- Conventional mining
- Solution mining
- Beneficiation and processing of potash ores and brines
- Crushing and grinding
- Heavy media separation
- Electrostatic separation
- Thermal dissolution-crystallisation
- Potash recovery from carnallite
- Beneficiation of carnallite ores and brines
- Disposal of brines and tailings
- Beneficiation of sulphate ores
Potash is normally found in underground deposits, 700-1,400 metres deep. Potash ore contains generally between 20 and 35% KCl, the remainder being NaCl, Mg-salts, etc. There are two basic types of potash mining: conventional mining, which encompasses several standard methods, modified as necessary for specific situations, and solution mining.
Each potash deposit has its own characteristics and mining challenges that must be handled on an individual basis. However, many standard methods have become widely accepted within the potash mining industry. For potash beds that are relatively flat and uniform in thickness, or for cutting entries, boring machines with two or four cutting arms have proven to be the most effective and economical mining method. For moderately inclined, undulating and/. or thickening-thinning potash seams, continuous miners with cutting “drums” mounted on one or two moveable arms are the most effective. For highly variable ore bodies, or ore with potential rock-bursts or other special needs, drilling and blasting are still required.
Most mine configurations are the conventional room and pillar design, except that “stress relief mining” (multiple entries with outer yield pillars) has proven highly effective for the deeper mines with plastic flow (of the ore) or unstable roof conditions. For thicker or highly inclined potash seams, the open slope (successive horizontal cuts for thick beds) or cut and fill methods allow both a high ore recovery and an easy means of disposing of both wastes produced underground and plant tailings.
For those mines without high-pressure aquifers above the potash and with somewhat plastic overlying strata, the long-wall mining method, which allows the roof to descend soon after the ore is cut, can provide very high ore recovery rates and result in high productivity. In the case of long-wall mining, the beds must be uniform and continuous (very minor faulting, etc.) over the width and length of individua mining panels. Although long-wall mining can produce ores at relatively low costs per tonne, equipment and development costs can be high and the method is not as flexible as other methods.
Underground ore transport is now accomplished almost entirely by conveyor belts. Conveyor belts may be suspended from the roof or set on the floor. For the highest capacity mines with continuous (borer or drum) mining machines, flexible belt conveyors attached to or directly fed from the miner and directly feeding a panel belt have produced considerable savings. Some mines employ either shuttle cars to take the ore from the mining machines to the feeder-breaker at the panel belt, or load-haul-dump (LHD) machines that pick up ore from the floor and then haul it to a feed station. Most mines have considerable underground ore storage capacity to even out the flow, as well as feed bins where the ore is loaded into high-speed “skips” (elevator bins) that take the ore to the surface and dump it, with all aspects of the loading-lifting-dumping cycle being done automatically. A few shallower mines bring the ore to the surface with conveyor belts in long inclined tunnels (ramps).
The costs of potash mining by conventional methods can vary tremendously. Mining is the major expense of potash production in operations that require more manpower because of low-grade ore or difficult mining conditions. For ore bodies and situations that allow high-capacity operations, mining can be somewhat less expensive than the processing of potash ores at the surface.
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This type of mining is mainly found in Canada and the United States. The most recent mine of this type brought into operation is the Bethune potash mine from K+S in Saskatchewan, with a nominal capacity of 2 mln t/a KCl.
The process involves dissolving underground water-soluble minerals with water, resulting in a solution commonly referred to as brine. The water is then extracted from underground and the dissolved minerals are recovered from it. Basically there are two extraction flows: the primary being fresh water dissolving potash from the rock, the secondary being a saturated NaCl solution to selectively dissolve KCl from the walls and the roof of the caverns.
The major BHP-owned Burr project, located near Saskatoon, is under construction (2021). The aim is to produce 4 mln t/a KCl, using the same solution mining technology.
Solution mining is also a method to be applied on mines in the last stage of their lives, where conventional mining gets too expensive and solution mining can recover the last patches of valuable material out of the mine (for instance from the pillars that conventional mining cannot take out).
Also, examples are known from mines that were accidentally flooded by underground aquifers and were turned into solution mining operations.
Beneficiation and processing of potash ores and brines
The beneficiation of raw potash ore into marketable products, compared to potash mining operations, requires a somewhat different scheme and equipment for each processing plant. The ore composition (mineralogy), ore grade (K2O content), liberation size (the maximum particle size that allows separation of the different minerals), and type and quantity of slimes (insoluble such as clay, anhydrite, dolomite, silica,) are different for every deposit, and possibly highly variable within a single deposit. Accordingly, the modern potash processing plant must have an efficient, highly versatile process to accommodate variations in the feed yet still maintain product quality at as low a production cost as possible.
Four basic beneficiation techniques are used in the potash industry: flotation, heavy media separation, electrostatic separation, and thermal dissolution crystallisation. Other techniques have also been experimented with. These include ammoniacal leaching and gravity separation using shaking tables. Ammoniacal leaching has proven too costly and gravity separation can only produce impure concentrates (about 80% KCl). Most potash processing plants use a combination of beneficiation techniques. Each method will be reviewed separately in the following sections.
Crushing and grinding
The first step in the beneficiation of a solid potash ore is to reduce the ore to a size where the potash is liberated from the other ore constituents and can be separated. A prime consideration in crushing and grinding is that a minimum amount of fines be produced. Any fines produced by grinding must be processed by more costly methods. Larger and closely sized product is more valuable, and less energy is spent on grinding. Another prime consideration is that both the capital costs involved in establishing the operation and the production costs be minimised. To accomplish these goals, processing usually begins with various size reduction equipment and coarse (0.6-1. 2 cm) screens working in closed circuit (recycling the oversize) because coarse ore crushing and screening are relatively simple and less expensive than other steps. There are many equipment choices, but most modern plants start out with dry screening. The oversize is recycled to an impact crusher, and the undersize is sent to other types of size classification equipment.
The second and third stages of grinding and classification involve even greater choices of equipment but are relatively more expensive. For small liberation sizes (usually between 100- and 150-mesh, Tyler), wet ball or rod mills are required. For larger liberation sizes (plus 100-mesh, Tyler), wet or dry impactors or cage mills are used. The oversize closed-circuit separators are either screens or, if wet processing is utilised, screens and rake, or screw classifiers. Undersize separators (usually to split at 100-200-mesh, Tyler) are typically wet cyclones and/or hydro classifiers.
Fine particles, both ore minerals and insolubles, are always removed to the greatest extent possible before the potash separation step. In flotation, fine particles with high surface areas tend to adsorb excessive quantities of flotation reagents. Excessive amounts of fines can significantly raise reagent costs and cause contamination of the product. Fine gauge particles may float or become mechanically entrapped in the product. The ore slurry is first “attrition scrubbed” to liberate and disperse the fines. The equipment used varies widely, in that any form of intense agitation is effective in slimes removal. In a few plants, some or all of the slimes are separated by selectively floating the slimes from the sylvinite (Figure 1). However, most plants deslime by using one or more stages of cyclones and hydro separators (Figure 2). A liberation size of about 150-mesh has been determined to provide the best cut-off range for flotation of most potash ores. Many plants (especially those employing fine grinding) make a second size classification at 200- to 400-mesh to reduce potash losses. These fine solids are then floated in a fines flotation circuit, and the impure float is usually sent to a hot leach-crystallisation plant for further processing. In several plants, the slimes-slurry is hot leached to further recover potash (Figure 3). Whatever treatment is used, the residual slimes are finally thickened and sent to tailings areas for disposal. If regrinding of coarse flotation tails is practiced, these solids are also deslimed again before being re-floated.
Flotation is a selective beneficiation process that utilises the differences in surface properties of various minerals. By conditioning ores with specific reagents, selected minerals can be induced to become either hydrophobic (water repellent) or hydrophilic (water attracting) in solutions. If a solution is then agitated and aerated by introducing air bubbles, mineral particles that are hydrophobic will preferentially attach themselves to the air bubbles and float to the surface where they can be removed. The waste (gangue) materials in the pulp at the bottom of the cells are disposed of as tailings. This type of flotation, where the valuable minerals are removed in the froth, is termed direct flotation and is the most common flotation technique employed in the potash industry.
The most important reagents used in flotation are collectors. In the case of flotation to treat sylvinite ores, a cationic collector is added to the closely sized and deslimed ore slurry. These collectors are mostly straight-chain aliphatic primary amines derived from natural fats and oils that are neutralised by acetic or hydrochloric acid before use. The polar end group of the reagent is selectively adsorbed on potash, rendering its surface hydrophobic. The salt (NaCl) and other gangue minerals sink in the pulp and are removed from the bottom of the flotation cell.
Another category of reagents that is commonly used is known as frothers. Frothers are used to aid the formation of, and stabilise, the flotation froth. Generally, these reagents are organic heteropolar compounds. Pine oil, a widely used frother, contains aromatic alcohols. A wide range of synthetic frothers is also available. A conventional frother, MIBC (methylisobutyl carbinol), is used in potash flotation as a modifier to inhibit the formation of excessive amine froth.
Modifiers or regulators constitute another class of flotation reagents that are used to control the process. In potash flotation, slimes depressants such as starch, guar gum, dextrin, and synthetic compounds are typically used.
In potash flotation the objective is usually to separate sylvite from halite by using cationic collectors. When the potash ores contain sulphate minerals like kieserite, sulphate salts can be floated by using fatty acids or sulphated fatty acids as collectors. Combined salts such as kainite (KCI•MgSO4•3H2O) can be floated by using coconut amine, as performed in Sicily, Italy. Schoenite (K2SO4•MgSO4•6H2O), processed from Great Salt Lake, U.S.A., brines, can be floated by using coconut fatty acid. Ores containing carnallite may be pre-treated to remove magnesium chloride and give simple potassium chloride and halite, which then can be floated with tallow amine.
Potash flotation circuits always have both rougher and cleaner cells. In the rougher cells, an attempt is made to float as much potash as possible, while in the cleaner cells the purity of the floated product is the prime consideration. The coarse particles of the rougher tailings and the tailings from the cleaner cells are usually reground to a smaller size range and re-floated. In situations where the potash ore is liberated at coarse sizes, the coarse size fraction is conditioned with reagents and floated separately from the fines, thus decreasing reagent consumption and generally improving recovery (Figure 4). With very fine ores, some companies use counter current flotation because the use of traditional flotation cells leads to excessive mechanical entrainment of the salt (Figure 2). In this case, the tailings from each stage are returned to the previous stage and the float advances to the next stage, both to be repeatedly re-floated. If the fine-floated potash product is not quite up to grade, it may be hot or cold leached to dissolve halite and to increase the purity,
Heavy media separation
Heavy media separation is a useful beneficiation process as a preliminary scalping operation or in combination with other techniques. The process utilises the difference in specific gravity of sylvite (KCl) and halite (NaCl). Halite is the denser mineral (specific gravity 2.13 versus 1.9 for sylvite). In a liquid of intermediate specific gravity, halite will sink and sylvite will float. Commercial heavy media operations use a very finely divided weighting agent, typically ferrosilicon or magnetite of minus 200- mesh, which is slurried to create an artificial heavy medium at the specific gravity required for separation. After separation, the magnetite or ferrosilicon is recovered by magnetic separation and recirculated to the system.
Electrostatic separation is a dry technique in which a mixture of minerals may be differentiated according to their electrical conductivity. For potash minerals, which are not naturally conductive, the separation must be preceded by a conditioning step that induces the minerals to carry electrostatic charges of different magnitudes and, if possible, different polarities. For potash, fractional or triboelectric charging is used; the charges are induced through repeated physical contact between the different minerals.
In Germany, the potash ore is ground to between 1 and 2 mm. The ground ore is conditioned by one or more reagents, preferably aromatic and aliphatic monocarboxylic acids. The mixture is then heated in a fluidised bed, and the relative humidity is adjusted to enhance charging of the particles. The ore is then fed into the electrostatic separator, to yield three fractions, i.e., product, residue, and middlings. The middlings, after further grinding, are recycled to the conditioning stage.
Using this process provides two main advantages in terms of costs and environmental impact. By using a dry pre-enrichment step in the mine (underground), hoisting costs are reduced because only valuable preconcentrated ore is moved to the surface. Thus, the process decreases the environmental impact of disposing of salt tailings on the surface or the costs of returning salt tailings to the mine. Using a dry process to improve product grade also eliminates drying costs.
Thermal dissolution-crystallisation is possible because potassium chloride is much more soluble in hot water than in cold, and sodium chloride is only slightly more soluble at 100°C than at 20°C. In saturated solutions containing both salts, sodium chloride is actually less soluble at higher temperatures. When a brine that is saturated with both salts at 20°C is heated to 100°C, it is capable of dissolving substantial additional amounts of KCl but not NaCl.
A typical dissolution-crystallisation flow diagram is illustrated in Figure 5. Sylvinite ore is crushed to minus 3-mesh and washed with cold, saturated NaCl-KCl solution. Clays are removed from the solution by desliming, and the clarified solution is then heated and used to dissolve the potash in the washed ore. Undissolved NaCl is discarded as tailings. The brine solution is then cooled under vacuum, and the KCl crystallises out and is separated, washed, and dried. The remaining brine is recycled. When a high-purity KCl for chemical use is required, the crystalline KCl is redissolved and recrystallised to produce a double-refined product containing more than 99.9% KCl.
Commercially, the thermal dissolution-crystallisation process is used in Germany by K+S at various refineries. In eastern Europe and the former Soviet Union, this process is used to treat potash ores that contain high levels of clays and magnesium chloride levels that exceed 30% MgCl2. The main disadvantage of this process is the high energy consumption required to heat all the mother liquor. Corrosive hot brines must be handled, which results in higher plant maintenance costs and requires the process equipment to be manufactured of special high-cost materials. Although thermal dissolution-crystallisation is used only in solution mining, a wider application would be the extraction of potash from brines.
Potash recovery from carnallite
The conventional mining and beneficiation sequence for potash recovery from carnallite encompasses the operations of mechanically mining the ore, hoisting ore to the surface, beneficiating the ore to a carnallite concentrate, dissolving the carnallite, recrystallising potassium chloride, disposing of saline solid wastes, and disposing of saline liquid effluents. Carnallite ore processing methods vary widely and depend upon accessory minerals that are associated with the ore.
There are two general methods of carnallite processing: cold leaching and hot leaching. Cold leaching processes are carried out at 20°-25°C. Carnallite ore is processed to recover a liquor saturated in magnesium chloride, a coarse sodium chloride solid product, and a fine residue of sylvite and kieserite. Initial grinding of the ore determines the eventual purity of the coarse and fine decomposition products. The temperature used to decompose the carnallite is not critical but must be below 25°C to prevent sulphate from remaining in the magnesium chloride liquor. Decomposition of the ore is carried out using unsaturated brine recirculated from a later filtration stage. To prevent reprecipitation of carnallite, the decomposition liquor is allowed to reach only 90% saturation in magnesium chloride. The fine decomposition product contains sylvite and kieserite and meets the specification for some potash-magnesium fertilisers (46% potassium chloride and 45% magnesium sulphate). However, flotation can be used to separate kieserite from sylvite.
Beneficiation of carnallite ores and brines
Carnallite ores are the source of only a small percentage of the world’s potash supply, even though there are large deposits of carnallite-containing ores in many regions. One disadvantage of carnallite ores is their low grade. Pure carnallite (KCl•MgCl•6H20) contains only 17% K2O. Carnallite cannot be used as a direct-application potash fertiliser because it is deliquescent. Dissolution and recrystallisation methods must be used to process carnallite ores, and these methods are energy intensive and expensive. A large volume of by-product magnesium chloride solution is produced in carnallite processing, which is likely to pose a disposal problem. Despite many disadvantages, ores or brines containing carnallite are processed to produce KCI in several countries including Germany, Israel and Jordan.
Carnallite has been mined, beneficiated, and processed in Germany for approximately 130 years. These operations were economically sustainable only as long as the German mines supplied a very significant portion of the world potash market. As a result of international competition based on the production from sylvinite ores, carnallite processing plants in Germany that only produced KCl were forced to close and potash production in Germany was concentrated at mines with significant amounts of sylvinite ores.
Carnallite ores may also be treated by a hot leach at about 100°C to dissolve MgCl2, KCI, and any NaCl that may be present. The hot solution is then clarified to remove solid impurities and then evaporated and cooled. After the KCl and NaCl crystals are recovered, they are separated by one of several beneficiation techniques.
In Germany, at the Hattorf and Wintershall mines, carnallite ore is first ground to less than 4 mm. The ore is then leached, and the carnallite decomposes. Saturation in potassium chloride is not reached, and only sodium chloride and kieserite remain as solids. The solids and brine are separated by filtration; the brine is cooled in vacuum crystallisers, and KCI is precipitated. The 40% K20 product is upgraded by the addition of a liquor from the potassium sulphate section of the refinery. On centrifuging, a 60% KCl product is recovered.
Processing of the Dead Sea brine by the Arab Potash Company· (APC) in Jordan traditionally utilises solar evaporation and the hot leach process to produce potash. However, APC’s 2021 expansion involves processing by flotation followed by cold crystallisation. The Dead Sea brine typically contains 11.5 g/l KCl. The concentration of other salts is: MgCl2 130 g/l, NaCl 87 g/l, CaCl2 37 g/l, and MgBr2 5 g/l. Brine is pumped from the Dead Sea into a 10-km-long gravity canal feeding a series of solar ponds. The solar ponds are divided into three basic sections: the salt ponds, the pre-carnallite ponds and the carnallite ponds. Sodium chloride is precipitated in the salt ponds. The pre-carnallite ponds are used to regulate the concentration of the brine that is fed to the carnallite ponds as close as possible to the point at which carnallite precipitation begins. Carnallite is recovered with a specially designed floating-track harvesting machine. A screw-type cutterhead harvests the 30- to 50-cm-thick carnallite layer. The carnallite slurry (20%-25% solids) is cycloned to reduce the water content and pumped at 40% solids to the processing plant.
The slurry is first wet screened to separate the large high-grade carnallite crystals, which are fed to cold crystallisers. Undersize slurry from the screens is mixed with brine discharged from the cold crystallisers. Additional carnallite is then crystallised, and the slurry is thickened. The thickener overflow is returned to the evaporation ponds. The underflow is subjected to salt flotation. Floated NaCl slurry is pumped to the tailings area. The carnallite sink fraction is settled in a thickener. Thickener underflow is dewatered in centrifuges. The cake from the thickener underflow is conveyed to the cold crystallisers where carnallite is decomposed by the addition of water and recycle brine. Crystalliser discharge slurry is wet screened to remove large particles. Crystalliser screen undersize slurry is subjected to leaching in agitated pans to dissolve any remaining sodium chloride. The leach tank discharge slurry is transported to a thickener, then to two stages of dewatering centrifuges, and finally to a rotary dryer. The dried potassium chloride product passes through screens to be separated into two fractions, standard and fines, according to the required size specifications. The products are collected and stored. A typical block diagram for the cold crystallisation process is shown in Figure 6.
Disposal of brines and tailings
A major problem in all plants that process potash ores is the disposal of excess brine and tailings. Until recently, plants merely stacked their waste salt and slimes and injected the excess brine underground or dumped both into the nearest river. New plants have been required to eliminate these practices. Some potash operations in Canada run a much tighter water balance in their plants and evaporate excess brine. This results in a 5-15% higher yield than in other potash plants. Tailings containing small amounts of residual MgCl2 brine are mechanically or hydraulically backfilled into the areas previously mined. The tailings are compacted as equipment is driven over them to make the next cut in inclined potash ore bodies. This process of dewatering and backfilling costs more than surface impoundment or river disposal, but greater potash recovery and other savings help to defray additional costs. Besides that, this backfilling is more sustainable than river or surface disposal.
Beneficiation of sulphate ores
The processes for beneficiating langbeinite, alunite, and complex sulphate ores into marketable products are discussed in this section.
Langbeinite (K2SO4•2MgSO4) is separated from sylvite and halite by selective washing, froth flotation, or heavy media separation. Langbeinite may be marketed as fertiliser or animal feed as a source of K; Mg, and S, or it may be converted to other products. Potassium sulphate is produced by reacting one molecule of langbeinite with four molecules of KCl. Potassium magnesium sulphate is produced by refining langbeinite through the reaction with KCl. The reactions to produce potassium sulphate and potassium magnesium sulphate are as follows:
K2SO4•2MgSO4 + 4KCl → 3K2SO4 + 2MgCl2
2(K2SO4•2MgSO4) + 2KCl → 3(K2SO4•MgSO4) + MgCl2
A flowchart of the production of potassium sulphate from the Great Salt Lake, U.S.A., is shown in Figure 7. Water from the lake is pumped and distributed to shallow ponds covering an area of 19,200 acres. Solar evaporation causes the brine to reach saturation with sodium chloride, and the salt settles on the pond floor. Further evaporation of the brine in subsequent ponds results in the precipitation of kainite, sylvite, and carnallite, which are harvested from the ponds for further processing. Fractional crystallisation is used to separate the final potassium sulphate product from the other minerals.
Alunite (K2•Al6(OH)12•(SO4)4) is a potential source of potassium sulphate, alumina, and sulphur dioxide by-product. The economics of production hinge on the value of the alumina. Processing involves ore comminution, roasting, and leaching to recover potassium sulphate solution. Filtered solids can be processed in a Bayer-type process plant for alumina recovery. (The Bayer process is the principal industrial means of refining bauxite to produce alumina (aluminium oxide) and was developed by Carl Josef Bayer. In the Bayer process, bauxite ore is heated in a pressure vessel along with a sodium hydroxide solution (caustic soda) at a temperature of 150 to 200 °C. At these temperatures, the aluminium is dissolved as sodium aluminate (primarily [Al(OH)4]−) in an extraction process. After separation of the residue by filtering, gibbsite is precipitated when the liquid is cooled and then seeded with fine-grained aluminum hydroxide crystals from previous extractions.)
Production of 1 million t/a of alumina (Al2O3) would require about 10 million tonnes of alunite ore and would yield about 500,000 tonnes of potassium sulphate and nearly 800,000 tonnes of sulfuric acid. Energy costs and the utilisation of the coproducts limit the potential for the production of alumina or potassium sulphate from alunite.
Complex mixtures of potash ores may contain any or all of the following minerals: anhydrite, epsomite, halite, kainite, kieserite, langbeinite, polyhalite, and sylvite along with clays. Extraction of potassium salts from such ores becomes quite complicated because the mineralogical forms of the various components can be unstable. Whenever potassium sulphate is the desired product, free sylvite and kainite must be present in the same molecular proportions. Ideally the reaction is as follows:
KCl + KCl•MgSO4•3H2O → K2SO4 + MgCl2 + 3H2O
When the ore is deficient in sylvite, schoenite is produced:
2KCl•MgSO4•3H2O → K2SO4•MgSO4•6H2O + MgCl2
Potash ores from the Carpathian deposits in the Ukraine contain several minerals, and processing involves grinding and hot leaching with a “synthetic kainite” solution. The leaching process dissolves epsomite, kainite, and sylvite. Langbeinite, halite, and polyhalite are insoluble in this liquor. Clay remains suspended until it is clarified. As the liquor begins to cool, potassium salts begin to crystallise out of solution. Normally, potassium chloride and sodium chloride would precipitate first, but their precipitation is inhibited by the addition of a solution that is saturated with magnesium and potassium sulphate. Under these circumstances, only schoenite will precipitate:
2KCl + 2MgSO4 + 6H2O → K2SO4•MgSO4•6H2O + MgCl2
The precipitated schoenite is thickened and filtered, and part of it is mixed with water to produce potassium sulphate:
K2SO4•MgSO4•6H2O → K2SO4 + MgSO4 + 6H2O
The potassium sulphate is removed as a product by centrifugation. The hot liquor is recycled to the settling tanks, and the clear solution is recycled to the dissolving tanks.
During the conversion of kainite to schoenite, magnesium chloride is also formed. The concentration of magnesium chloride in solution is controlled by withdrawing a part of the schoenite filtrate, which is evaporated to remove enough potassium and sodium to maintain a constant concentration of these materials in the system. This process permits the use of a complex ore to produce potassium sulphate, schoenite, and sodium sulphate and magnesium chloride liquors.
Anonymous. 1986. “Beneficiating Potash to Commercial Product,” Phosphorus and Potassium, No. 145 (Nov.-Dec.): pp 30-36.
Foot, D. G., Jordan, C. E. and Huiatt, J. L. 1980. “Direct Flotation of Potash from Carnallite,” U.S. Bureau of Mines Report of Investigations 8678.
Anonymous. 1992. ”Potash Extraction by Flotation,” Phosphorus and Potassium, No. 182 (Nov.-Dec.): pp 20-26.
Medemblik, L. 1991. “Treatment of Clay Suspensions in MDPS’s Potassium Chloride Flotation Plant,” In Preprints of the XVII International Mineral Processing Congress-PP. 104-120, Dresden, Germany.
Anonymous. 1983. “Mining and Beneficiating Potash,” Phosphorus and Potassium, No. 128 (Nov.-Dec.): pp 26-31.
Larmour, D. 1983. “Electrostatic Separation of Potash-PCS Mining Experience,” Paper presented at the First International Potash Technology Conference, Saskatoon, Canada.
Breniaux, J., and Marx, J.P. 1991 “Equipment of a Long-Life Panel at MDPA,” Paper presented at “Kali 91,” Second International Potash Technology Conference, Hamburg, Germany, May 26-29, 1991.
Duchrow, G., Fitz, I., and Grushow, N. 1990. “Solution Mining Provides the Solution to an Economic Problem, Possibilities for Profitable Carnallite Extraction in East Germany,” Phosphorus and Potassium, No. 167 (Sept.-Oct.): pp 20-25.
Amira, J., Stanley, W., Armstrong, H. and Frimodig, M. 1992. “From Hot to Cold Crystallization,” Phosphorus and Potassium, No. 182(Nov.-Dec.): pp 28-32.
Anonymous. 1992. “Optimization of Solar Evaporation Systems,” Phosphorus and Potassium, No. 179(May-June): pp 17-25.
Anonymous. 1972. “Technical Advances by the U.S.S.R. Potash Industry,” Phosphorus and Potassium, No. 53(May-June): pp 52-55.
Links to Related IFS Proceedings
536, (2004), Underground Disposal of Process Waste in a Potash Mine – System Development and Installation, M. J. Wilkins, M. J. Fehrsen, M. E. Dodds-Smith
667, (2010), Land Remediation and Workforce Redeployment Resulting from the Ending of Potash Mining in France, R Giovanetti
780, (2016), Evaporation / Crystallisation Challenges in the Fertiliser Market, K. Schooley, V. Bourgier, R. Lawson
857, (2021), Mineral Sizing Journey from Pit to Port and its Influence on Fertiliser Products, R McConnell
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