with high density bittern in proper proportion and heated to a temperature of 110 C
when kieserite (MgSO4.H2O) is formed which is separated by filtering the slurry under hot conditions. The filtrate is cooled to ambient temperature, when carnallite crystallizes out. Camallite is decomposed with water to get a solid mixture of sodium chloride and potassium chloride while magnesium chloride goes into solution. Solid mixture of potassium chloride and sodium chloride is purified using known techniques to produce pure potassium chloride. The drawbacks of this process are that it fails to make use of the sulphate content in bittern and, instead, offers an elaborate process for manufacture of MOP, which, in any case, is inferior to SOP as fertilizer.
US Patent Application Number 20030080066 dated October 29, 2001 by Vohra, Rajinder N. et. al. discloses an integrated process for recovery of high purity salt, potassium chloride, and end bittern containing 7.5 gpl Br. The process is based on desulphatation of brine with distiller waste of soda ash industry or calcium chloride generated from limestone and acid. The main drawback of the patent application is that the process is less attractive when distiller waste is not available in the vicinity and the process becomes less economical when carnallite has to be obtained from bittern without production of industrial grade salt. Moreover, as in the case referred to above, it is desirable to utilize the sulphate content in bittern and produce SOP in preference to MOP.
Michael Freeman ('Great Salt Lake-A fertile harvest for IMC'. in Phosphorus &
Potassium, 225, Jan-Feb, 2000) describe a process comprising concentrating the brine containing 0.2-0.4% KC1, harvesting mixed salt, separation of high sodium chloride fraction through floatation, leaching with sulphate rich brine to produce schoenite, hot water dissolution of schoenite, fractional crystallization of SOP and recycling of mother liquor containing up to 30% of original K to evaporation pond. The main drawbacks of the process are: (i) need for floatation which involves use of organic chemicals whose disposal is problematic, (ii) need for external heat for recovery of SOP from schoenite through fractional crystallization at elevated temperature, (iii) need for recycling of as much as 30% of K to evaporation ponds where it again gets contaminated with 'other components of the brine.
In Ullmann's Encyclopedia of Industrial Chemistry, Sixth Edition, 1999, under the Chapter, Potassium compounds, a description of a process for production of SOP in Sicily is given. Kainite (KCl - MgSO4 = 2.75 H20), is obtained from a potash ore by flotation. It is then converted into schoenite at ca. 25 C by stirring with mother liquor containing the sulfates of potassium and magnesium from the later stages of the process. Schoenite is filtered off and decomposed with water at ca. 48 C. This causes magnesium sulfate and part of the potassium sulfate to dissolve and most of the potassium sulfate to crystallize. The crystals are filtered and dried. The sulfate mother liquor is recycled to the kainite - schoenite conversion stage. The main drawbacks of the process are that there is no mention of the fate of the mother liquor obtained upon conversion of kainite into schoenite which woul& inevitably entail considerable loss of K, and the need for external source of heat to effect the fractional crystallization of SOP.
Chinese Patent CN 2000-112497, 29 Aug. 2000, by Song, Wenyi; Liu, Yu;
Zhao, Shixiang; Dai, Fangfa, titled method for preparing K2S04 from sulphate type K-containing bittern. The method comprises concentrating the bittern, separating NaC1, concentrating to obtain crude K-Mg salt containing 10-45% NaCI, crushing, mixing with saturated bittern to obtain a solution witli concentration of 20-40%, removing NaCl by back-floatation, concentrating, dewatering to obtain refined K-Mg salt containing less than 5% NaCI, mixing the K-Mg salt and water at specified ratio, allowing the mixture to react at 10-60 F for 0.5-3hr, separating to obtain schoenite, mixing with KC1 and water at specified ratio, allowing the mixture to react at for 0.25-3hr and separating to obtain K2SO4. The drawbacks of the process are (i) need for elaborate method of purification of mixed salt that includes removing NaC1 by the less desirable method of back floatation that involves use of organic chemicals, (ii) lack of any mention of the manner in which the various effluent streams are dealt with, and (iii) dependence on outsourced KC1 since no mention is made of any process for production as part of the process.
J. H. Hildebrand ('Extraction of Potash and other Constituents from sea water Bittern' in Journal of Industrial and Engineering Chemistry, Vol. 10, No. 2, 1918, pp 96-106) describe theoretical aspects of the recovery of potash from sea bittern and propose a process for extraction. According to this process, bittern is evaporated at a temperature between 100-120 C, thereby forming a solid mixture of sodium chloride and kieserite (MgSO4.H20), separating this mixture under hot conditions in a heated centrifuge, and cooling the mother liquor in a cooler for separation of carnallite.
Carnallite is decomposed and washed with water to produce potassium chloride.
The drawbaclc of this process is that it is demanding in terms of energy requirement and sufficiently pure. carnallite cannot be obtained. The main drawback of the process is the contamination of kieserite with NaCl, which would necessitate further purification to obtain products in saleable form. Another drawback of the process is that it requires energy to remove sulphate from bittern in the form of kieserite whereas it would be preferable to utilize the sulphate for the production of SOP.
D. J. Mehta et al ('Production of Potassium Sulphate from Mixed Salt obtained from Salt Works of Little Rann Of Kutch' Salt Research & Industry, Vol. 2, No.
4, October 1965) describe a process using floatation technique for the production of potassium sulphate from two types of mixed salt available from the salt works of the Little Rann of Kutch. The process suffers from the drawback of lack of suitability when high sulphate containing sea bittern is used and the need for froth floatation, which is costly, cumbersome and polluting.
Reference is made to the Chapter in Ullmann's Encyclopedia of Industrial Chemistry, Sixth Edition, 2002, (Electronic Version) dealing with Magnesium Compounds written by Margarete Seeger, Walter Otto, Wilhelm Flich, Friedrich Bickelhaupt and Otto. S. Akkerman, wherein the process of preparation of magnesium llydroxide from seawater is described. It is mentioned therein that preparation of low boron containing magnesia requires over liming of the seawater up to pH 12 to maintain B2O3 content less than 0.05% in magnesia. Over liming involves higher lime cost, need for neutralization of supernatant and results in a colloidal suspension which is not easy to filter. Another drawback is a lack of application of calciuin chloride-containing effluent which is discharged back into the sea.
Patent Application No. 423211, CA 1203666, by Wendling et al titled, "Process for the manufacture of potassium sulphate by treatment of solution containing magnesium chloride and potassium chloride" describes a process for the production of potassium sulphate from solutions containing magnesium chloride, such as solutions of carnallite ore and, in particular, the equilibrium mother liquors of a unit for the treatment of carnallite. According to this process, sodium sulphate and potassium chloride are added to the solutions containing magnesium chloride, so as to precipitate sodium chloride and schoenite, K2SO4MgSO46HaO, and the schoenite obtained is treated in a lcnown manner to produce potassium sulphate. The main drawback of the process is the need to outsource sodium sulphate and the lack of any mention of a solution to the problem of KCl loss in effluent streams.
H. Gurbuz et al. (`Recovery of Potassium Salts from Bittern by Potassium Pentaborate Crystallisation' in Separation Science & Technology, 31(6), 1996, pp. 857-870) disclose the preparation of sodium pentaborate from the reaction of Tincal and recycled H3BO3 in presence of water and thereafter treated with bittern to selectively precipitate out potassium pentaborate, which in turn is acidulated with sulphuric acid and fractionally crystallized to remove K2S04 and recycle the H3BO3 in the process.
The main drawbacks of the process are that the mother liquor contains significant quantities of boron, which entails elaborate procedure to recover boron and, moreover, the MgO obtained from such mother liquor would be unfit for industrial use.
Moreover, although such a process can still be thought of for sulphate poor bittern, it would not be a preferred route when the bittern is rich in sulphate content. Yet another drawback is the need to chill the acidulated product for high yield.
A. S. Mehta (Indian Chemical Engineer, 45(2), 2003, p. 73) describes a process of bromine manufacture from bittern. Bittern is acidified with sulphuric acid to a pH of 3.0-3.5 and the bromide ion is then oxidized with chlorine and stripped off with the help of steam. The acidic de-brominated bittern is neutralized with lime, the sludge thus formed removed, and the effluent discharged. Bromine plants located in the vicinity of natural salt beds in the Greater Rann of Kutch in Gujarat, India utilize natural bittern for bromine production by the above method and discharge their effluent back into the Rann. Disposal of sludge poses a formidable challenge in these plants.
Chr. Balarew, D.Rabadjieva and S. Tepavitcharova ("Improved Treatment of Waste Brines " International Symposium on Salt 2000, page 551-554) describe recovery of marine chemicals. The authors describe the use of lime for precipitation of Mg(OH)2 from a part of available bittern, and desulphatation of balance bittern with the resultant CaC12 solution for recovery of KCl via carnallite. The authors have not discussed any scheme of utilizing such methodology for production of SOP from sulphate-rich bittern.
Moreover, as will be evident later, Mg(OH)2 produced directly from raw bittern has much higher B203 content compared to Mg(OH)2 prepared from the Mg2+ source of the present invention, which is linlced to production of SOP.
Chinese Patent No. 1084492, Lu Zheng, describes a process of manufacture of SOP from bittern and potassium chloride. In this process, bittern is processed by evaporation, cooling, floatation, and is then reacted with potassium chloride to make potassium sulfate and by-products of industrial salt and residual brine. The main drawbacks of this process are that it requires involved separation techniques like floatation to remove NaCl from mixed salt and KC1 required for production of SOP
from schoenite has to be procured separately. Moreover, although overall yield in terms of potash recovery is 95%, yield with respect to such procured KC1 is not mentioned.
Objects of the invention It is an important object of the present invention to produce superior fertilizer, S'OP, from sulphate-rich bittern sources such as sea bittern and natural bittern in a cost effective manner through integration with production of valuable co-products.
Another object is to dispense with the need for floatation to remove NaC1 from mixed salt and instead to leach away NaCI in the mother liquor (SEL) and simultaneously convert kainite into schoenite.
Another object is to produce SOP from schoenite under ambient conditions through the known method of reaction witli KCl in presence of water and wherein the MOP is generated from SEL eliminating the need to source it externally.
Anotlier object is to maximize recovery of potash in the form of SOP from mixed salt.
Another object is to desulphate the SEL cost-effectively to promote carnallite formation.
Another object is to evaporate desulphated SEL in a multiple effect evaporator to recover water for reuse.
Another object is to utilize the NaCI separated as edible salt.
Another object is to utilize the MgC12-rich carnallite decomposed liquor (CDL) for cost- effective production of CaCl2 and Mg(OH)2 tlirough treatment with lime.
Another object is to utilize the washings from Mg(OH)2 filtration for preparation of slaked lime from quick lime which conserves water and recycles residual CaC12 in the washings.
Another object is to utilize the CaC12 solution above for desulphatation of SEL.
Another object is to recover KCl that is lost in CDL by recycling the latter in the manner described above.
Another object is to show that MgO produced from the above Mg(OH)2 contains very low (< 0.03 %) levels of B203 impurity.
Another object is to minimize effluent generation in the process and instead to utilize effluent to enhance potash recovery or to convert into value-added products.
Another object is to replace conventionally employed slalced lime with Mg(OH)2 generated in the process of the invention for neutralization of acidic debrominated bittern to eliminate sludge formation when acids such as sulphuric acid are employed for acidulation of bittern and instead make such bittern immediately useful for mixed salt production.
Summary of the invention An aspect of the present invention provides an integrated process for the preparation of sulphate of potash from bitterns, comprising:
(i) subjecting bittern to fractional crystallization to obtain kainite type mixed salt with high kainite content and MgC12-rich end bittern, and subjecting the MgCIZ rich end bittern to desulphation;
(ii) treating the kainite type mixed salt with water and mother liquor obtained in step (xiii) below to leach out substantially all NaCI from the mixed salt and simultaneously convert kainite into schoenite;
(iii) filtering the schoenite and separating the filtrate;
(iv) desulphating the filtrate with aqueous CaC12;
(v) filtering the gypsum produced in step (iv) and mixing the filtrate with the MgC12-rich filtrate obtained in step (vii) below, (vi) evaporating the resultant solution of step (v) and cooling to ambient temperature to crystallize crude carnallite, (vii) centrifuging the crude camallite and recycling the required quantity of filtrate to step (v), (viii) decomposing the crude carnallite with appropriate quantity of water from step (vi) to yield crude KCI and carnallite decomposed liquor;
(ix) filtering the crude KCI, and washing with water to remove adhering MgC12 and subjecting to hot leaching for production of MOP and NaC1, (x) mixing the carnallite decomposed liquor from step (viii) and washing from step (ix) and treating with hydrated lime, (xi) filtering the slurry and washing the cake to produce Mg(OH)2 and CaC12-containing filtrate for the desulphatation process of step (iv).
(xii) treating by known method the schoenite produced in step (iii) with MOP
produced in step (ix) to produce SOP under ambient condition, (xiii) filtering the SOP and collecting separately the mother liquor hereinafter referred to as KEL, (xiv) recycling the KEL of step (xiii) in the process of step (ii).

In accordance with another aspect of the present invention, there is provided an integrated process for preparation of sulphate of potash (SOP) from bittern comprising:
(i) subjecting bittern to fractional crystallization to obtain kainite mixed salt with high kainite content and MgC12-rich end bittern, (ii) treating the kainite mixed salt with water and mother liquor obtained as a filtrate during sulphate of potash preparation in (xiv) below, referred to as KEL, to leach out substantially all NaCI from the mixed salt and simultaneously convert kainite into schoenite, (iii) filtering the schoenite and separating the filtrate, hereinafter referred to as schoenite end liquor (SEL), (iv) desulphating the SEL with aqueous CaC12 containing filtrate generated in (xi) below to produce gypsum, (v) filtering the gypsum and mixing the filtrate hereinafter referred to as desulphated schoenite end liquor with MgC12-rich filtrate obtained from centrifuging crude camallite in step (vii) below, (vi) evaporating the resultant solution of step (v) and cooling to ambient temperature to crystallize out crude camallite, (vii) centrifuging the crude camallite and recycling of MgC12-rich filtrate to step (v) above, (viii) decomposing the crude camallite with an effective amount of water to yield crude potassium chloride and camallite decomposed liquor referred to as CDL, (ix) filtering the crude potassium chloride and washing it with water to remove adhering MgC12, (x) crude potassium chloride is subjecting to hot leaching for the production of Muriate of Potash (MOP) and sodium chloride, (xi) mixing the carnallite decomposed liquor (CDL) from (viii) and washing from (ix) and treating it with hydrated lime to form Mg(OH)2 slurry, (xii) filtering the slurry to recover and recycle the CaC12- containing filtrate for the desulphatation process of step (iv), and washing the resultant filter cake with water to produce Mg(OH)2 which on calcination produce high purity MgO, 8a (xiii) treating the schoenite produced in (iii) with Muriate of Potash (MOP) produced in (ix) to produce Sulphate of Potash (SOP) under ambient condition, (xiv) filtering the Sulphate of Potash (SOP) and collecting separately the mother liquor hereinafter referred to as KEL, (xv) recycling the Potassium sulphate mother liquor in the process of (ii) above.
8b It may be noted that certain steps of the above process are triggered initially with CaC12 and water procured externally and thereafter these are largely generated in the process of the invention as described above. ' In an embodiment of the present invention, bittern of density in the range of 34 13e' (sp. gr. 1.25-1.31) is used for production of mixed salt as described in the prior art and then converted into schoenite with simultaneous ,leaching of NaCI from the solid mass.
In another embodiment of the present invention, mixed salt is treated with a 0.3-0.5:1 ratio of water and KEL rich in KCl and MgSO4 and low in NaC1 and MgC12 to minimize loss of K from mixed salt without hampering transformation of kainite into schoenite and leaching of NaCl from the mixed salt.
In another embodiment of the present invention, schoenite is reacted with MOP
and water in the ratio of 1:0.3-0.6:1-2 to produce SOP.and KEL and wherein the MOP
is produced in situ from SEL.
In another embodiment of the present process, MOP is produced from carnallite which, in turn, is obtained through desulphatation of SEL, treatment with 400-440 g/L
MgClz liquor into the ratio of 1 part of desulphated bittern and 0.7-0.9 parts of MgC12 liquor, and forced evaporation till the solution attains a temperature of 120-128 C at atmospheric pressure.
In another embodiment of the present process, the filtrate obtained after removal of NaCl is cooled to room temperature wllereupon carnallite is obtained upon filtration while the filtrate contains 400-440 g/L of MgC12 and is recycled back into a fresh lot of desulphated SEL for further production of carnallite.
In another embodiment of the present process, the wet carnallite is treated with water in the ratio of 1:0.4-0.6 to obtain crude KCI.
In another embodiment of the present process, the magnesium chloride in carnallite decomposed liquor is supplemented with MgC12 in the end liquor and treated with lime to produce Mg(OH)2 and required quantity of calcium chloride solution (20-30% strength w/v)for desulphatation of SEL.
In another embodiment of the present process, the Mg(OH)2 is calcined in the temperature range of 800-900 C to produce MgO with < 0.04% B2O3.
In another embodiment of the present process, fresh water requirement is kept to a minimum by recycling water from forced evaporation step along with washing generated in the purification of gypsum, Mg(OH)Z and KC1.

In another embodiment of the present process, acidified de-brominated bittern, which is an ideal raw material for mixed production, is neutralized with crude Mg(OH)2 instead of with lime to eliminate sludge formation.

Detailed description The main inventive step is the recognition that the step of transforming kainite in mixed salt into schoenite and leaching of NaC1 from mixed salt can be simultaneously performed in a single operation with minimum loss of KC1 in mixed salt. Another inventive step is self reliance wherein the need for outsourced MOP is minimized by producing it instead from the waste filtrate of schoenite manufacture.
Another inventive step is the desulphatation of SEL required for MOP
production using calcium chloride generated in situ from the MgC12 in desulphated SEL that shows up as MgC12 -rich streams of carnallite decomposed liquor and end liquor. Another inventive step is the coupling of Mg(OH)2 production with desulphatation of SEL and thereby eliminating the problem of CaC12 waste management otherwise encountered in production of Mg(OH)2 from brine or bittern. Another inventive step is the use of CDL primarily for Mg(OH)2 production which greatly reduces BaO3 impurity in Mg(OH)2 and, as a result, in MgO obtained there from. Another inventive step is the local use of crude Mg(OH)2 for neutralization of acidified de-brominated bittern prior to production of mixed salt. Another inventive step is the recycling of liquid effluents to minimize requirement of fresh water while simultaneously enhancing recoveries and addressing the problem of effluent disposal.
The following examples are given by way of illustration and should not be construed to limit the scope of present invention.

In a typical process, 200 M3 of 29.5 Be' (sp. gr. 1.255) sea bittern was subjected to solar evaporation in a lined pan. The first fraction (20 Tons) containing mainly crude salt was removed at 34 Be' (sp. gr. 1.306). The bittern was further evaporated to 35.5 Be' (sp. gr. 1.324) and sels Mixt. fraction (15 tons) was separated.
The resultant bittern (100 M) was transferred to a second lined pan and solar evaporation was continued whereupon 16 tons of kainite type mixed salt and 26 M3 of end bittern were obtained. The mixed salt was further processed for production of schoenite as described in subsequent examples while a part of the end bittern was desulphated with outsourced calcium chloride to generate desulphated end bittern. A
part of the desulphated end bittern was subsequently treated with hydrated lime 'to produce calcium chloride and magnesiuin hydroxide. The calcium chloride solution was filtered and used for desulphatation of SEL of Example-6. The other part of the desulphated end bittern was used as MgC12 source in the same example to promote camallite formation from desulphated SEL. Similar experiments were also conducted with other sources of bittern such as sub-soil bittern and bittern obtained after bromine recovery.

142.0 kg of kainite type mixed salt, having chemical composition: KCl -15.5%, NaCl -14.6%,.MgSO4 -39.5% and, was treated with 140 L of water and stirred for 2.5 hr in a vessel. The slurry was filtered using basket centrifuge and yielded 32.0 kg of schoenite as solid product, analysing K2S04 - 38.0%, MgSO4 - 30.2%, and NaC1 - 1.2%, and 200 L of filtrate (SEL), analysing KC1 - 7.6, NaCl - 16.1%, MgSO4 -21.1 %, and MgC12 - 8.4%. The schoenite was treated with solution of 12.5 kg of MOP
in 49.0 L of water under agitation for 3.5 hr. The slurry was filtered to obtain 16.0 kg SOP, analyzing K2S04 - 95.0%, NaCl - 1.0 10, MgSO4 - 1.0%, and 60 L of filtrate (KEL) analyzing KCl - 15.0%, NaCl -1.5%, MgSO4 - 9.7%, and MgC12 - 3.9%.

60.0 kg of the mixed salt having the same composition as in Example-2 was taken along with the KEL obtained in Example-2. 27 L of water was additionally added and the contents were stirred for 2.5 hr. The slurry was filtered in a centrifuge to obtain 26.0 kg of schoenite analysing K2S04 - 39.7%, MgSO4 - 29.5%, NaCl -0.7%, and MgC12 - 0.6%, and 95.0 L of filtrate (SEL), analysing as KC1 - 9.9%, NaC1 -13.0%, MgSO4 - 18.6%, and MgC12 - 6.0%d. The schoenite was reacted with solution of 10.4 kg of MOP in 38 L of water in a vessel under stirring for 3.5 hr. The resultant sluny was filtered using centrifuge to obtain 14.5 kg SOP analyzing K2SO4 -98.1, NaCl - 0.2%, MgSO4 - 1.4%, and 45 L of filtrate (KEL) analysing as K2SO4 -12.4%, KC1- 6.15%, NaC1- 0.9%, MgSO4 - 1.0%, and MgC12 -10.2%, 104 kg of mixed salt analyzing KC1- 14.1 %, NaCl - 16.5%, MgSO4 - 41.6%, was reacted with 100 L of KEL analysing as K2S04-13.9 l0, NaC1- 2.8%, and MgC12 - 11.6%, and 40 L of water for 2 hr. The slurry was centrifuged to get 34.8 kg of schoenite analyzing K2S04 - 37.0%, MgSO4 - 30.3%, and NaC1- 4.9%,, and 190.0 L
of filtrate (SEL) analysing as KCl - 9.5%, NaCl - 13.0%, MgSO4 - 15.1% , MgCl2-8.0%, and. The schoenite was further reacted with a solution of 12.5 kg MOP in 46.0 L

of water for 3.5 hr to yield 17.5 kg SOP and 80 L of KEL. The SOP analysed as - 97.3%, NaC1- 0.2%, and MgSO4 - 3.0% and KEL as KC1 - 16.7%, NaCl - 1.3%, MgSO4 - 11.0%, and MgC12- 2.7%.

In this experiment 150.0 kg of mixed salt analysing as KCl - 13.1%, NaCI -19.8%, MgSO4 - 38.0%, MgC12- 1.9%, and was taken in a vessel along with 160 L
of KEL, analyzing KCl - 17.0%, NaC1- 3.3%, MgSO4 - 9.0% , MgCl2- 1.9%, and 60 L
of water and stirred for 2 hr. The resultant slurry was centrifuged to get 49.9 kg of schoenite analyzing K2S04 - 42.0%, MgSO4 - 32.2%, NaCl - 0.7%, and 255 L of filtrate (SEL) analysing as KCl - 10.5%, NaC1 - 12.3%, MgSO4 --13.7%, MgC12-6.70%. The schoenite was reacted with a solution of 19.0 kg of MOP in 75 L of water for 3.5 hr in a vessel with continuous stirring. The slurry was centrifuged to get 27.0 kg of SOP analysed as K2S04 - 94.3%, NaCl - 0.2%, and MgSO4 - 3.7%, and 85 L of filtrate (KEL), analysing as KCl - 15.5%, NaCl - 0.8%, MgSO4 - 10.5%, and MgC12-3.0%.

59 L of desulphated end bittern obtained in Exainple-1 having chemical composition: KCI - 1.15 %, NaCl 1.3%, MgCla - 41.2%, CaSO4 - traces was diluted with 40 L of water and treated with 14.7 kg of freshly prepared hydrated lime (87.7 %
active strength) for 1 hr. The resultant slurry was filtered and the cake was washed with 30 L of water. 90 L of total filtrate containing CaCl2-22.3% and MgC12-3.0% was obtained. The solid magnesium hydroxide was further washed with 100 L of water to malce it free from soluble impurities. 15.7 kg of Mg(OH)2 with 86.9 % Mg(OH)2 content was obtained on drying in a tray drier. A part of the Mg(OH)2 was calcined at 850 C yielding MgO of 90.0%. The 90 L of filtrate containing 22.3% CaC12 was used to desulphate 90 L of SEL obtained in Example-3. The resultant slurry was filtered to obtain 142 L of desulphated SEL and 21.0 kg of gypsum by-product. 57 L of desulphated SEL was mixed with 41 L of desulphated end bittern from Example-1 having Mg concentration of 10.3 %. The resultant solution was subjected to forced evaporation in an open pan evaporator till the solution attained a boiling point of 120 C. The hot liquor was filtered to separate 5.5 kg of crude NaCl having composition: NaC1- 85%, KC1-2.9% and MgCl2-12.1%. The filtrate was cooled in a tank to crystallize camallite. The resultant slurry was filtered to obtain 11.3 kg of carnallite analysing as KCI -21.7%, NaCl-9.7 %, MgCl2 -31.4%, and CaSO4 -2.7%, and 48 L of end bittern analyzing as MgC12 -40.2%, KCl - 0.8 %, NaCI - 1.1 %.
9.2 kg of carnallite was decomposed using 3.6 of water and filtered to get 8.0 L
of carnallite decomposed liquor (CDL) having chemical composition : KC1-4.6 %, NaCI-2.8% ; MgC12 -30.5%; CaSO4- traces, and 2.9 kg CDP having chemical composition: KCl-75.3%, NaCI-20.2%, MgC12-2.0% and CaSO4-2.5%. The CDP was treated with 1.9L of water at ambient temperature (30 C) to obtain 2.0 kg KCl having composition : KCl - 90.0 %, NaCI - 3.3% ; MgC12 - 0.4% and CaSO4 - 6.0 and 2.2 L
of saturated solution having chemical composition KCl - 14.0% and NaCl -20.0%.

Of 10 L of CDL obtained in above experiment, 5.7 L of cold leachate with which crude salt produced in the previous example was also washed to recover magnesium content in it, having chemical composition: KCl-7.0%, NaCI-8.2%, MgC12-21.5%, and CaSO4- traces, and 15 L of water was treated with 2.5 Kg of freshly prepared hydrated lime having 90 % activity for 1 hr. The resultant slurry was filtered and solid cake washed with 10 L of water to obtain 34 L of filtrate containing 7.7 %
CaC12. The solid magnesium hydroxide was further washed with 30 L of water to make it free from soluble impurities. The Mg(OH)2 was dried to obtain 2.3 Kg of Mg(OH)2 which was calcined to get MgO analyzing as 92% MgO containing 0.034%
B203 as impurity. 34 L of CaCla containing brine was used to desulphate 17 L
of SEL
having chemical composition KC1 - 7.2%, NaCI - 12.4 %, MgSO4- 16.0 %, and MgC12 - 6.5 % . The resultant slurry was filtered to remove 5.2 kg of wet calcium sulphate and obtain 49 L of desulphated SEL having Mg content of 2.03 %. 75 L
of end bittern having Mg concentration of 9.6 % obtained from previous experiment was added to the desulphated SEL. The resultant solution mixture was subjected to forced evaporation in open pan evaporator till the boiling point of the solution is 126 C. The hot liquor was cooled in a tank to crystallize carnallite. The resultant slurry was filtered to obtain 18.8 kg of carnallite having chemical composition: KCl - 14.3 %, NaCI -12.7 %, MgC12 - 31.9 % and CaSO4 - 1.9 % and 46.5 L of end bittern having chemical composition. MgC12 - 46.1%, KCl - 0.2 %, NaCl - 0.5 %. 18.8 kg of carnallite was decomposed using 8 L of water and filtered to get 15.5 L of CDL having chemical composition : KCl - 4.8 %, NaCI 3.2%, MgC12 - 32.5% and CaSO4 - traces; and 5.7 kg CDP liaving chemical composition: KCl - 33.9% and NaCl - 46.3%, MgC12-1.4%, CaSO4-5.1% and Moisture-13%. The CDP was subjected to hot leaching along with CDP obtained in the following example by known method to separate KCI as detailed below.

15.5 L of CDL obtained in above experiment having chemical composition:
KC1- 5.0%, NaCI - 3.2%, MgC12 - 32.5% and CaSO4 - traces; and 15 L of water was treated with 3.0 kg of freshly prepared hydrated lime having 90.0 % activity for 1 hr.
The resultant slurry was filtered and solids washed with 10 L of water to obtain 27.5 L
of filtrate containing 10.60 % CaC12. The solid magnesium hydroxide was further washed with 30 L of water to make it free from soluble impurities. The Mg(OH)2 was dried to obtain 2.9 kg of Mg(OH)2 and subsequently calcined to obtain caustic calcined MgO having 95% MgO content and 0.03 % B203 impurity. The CaC12 containing solution was used to desulphate 25 L of SEL having chemical composition KC1-7.2%, NaC1- 12.4 %, MgSO4 16.0 % and MgC12 - 6.5 %. The resultant slurry was filtered to remove 5.7 kg of calciuin sulphate and obtain 46 L of desulphated SEL
having Mg content of 3.05 %. 33 L of end bittern having Mg concentration of 11.8 %
obtained from previous experiment was added to the desulphated SEL. The resultant solution mixture was subjected to forced evaporation in an open pan evaporator till the boiling point of the solution is 125 C. The hot liquor was cooled in a tank to crystallize carnallite. The resultant slurry was filtered to obtain 14 kg of camallite having chemical composition: KCI - 15.0 %, NaCI - 24.7 %, MgCl2 - 25.1%, and CaSO4 - 4.0% and 33.8 L of end bittern having chemical composition. MgC12 - 44.8%, KCI - 0.1 %, and NaCl - 0.46 %. 14.0 kg of carnallite was decomposed using 6.3 L of water and filtered to get 12 L of CDL having chemical composition: KC1- 5.6 %, NaCI 4.4%; MgCl2 -27.6% and CaSO4 - traces; and 5.0 kg of CDP having chemical composition: KC1-26:1 % and NaCI-51.1 %, MgC1Z-7.1 %, CaSO4-5.1 % and moisture - 9.0%. The CDP
obtained along with CDP from Example-7 weighing 10.8 kg, was subjected to hot leaching by known method to obtain 3.5 kg of MOP having 93.6 % KCI content.

_ In this example, MOP produced in the above Example-8 was used to prepare SOP. 9.0 kg of kainite type of mixed salt analyzing as KC1- 14.2%, NaCI -16.5%, MgSO4 - 40.2% , MgCI2- 1.2%, was reacted with 8 L of water for 2 hr. The slurry was centrifuged to get 3.0 kg of schoenite analyzing as K2SO4 - 35.5%, MgSO4 -31.0%, and NaCl - 3.3%, and 9.5 L of filtrate (SEL) analysing as and KCI -7.6%, NaCI - 12.6%, MgSO¾ - 15.1%, MgCl2- 9.5%, 0.488 kg of schoenite was further reacted with the solution of 0.190 kg MOP (from obtained in above Example-8) in 0.753 L of water for 3.5 hr to yield 0.255 kg SOP and 0.860 L of KEL. The SOP
analysed as .K2SO4 - 93.0%, NaC1- 0.6%, MgSO4 - 5.4% and KEL as KCl - 14.8%, NaCI - 1.4%, MgSO4 - 7.7%, MgC12- 4.1%.

Claims (30)

Claims:

1. An integrated process for preparation of sulphate of potash (SOP) from bittern comprising:
(i) subjecting bittern to fractional crystallization to obtain kainite mixed salt with high kainite content and MgCl2-rich end bittern, (ii) treating the kainite mixed salt with water and mother liquor obtained as a filtrate during sulphate of potash preparation in (xiv) below, referred to as KEL, to leach out substantially all NaCl from the mixed salt and simultaneously convert kainite into schoenite, (iii) filtering the schoenite and separating the filtrate, hereinafter referred to as schoenite end liquor (SEL), (iv) desulphating the SEL with aqueous CaCl2 containing filtrate generated in (xi) below to produce gypsum, (v) filtering the gypsum and mixing the filtrate hereinafter referred to as desulphated schoenite end liquor with MgCl2-rich filtrate obtained from centrifuging crude carnallite in step (vii) below, (vi) evaporating the resultant solution of step (v) and cooling to ambient temperature to crystallize out crude carnallite, (vii) centrifuging the crude carnallite and recycling of MgCl2-rich filtrate to step (v) above, (viii) decomposing the crude carnallite with an effective amount of water to yield crude potassium chloride and carnallite decomposed liquor referred to as CDL, (ix) filtering the crude potassium chloride and washing it with water to remove adhering MgCl2, (x) crude potassium chloride is subjecting to hot leaching for the production of Muriate of Potash (MOP) and sodium chloride, (xi) mixing the carnallite decomposed liquor (CDL) from (viii) and washing from (ix) and treating it with hydrated lime to form Mg(OH)2 slurry, (xii) filtering the slurry to recover and recycle the CaCl2- containing filtrate for the desulphatation process of step (iv), and washing the resultant filter cake with water to produce Mg(OH)2 which on calcination produce high purity MgO, (xiii) treating the schoenite produced in (iii) with Muriate of Potash (MOP) produced in (ix) to produce Sulphate of Potash (SOP) under ambient condition, (xiv) filtering the Sulphate of Potash (SOP) and collecting separately the mother liquor hereinafter referred to as KEL, (xv) recycling the Potassium sulphate mother liquor in the process of (ii) above.

2. The process as claimed in claim 1 wherein the bittern of (i) is of density 34°Be' (sp. gr. 1.25-1.31) and contains on an average K ion = 1.0 to 1.5%, Mg ion = 4.0 to 5.5%, and SO4 ion = 2.5 to 6.0% suitable for kainite production.

3. The process as claimed in claim 2 wherein the bittern is selected from the group consisting of sea bittern, sub-soil bittern, debrominated bittern effluent and bittern with higher potassium content, which also requires the least evaporation to produce kainite mixed salt.

4. The process as claimed in claim 1 wherein kainite mixed salt contains KCl-22%, NaCl-15-22%, MgSO4-28-40%, MgCl2-5-10%.

5. The process as claimed in any one of claims 1 to 4 wherein one part by weight of kainite mixed salt is treated with 0.75-1.25 parts by volume of potassium sulphate mother liquor (KEL) obtained in (xiv) and 0.3-0.7 parts by volume of water.

6. A process as claimed in any one of claims 1 to 4 wherein potassium sulphate mother liquor (KEL) obtained contains typically 15-17% KCl, 1-3% NaCl, 10-12% MgSO4, and 2-3% MgCl2.

7. A process as claimed in any one of claims 1 to 6 wherein the schoenite end liquor (SEL) obtained in (iii) contains typically 8-10% KCl, 6-12% NaCl, 12-14% MgSO4 and 5-7% MgCl2.

8. A process as claimed in any one of claims 1 to 6 wherein the schoenite obtained in (iii) contains typically 40-45% K2SO4, 30-35% MgSO4 and 0.5-2.0% NaCl.

9. A process as claimed in any one of claims 1 to 6 wherein the stoichiometric ratio of CaCl2 to sulphate for the desulphatation of schoenite end liquor (SEL) in (iv) is 1.1:1 to 0.9:1.

10. A process as claimed in claim 9 wherein the stoichiometric ratio of CaCl2 to sulphate for the desulphatation of schoenite end liquor (SEL) in (iv) is 1:1.

11. A process as claimed in any one of claims 1 to 6 wherein 1 part by volume of desulphated schoenite end liquor is mixed with 0.5-1.5 parts by volume of MgCl2 rich end bittern of 36-38° Be' (sp. gr. 1.33-1.38).

12. A process as claimed in claim 11 wherein 1 part by volume of desulphated schoenite end liquor is mixed with 0.7-0.9 parts by volume of MgCl2 rich end bittern of 37° Be' (sp. gr. 1.342).

13. A process as claimed in claim 11 or 12 wherein the MgCl2 rich end bittern contains almost nil sulphate.

14. A process as claimed in any one of claims 1 to 13 wherein the evaporation and concentration of desulphated schoenite end liquor to produce carnallite is carried out either in a solar pan or in a multiple effect evaporator with simultaneous recovery of water.

15. A process as claimed in claim 14 wherein evaporation in the solar pan is continued till the solution attained Mg2+ concentration in the range of 10 to % and KCl concentration in the range of 0.4 to 0.6% when crude carnallite is crystallized out.

16. A process as claimed in claim 14 wherein evaporation in the multiple effect evaporator is continued till the solution attained a temperature in the range of 120-128°C.

17. A process as claimed in claim 14 wherein evaporation in the multiple effect evaporator is continued till the solution attained a temperature in the range of 122-124°C.

18. A process as claimed in any one of claims 1 to 14 wherein the carnallite obtained has 15-20% KCl, 15-20% NaCl and 28-32% MgCl2.

19. A process as claimed in any one of claims 1 to 18 wherein one part by weight of the carnallite is decomposed with 0.4-0.6 parts by volume of water to yield crude potassium chloride and carnallite decomposed liquor (CDL), followed by washing of the crude potassium chloride with a small quantity of water.

20. A process as claimed in any one of claims 1 to 18 wherein the molar ratio of hydrated lime to MgCl2 of carnallite decomposed liquor (CDL) is in the range of 0.8-1.0, for the production of Mg(OH)2 with simultaneous recovery of CaCl2 containing solution for desulphatation of SEL.

21. A process as claimed in claim 20 wherein the molar ratio of hydrated lime to MgCl2 of carnallite decomposed liquor (CDL) is 0.90.

22. A process as claimed in any one of claims 1 to 18 wherein the Mg(OH)2 obtained from carnallite decomposed liquor (CDL) is calcined to produce MgO with 94-98 % purity and with 0.02-0.04 % B2O3.

23. A process as claimed in any one of claims 1 to 22 wherein the Muriate of Potash obtained in (x) has purity in the range of 92-98% KCl and NaCl content of 1-5 %.

24. A process as claimed in claim 23 wherein the Muriate of Potash obtained in (x) has purity in the range of > 95% KCl and < 2% NaCl.

25. A process as claimed in any one of claims 1 to 22 wherein the NaCl obtained in (x) has > 97% NaCl.

26. A process as claimed in any one of claims 1 to 24 wherein one part by weight of schoenite is treated with 0.3-0.6 parts by weight of Muriate of Potash and 2 parts by volume of water, in the ambient temperature range of 20-45 degree C to produce Sulphate of Potash.

27. A process as claimed in any one of claims 1 to 24 wherein one part by weight of schoenite is treated with 0.4 parts by weight of Muriate of Potash and 1.5 parts by volume of water, in the ambient temperature range of 20-45 degree C
to produce Sulphate of Potash.

28. A process as claimed in any one of claims 1 to 27 wherein Sulphate of Potash is filtered and potassium sulphate mother liquor (KEL) is separately collected and recycled in (ii) above.

29. A process as claimed in any one of claims 1 to 27 wherein the Sulphate of Potash (SOP) produced has K2O content in the range of 50-52% and chloride in the range of 0.5-2.0 %.

30. A process as claimed in any one of claims 1 to 27 wherein by utilizing all intermediate streams obtained between (ii) to (xiv), the recovery of potash in the form of Sulphate of Potash (SOP) is > 90%.