Original Research Article

An Integrated Process for Producing Potassium Sulfate from Potash Mining Wastes

Dr. Hussameldin Ibrahim,
Michael Fabrik, Andrea Echeverría, Paul Kindermann and Hussameldin Ibrahim*
Industrial Systems Engineering, Faculty of Engineering and Applied Science, University of Regina, 3737 Wascana Parkway,
Regina, SK, S4S 0A2, Canada.

In light of growing world demand for potash, increasing production and competition, and the potential effect on increased waste production by potash refineries, a study was completed to investigate the technical and economic feasibility of producing potassium sulphate (K2SO4) from potash wastes.

*Corresponding author:

Hussameldin Ibrahim,
hussameldin.ibrahim@uregina.ca

Keywords:

Potassium sulfate, economics, glaserite, nanofiltration, custom-built simulation.

In light of growing world demand for potash, increasing production and competition, and the potential effect on increased waste production by potash refineries, a study was completed to investigate the technical and economic feasibility of producing potassium sulphate (K2SO4) from potassium chloride (KCl) as an alternative fertilizer product from such potash wastes. Experiments were conducted to study the crystallization of glaserite and K2SO4 from KCl and sodium sulphate (Na2SO4). It was found that glaserite crystallization occurs at high NaCl-concentrated solutions while K2SO4 crystallization only proceeds in pure KCl solutions. Also, a custom-built process simulation was developed and used to evaluate the proposed two-stage crystallization process integrated into an existing potash mine. Simulation results indicate that the proposed process was technically feasible with a net present value of $42 million, internal rate of return of 27% and payback period of 3.7 years.

Growing world population, combined with the increased dietary requirements from the expanding wealth of developing nations, has placed greater productive demands on food production. To accommodate the challenge of producing more food with the same amount of land, fertilizer producers – particularly those involved with potash mining – have expanded their operations in recent years. Saskatchewan is the largest potash-producing region in the world, meeting over half of the total demand for potassium-based fertilizers (Potash Corp. 2010). The industry has entered a phase of fervent supply growth, with massive expansion projects in existing mines and construction of multiple new mines occurring mainly in Saskatchewan. New venture companies and multinational mining companies were presenting new challenges for an already-competitive industry, where expansion of production and improvement of product quality is becoming increasingly significant priorities. Addressing productivity and material losses of potash mines will be a key element in ensuring the competitiveness of Saskatchewan’s potash industry. Material losses in potash drying and sizing operations are currently addressed by recycling dust into potash granulation systems, which reduces the overall quality of the product or, by dissolving the potash and returning the brine to fluid-handling systems, which results in increased energy costs and in a decrease of the plant’s capacity. A superior alternative is to use waste brines for producing value-added potassium sulphate by taking advantage of Saskatchewan’s massive sodium sulphate reserves, which are the largest in the world (Hammer 1986). Upgrading potash-mining wastes allows producers to improve their gross capacity while diversifying their production into a high-margin market.

Muriate of potash (MOP), known chemically as KCl, is the most significant source of potassium available to agriculture. KCl is mined from underground salt formations, and can be produced from either conventional shaft mines or solution mines. Conventional mines remove material directly from the ore body and transport it to surface facilities that crush the rock to pass it through flotation cells. Solution mines dissolve the ore and transport it to the refinery facilities as brine, where fractional crystallization is used to first remove the sodium chloride (NaCl) as a waste, and then KCl as a product. The KCl must then be dewatered and dried, sized and, ultimately, sent to market. A major source of material and energy waste within the previously described processes is the handling and reprocessing of KCl dust and undersized product. Despite best efforts, there is always a fraction of KCl product that is undersized and, therefore, unable to be sold. This particular size fraction of KCl product consists of dust and extremely small particles that fluidize and aerate when passed through compaction systems, impairing the process and reducing product quality. Given the random nature of ore crushing, and of crystallization, elimination of this small size fraction of KCl product is impossible. Also, KCl produced from solution mining is especially susceptible to weathering, since it does not have the advantage of clay and salt impurities that aid in maintaining particle strength. The overall cost of the KCl waste dust to producers, in terms of lost production and sunk processing costs, is significant. Additionally, refinery capacity to process raw brine is reduced, and its overall efficiency and profitability is impaired. Reducing or entirely avoiding the recirculation of KCl waste dust into the refinery process would grant a more flexible and profitable operation for a KCl mine.

Amidst the production expansion occurring in potash mining, many smaller businesses rely on converting KCl into alternative forms of potassium fertilizer, the most common of which being K2SO4. While the market is smaller, K2SO4 is favorable for chloride-sensitive crops. Additionally, most of the world’s K2SO4 supplies are produced through the Mannheim processes that react KCl at considerable material and energy costs (IC Potash 2017). These factors result in K2SO4 being sold at a 30% to 60% higher price than that of KCl (CM 2013; IC Potash 2014; EPM 2014; PR 2014). In order to eliminate the intermediate supplier of K2SO4 and capture additional revenue, it could be advantageous for potash producers facilitate this conversion process themselves. Such a process would have to be economically justified by the gain in net margin per ton of KCl converted. Furthermore, this process must be of low capital cost, easily integrated into a mine site’s existing processes, and operate in a manner that does not interfere with the main operation. Existing sources of K2SO4 include limited natural resources and costly synthetic production from KCl. Given Saskatchewan’s considerable low-cost reserves of Na2SO4, a method for producing synthetic K2SO4 from mixing KCl and Na2SO4 in solution was examined. In this study, a two-stage crystallization process is proposed to facilitate the conversion of potassium chloride to potassium sulphate. Also, the techno economic feasibility of the process and some aspects of the design considerations were studied. While independent facilities can complete this conversion, there are significant advantages to integrating this process as an “upgrader” to complement the operation of potash mine refineries. Auxiliary processes to handle waste salts from the conversion are significantly reduced by relying on nanofiltration technology to achieve a separation between sulphate and chloride salts, where the chloride salts are returned for handling by existing refinery systems. With respect to other methods of potassium sulphate production worldwide, Saskatchewan’s potash producers will have a strong cost advantage in the marketplace while improving their existing operations.

The conversion of KCl to K2SO4 occurs as a double-exchange reaction with Na2SO4, written in stoichiometric form as follows:

6 KCl + 4Na2SO4 -> 3K2SO4*Na2SO4 + 6 NaCl    (1)

2 KCl + 3K2SO4*Na2SO4 -> 4K2SO4 + 2 NaCl      (2)

The intermediate product of this reaction is glaserite, a double-salt (K2SO4*Na2SO4). The solubility of this salt is particularly challenging for this process, as it prevents the reaction from occurring as one step. Glaserite must be physically removed from the initial solution, and used as a reactant in pure KCl brine to produce K2SO4. Typically, facilitating this conversion as a stand-alone process follows a common design as show in Figure 1.

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Figure 1. Process flow diagram of stand-alone glaserite process

The combination of KCl and Na2SO4 in solution will result in a reaction producing glaserite. While the exact proportions may vary, sufficient reactants are supplied to oversaturate the brine in glaserite, causing it to crystallize. The crystals are removed from the filtrate and combined with a pure KCl brine to produce K2SO4. This is removed as product, while the filtrate is returned to the glaserite reactor. The caveat with stand-alone processes is the treatment of filtrate from the glaserite reactor. As NaCl builds up within the process, its performance becomes impaired. Due to the solubility of Na2SO4, however, it is necessary to subject the brine to cooling crystallization, removing sodium sulfate dodecahydrate. Care must be made in this step to ensure that KCl does not crystallize as well. The filtrate must then be processed via evaporative crystallization, where the NaCl can be removed. Cooling the filtrate back down again yields crystal KCl, which can then be recycled into the K2SO4reactor. The depleted filtrate is recycled back into the glaserite reactor, or is discharged as waste. The use of this reaction as a conversion method has been well-documented, with a variety of different proposed processes. Gunn (1964), Gunn (1965) and May (1964) developed the initial continuous process using this conversion method. Later a number of researchers explored variations in the process in order to improve feeds retention while removing by-product sodium chloride (Sokolov et al. 1978; Scherzberg et al. 1992; Lampert and Holdengraber 1994; Grmzil and Kic 2005; Susalra et al. 2007). As noted by Lampert and Holdengraber (1994), the caveat of this process seems to be an unfavorable trade-off between using expensive and energy-intensive fractional crystallization systems to separate sodium sulfate, potassium chloride, and sodium chloride, or enduring significant material recycle flows, large retention times, and poor conversion rates. To circumvent the challenges posed by having to remove sodium chloride, more robust variations of the glaserite process were proposed. Fractional crystallization of glaserite via evaporative crystallization (Lampert and Holdengraber 1994) and cooling crystallization (James 2009) were proposed. The use of a graesser differential contractor was explored as a means of shrinking the process (Lampert 1996). Alternatively, a method using cationic exchange was discussed by Derdall (2006), replacing the formation of Glaserite with an ion exchange. However, the process was still limited by the need to recover salts from the filtrate after the final crystallization of potassium sulfate.

While it is clear that an independent plant using the glaserite process has significant challenges, integrating the process into an existing potash mine allows one to forego most of the required auxiliary systems. These mines already have processes to separate potassium chloride and sodium chloride from solution. Conventional underground mines use flotation to separate potash from the ore, while solution mines use a combination of evaporative and vacuum-cooled crystallization to separately remove sodium and potassium chloride. Employing a method to recover only sulfate salts back into the conversion process, the discharged brine containing sodium and potassium chloride salts can be reprocessed by the mine. Further efficiencies can be found when considering sharing power and heat supply, along with tailings management. The separation of sulfate salts from brine solutions, has been accomplished through the use of nanofiltration to high percentages of recovery. In particular, Twardowski (1995) and Twardowski and Ulan (1996) described a nanofilter design capable of rejecting 98.8% of sodium sulfate from concentrated Na-Cl-SO4-H2O brine, and rejecting 96.2% of potassium sulfate from concentrated K-Cl-SO4-H2O brine (where, in the case of our process, rejected brine is recycled back into the process). While there is a need for rigorous testing, and for a more detailed design of the nanofiltration system, it is within reason to assume that a nanofiltration system could be designed to recover nearly all sulfate salts before discharging the filtrate brine back to a potash mine. The specific design of such a process is dependent on the specific mine at which it will be used. For demonstration purposes, a hypothetical installation at a potash solution mine will be chosen, given the presence of crystallization systems dedicated to handling brine (as opposed to flotation processes used at most conventional potash mines). In particular, it is assumed that slurry of potash waste (produced from dryer dust and undersized product) is supplied to the proposed SOP process. The particulars of the process at steady-state are shown in Figure 2. Potash waste slurry is conditioned by a heat exchanger, and solids are dissolved in a tank. The KCl solution is then fed into the second-stage crystallizer, where it is mixed with glaserite to produce K2SO4, which then crystallizes and is separated as a product. The filtrate is returned to the first-stage crystallizer, where it is mixed with brine containing Na2SO4 and recycled K2SO4. The NaCl – produced in the second-stage crystallizer – improves the crystallization of glaserite in the first stage. Filtrate from the first stage is diluted (to prevent scaling) and pass through nanofilters, separating sulfate salts from chloride salts. Na2SO4 is added to saturate the recycled sulfate salt brine, while the chloride salt brine is returned to the mine.

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Figure 2. Process flow diagram of proposed integrated glaserite process

To assess the efficiency of the proposed process, a simulation model was developed and implemented. Due to the limitations with existing commercial process simulators to offer a suitable representation of the aqueous system in question, especially with regards to predicting solubility, an in-house spreadsheet-based model was developed and used to determine the mass and energy balances, phase equilibrium, and physical properties of the process.

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Specifying the temperature, and the concentration of KCl and Na2SO4, these two solubility products can then be set equal to one another and solved for the requisite saturation of K2SO4. The prediction of the solubility of glaserite required a simplification of the K-Na-Cl-SO4-H2O system. The phase diagrams, given in Figure 3, suggest that the solubility of glaserite is not significantly affected by the presence of chloride ions in solution. This permits the use of empirical equations presented by Mazzarotta (1990), which calculated the solubility of glaserite for the K-Na-SO4-H2O system based on the mass ratio of K2SO4 to Na2SO4, specifically at 30°C and 45°C. Calculating the molar ratio of potassium ions to sodium ions for a solution within the K-Na-Cl-SO4-H2O system, and converting this ratio into an equivalent mass ratio of K2SO4 to Na2SO4, Mazzarotta’s equations can be used. A particularly useful feature of Mazzarotta’s equations, is the ability to linearly interpolate between them to account for temperature variation; given the hydration of Na2SO4 at about 32°C (which can lead to process instability), it is in fact most favorable to operate this process between the equation’s temperature limits. This simplification of the K-Na-Cl-SO4-H2O system does require an assumed retention of sulfate ions in solution. This can be determined by the saturated composition at the glaserite-KCl-K2SO4, glaserite-KCl-NaCl, or glaserite-NaCl-Na2SO4 invariant points. In the first two cases, an anionic ratio of 2.5% for sulfate can be assumed. In the third case, an 8.5% ratio can be assumed. Based on the low NaCl content of the process being proposed, a 2.5% ratio is used to help calculate glaserite solubility.  This process does require a further adjustment to account for the increasing solubility of glaserite as it crystallizes. A single calculation of the crystallization of glaserite, followed by an adjusting mole balance, reveals that the calculation has not reached equilibrium. Instead, an iterative calculation is required to repeatedly calculate the mole balance across the first stage crystallizer while ensuring that the crystallizer is truly saturated in glaserite after the crystals are removed.

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Figure 3. Phase diagram of the K-Na-Cl-SO4-H2O system (Scherzberg et al., 1992)

3.2 Simulation development

The simulation was built to reflect the process operation at steady state, assuming perfect mixing in each unit. The density and heat capacity of the brine within each unit was calculated from the brine’s composition and temperature using empirical models (Laliberté and Cooper 2004; Laliberté 2009). In units where different brines are mixed, it was noted that an iterative solution would be required to reconcile the difference between outlet temperature and the outlet brine properties - as these are co-dependent. Instead, the outlet brine properties were calculated based on the perfectly mixed inlet brine, with the temperature based on the amount of energy into the unit. The error between the energy into the process and the energy out of the process (as calculated based on the outlet brine properties and temperature) did not exceed 3% on an absolute basis. The solubility models for glaserite and K2SO4 were used to calculate the production of the first and second stage crystallizers, respectively. Any heat of crystallization accounted for in the crystallizers’ energy balances. Solubility models were also used to maintain undersaturation in all other units, and were used to calculate the required dilution for the nanofiltration system. Simulation of the nanofiltration system involved splitting the flow of the brine based on volume recovery and divalent anion rejection. The percent recovery and percent rejection of the nanofilter was assumed based on published performances for similar brine compositions and operating conditions. To ensure good convergence, calculations for the simulation were controlled using in- house developed Visual Basic user-defined function, with critical processes having dedicated spreadsheets. Dedicated spreadsheets were used in order to take advantage of the ability to control their calculation individually, allowing sensitive equilibrium balances to be completed before iterating through the entire simulation again. To monitor the simulation convergence, readings within the simulation were recorded following iteration. Figure 4 shows the example, convergence of the crystallization processes.  In order to maintain the stability of the simulation while it converged, it was necessary to control the addition of Na2SO4 into the process. This was accomplished by defining a percent saturation of Na2SO4 in the recycle brine, and incrementing or decrementing the actual mole flow of Na2SO4 added in the process. The increment steps were controlled by specifying a gain, which was used to divide the difference between the Na2SO4 required to achieve the required saturation and the molar flow of Na2SO4 calculated for a particular iteration. The simulation achieved convergence when the percent change in the monitored values between the current and previous iterations was within one 0.01%. Under this precision, the simulation converges within 50 to 100 iterations.

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Figure 4. Example of process simulation’s convergence for crystallization

The solubility models were experimentally validated following the proposed conversion process. To produce glaserite, saturated solutions of KCl and Na2SO4 were prepared and mixed together under temperature-controlled conditions. The resulting crystal product was filtered, dried, and stored for conversion into K2SO4. In similar fashion, the produced glaserite was then used to prepare a saturated glaserite solution, which was then mixed with a saturated KCl solution. The final product was removed after crystallization was complete. A variant of this experiment was completed to assess the effects of the presence of NaCl in the KCl solution, with regard to both the feasibility of the process as well as the validity of the solubility models used. In examining the crystal yields of glaserite and K2SO4, it was found that the presence of NaCl had a significant influence on the glaserite yield as given in Tables 1 & 2. However, NaCl at a certain limit during the crystallization of glaserite had no significant effect on crystal yield compared to scenarios of pure KCl brine. However, NaCl was found to have a detrimental effect on the crystallization of K2SO4 from glaserite as shown in Table 2. Results from Table 3 show that, when the pure KCl solution was considered, the proposed models developed to predict the solubility of glaserite and K2SO4, showed strong agreement with experimentally measured results. When considering the effects of mixing NaCl into the KCl brine, it was found that the prediction of glaserite solubility was satisfactory up to 60 grams-per-liter of NaCl. In all other cases, however, an unacceptable level of error occurred. In the case of glaserite at 120 grams-per-liter of NaCl, it is suspected that the filtrate solution resided at the glaserite – NaCl – Na2SO4 invariant point, resulting in an overestimation of glaserite crystallization. It was also found that the prediction of K2SO4 solubility could not be completed if significant NaCl was present in the solution.

Table 1. Experimental yields of glaserite crystals

Sample ID

Volume (L)

Temperature (°C)

KCl (g)

NaCl (g)

Na2SO4 (g)

Glaserite Yield (g)

GPB1

0.5

36°C

175.67

0.00

241.39

98.25

GPB2

0.5

41°C

172.31

0.00

238.95

93.97

GPB3

0.5

41°C

173.64

0.00

240.02

79.66

GPB4

0.5

40°C

175.40

0.00

240.91

98.90

GPB5

0.5

41°C

175.42

0.00

242.28

100.72

GPB6

0.5

41°C

177.19

0.00

242.42

100.46

GB7A

0.5

36°C

170.00

30.00

235.00

107.13

GB7B

0.5

35°C

170.00

30.00

235.00

109.24

GB8A

0.5

35°C

155.00

60.00

235.00

94.70

GB8B

0.5

36°C

155.00

60.00

235.00

99.58

GB9A

0.5

35°C

127.50

120.00

235.00

48.63

GB9B

0.5

35°C

127.50

120.00

235.00

32.87


Table 2. Experimental yields of potassium sulfate crystals

Sample ID

Volume (L)*

Temperature (°C)**

KCl (g)

NaCl (g)

Glaserite (g)

K2SO4 Yield (g)

KPB1

0.82

40

142.02

0.00

73.92

30.94

KPB2

0.86

40

175.23

0.00

86.43

42.41

KPB3

0.86

40

152.67

0.00

79.28

33.03

KPB4

1.1

25

195.28

0.00

98.71

58.57

KPB5

1.1

20

190.83

0.00

99.25

57.84

KPB6

1.1

20

190.00

0.00

99.48

60.11

KB7A

0.65

30

220.00

40.00

90.00

11.71

KB7B

0.65

22

220.00

40.00

90.00

15.78

KB8A

0.65

35

200.00

80.00

90.00

2.27

KB8B

0.65

23

200.00

80.00

90.00

2.89

KB9A

0.2

35

51.00

48.00

30.00

1.61

KB9B

0.2

22

51.00

48.00

30.00

2.75


Table 3. Measured and model predicted yields of glaserite and K2SO4

Sample ID

Glaserite

Measured

(g)

Glaserite Calculated (g)

Error

(%)

K2SO4

Measured

(g)

K2SO4

Calculated (g)

Error

(%)

G/KPB1

98.25

100.48

2.22

30.94

30.12

-2.72

G/KPB2

93.97

97.46

3.58

42.42

44.60

4.89

G/KPB3

79.66

78.32

-1.71

33.03

33.94

2.68

G/KPB4

98.90

98.64

-0.26

58.58

56.48

-3.72

G/KPB5

100.72

100.48

-0.24

57.84

60.52

4.43

G/KPB6

100.46

102.09

1.60

60.11

60.58

0.78

G/KB1A

107.13

105.59

-1.46

11.72

48.94

76.05

G/KB1B

109.24

105.59

-3.46

15.78

48.94

67.76

G/KB2A

94.70

88.29

-7.26

2.27

45.70

95.03

G/KB2B

99.58

88.29

-12.79

2.89

45.70

93.68

G/KB3A

48.63

59.98

18.92

1.61

Nil

N/A

G/KB3B

32.87

59.98

45.20

2.75

Nil

N/A


The process was simulated under the assumption of supplying it with a saturated solution of KCl at 40 oC, containing 100 grams-per-liter of KCl solids. The rate at which the KCl brine and Na2SO4 were supplied to the process was adjusted to produce approximately 40,000 tons per year of K2SO4. This is a reasonable scale for this process, considering the size of the North American market and the logistical challenges that would otherwise arise from pursuing international export. Another critical assumption of the model is the recovery and rejection performance of the nanofiltration system. Simulation of the process assumed a recovery rate of 45% of the solution volume, and a rejection of 97.5% of sulfate ions and 5% of chloride ions. It should be noted that rejected ions are in fact retained by the process, constituting its recycle stream. Using the simulation under these assumptions, for brine feed flow of 6.25 liters-per-second, 40,860 tons-per-year of K2SO4 was produced from 34,961 tons-per-year of KCl and 33,423 tons-per-year of Na2SO4. The utilization of KCl (this being the ratio between actual consumption of KCl and the mass actually fed into the process) was 38%. While the simulation of this process is not perfect, this result indicates that such a process could be technically feasible.

5.1 Capital Cost Estimate

The capital cost estimate was prepared on the basis of the major equipment required as described by the block flow diagram shown on Figure 2. Detailed information from the model simulation was employed to size the individual pieces of equipment appropriately. Instrumentation and contingency costs were also taken into account in the estimate. The target K2SO4 production for the capital cost estimate was considered to be 40,860 tons/year. The total plant capital cost estimate was determined to be approximately $26.7 million, which translates into a capital expense of $654/ton of K2SO4 produced.

5.2 Operating Cost Estimate

An operating cost estimate was completed for an annual basis before income tax expenses by considering major operating expenses at the preliminary study stage. The operating costs considered in the estimate include direct and general expenses, as well as depreciation expenses over the proposed 30-year lifetime of the project. The annual operating cost estimate that includes the annual production cost of KCl was determined as $11.1 million/year. If annual material cost of KCl is excluded from the calculation, the annual operating cost estimate is reduced to $6.7 million/year.

5.3 Profitability Analysis

A profitability analysis was performed in order to evaluate the economics of the project using the net present value (NPV) method. The analysis was completed by considering a capital cost estimate of $26.7 million, a gross annual income of $21.2 million/year based on a production capacity of 40,860 tons/year and a K2SO4 market price of $520/ton (KCl market price of $400/ton and a 30% K2SO4 premium), and a net annual income after taxes of $7.3 million/year ($178/ton) for the case that includes KCl as an annual raw material cost and a net annual income after taxes of $10.4 million/year ($255/ton) for the case that excludes KCl as an annual raw material cost. The simple payback periods were calculated by considering the ratio of the capital cost estimate and the net annual income after taxes for the previously mentioned cases. A payback period of 3.7 years was determined for the case that includes the annual KCl material cost; whereas the payback period for the case that excludes the annual KCl material cost is 2.6 years. Assuming a 10% discount rate to be appropriate for further analysis, a NPV of $42 million and an internal rate of return (IRR) of 27% are realized for the case that includes the annual KCl material cost. On the other hand, when excluding the annual KCl material cost the NPV considerably increases to $71.4 million with an IRR of 39.0%.

An implementation of the glaserite process as an integrated part of an existing potash mine was evaluated. An integrated process can forego the costly auxiliary crystallization systems typically required by a stand-alone glaserite process, along with realizing other capital and operating efficiencies from sharing resources with the larger mine. A custom-built simulation was developed to evaluate this process, and indicate that the process is technically feasible. Specifications for a particular design will require a more robust solubility modeling. It should be noted that the chosen scenario for the simulation is idealized, and that other conditions should be tested once a better simulation has been conceived. Particularly, if the crystallization of glaserite from tailings pond brine or disposal brine can be accomplished, then potash otherwise lost as waste can be used to produce potassium sulfate, significantly improving the economics of this process. Simulation results indicate that the proposed process was technically feasible with a net present value of $42 million, internal rate of return of 27% and payback period of 3.7 years.

This work was supported by the Natural Sciences and Engineering Research Council of Canada (NSERC), the Canada Foundation for Innovation (CFI) and the Clean Energy Technologies Research Institute (CETRi) of the University of Regina.

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Published: 04 April 2017

Reviewed By : Dr. Cristina Carbone.Dr. Afif Khouri.

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Copyright:© 2017 Hussameldin Ibrahim. This is an open-access article distributed under the terms of the Creative Commons Attribution License (CC BY). The use, distribution or reproduction in other forums is permitted, provided the original author(s) or licensor are credited and that the original publication in this journal is cited, in accordance with accepted academic practice. No use, distribution or reproduction is permitted which does not comply with these terms.