REACTION KINETICS AND MASS TRANSFER...

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REACTION KINETICS AND MASS TRANSFER STUDIES IN SELECTIVE LEACHING OF LOW-GRADE

CALCAREOUS PHOSPHATE ROCK

A thesis submitted

To

BAHAUDDIN ZAKARIYA UNIVERSTY, MULTAN

in

Fulfillment of requirements for the degree

of

DOCTOR OF PHILOSOPHY

in

CHEMISTRY

by

MUHAMMAD ASHRAF

DEPARTMENT OF CHEMISTRY

BAHAUDDIN ZAKARIYA UNIVERSITY

MULTAN, PAKISTAN

2010

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DEDICATED TO

My wife

and

Elder brother Ch. Jahan Khan

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DECLARATION I hereby declare that the work described in this thesis was carried out by under the supervision of Prof. Dr. Zafar Iqbal Zafar and Prof. Dr. Tariq Mehmod Ansari at the Department of Chemistry, Bahauddin Zakariya University, Multan, for the degree of Doctor of Philosophy in Chemistry. I also hereby declare that the substance of this thesis has neither been submitted elsewhere nor is being concurrently submitted for any other degree. I further declare that the work embodied in this is the result of my own research and where work of any other investigator has been used, that has been duly acknowledged.

Muhammad Ashraf

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CERTIFICATE

It is certified that the work contained in this thesis has been carried out under our supervision and is approved for submission in fulfillment of the requirement for the degree of Doctor of Philosophy in the subject of chemistry. 1. Prof. Dr. Zafar Iqbal Zafar 2. Prof. Dr. Tariq Mehmod Ansari Department of Chemistry Department of Chemistry Bahauddin Zakariya University, Bahauddin Zakariya University, Multan, Pakistan Multan, Pakistan

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ACKNOWLEDGEMENTS

All praises be to ALMIGHTY ALLAH (SWT), who bestowed man with intelligence,

knowledge and sight to observe and ponder. Peace and blessing of Allah (SWT) be upon the Holy

Prophet MUHAMMAD (SAW), who exhorted his followers to seek knowledge from cradle to

grave.

I feel great pleasure in expressing my sincere gratitude and profound thanks to my research

Supervisor, Prof. Dr. Zafar Iqbal Zafar, Department of Chemistry, Bahauddin Zakariya

University, Multan for his kind guidance and full cooperation throughout the research work. I am

also thankful to my second Supervisor Prof. Dr. Tariq Mahmood Ansari, Department of

Chemistry, Bahauddin Zakariya University, Multan, for his guidance and thought provoking

discussions during these studies. I am also grateful to Prof. Dr. Humayun Pervez, Ex- Chairman

Department of Chemistry, Bahauddin Zakariya University, Multan for providing the research

facilities. I am obliged to express my thanks to Prof. Dr Shahida B. Niazi, Dean of Faculty

Science & Agriculture, who encouraged and facilitated the whole PhD study registration process.

I feel greatly obliged to Prof. Dr. Muhammad Arif (Chairman Department of Chemistry)

and Prof. Dr. Muhammad Aslam Malana, Department of Chemistry, Bahauddin Zakariya

University, Multan for their help and valuable guidance during the research work. I am also

thankful to all the other teachers of Department of Chemistry, Bahauddin Zakariya University,

Multan for their continuous support. I am thankful to the Department of Chemistry, Bahauddin

Zakariya University, Multan for providing me the research facilities, Pak-Arab Fertilizer and

NFC Institute of Chemical Engineering and Technological Centre, Khanewal Road, Multan, for

help in sample crushing, grinding and sieving.

I would like to express my gratitude for Prof. Dr. Muhammad Shakirullah Khan, Institute of

Chemical Sciences, University of Peshawar, for his cooperation in providing the facilities for

instrumental analysis, P. C. S. I. R. laboratories, Lahore is gratefully acknowledged for

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providing the facilities of thermal analysis. I am thankful to Kohistan Engineering Ltd,

Abbot Abad, for their help in collection of phosphate ore samples. I am also thankful to

the Hazara Phosphate Fertilizers for their help to complete this research work.

I am highly obliged to my elder brother (Gen) Dr. Ahmad Khan and his wife Razia

Ahmed for their moral support and encouragement during the higher education. I feel

myself lucky enough to find very sincere friends, like Sajjad Hussain, Zahoor Ahmed,

Muhammad Saleem, Muhammad Shafi, Nazar, Ajmal and Javeed, who were eager to

listen about my success.

Many thanks to the staff of the laboratories of Inorganic, Organic, Physical and

Analytical Chemistry, Shafi Dogar, Ajmal, Riaz, Ishfaq, Yousaf, Ashraf, Akram, Riaz

Mettela and Javeed, who arranged all the requirements whenever I needed.

I also acknowledge the prayers and well wishes of my elders, Asghar Ali Jalli, Rana

Sajjad Hussain, who encouraged me all the time during the research work. I am very

grateful to my college Principal and all the Colleagues whose moral support encouraged

me to complete the research work. I also gratefully acknowledge the pleasant feelings of

my family, brothers, sisters and my children, Ali, Fizza and Ghana, who also felt my

absence during the research work, as I could not spare time for them. Concluding this

work, I feel pleasure that I may spend much time with them in future.

Muhammad Ashraf

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LIST OF CONTENTS Page No

DEDICATION 1

DECLARATION 2

CERTIFICATE 3

ACKNOWLEDGEMENT 4

CONTANTS 6

LIST OF TABLES 9

LIST OF FIGURES 11

ABSTRACT 15

CHAPTER NO 1 17

1. INTRODUCTION 17

1.1 Flotation Method 19

1.1.1 Siliceous Ores 20

1.1.2 Dolomite Ores 22

1.1.3 Highly Selective Collectors for Apatite in Igneous Ores 24

1.1.4 Depressants for Apatite in Sedimentary Ores 25

1.1.5 Selective Collectors for Carbonates in Sedimentary Ores 26

1.2 Other Flotation Separation Techniques 27

1.3 Magnetic separation Techniques 31

1.4 Calcination Method 32

1.5 Electrostatic Upgrading 35

1.6 Bioleaching of phosphate ores 36

1.7 Partial Acidulation of phosphate ores 37

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1.7.1 Run-of-Pile (ROP) Process for PAPR 38

1.7.2 Single-Step Acidulation/Granulation (SSAG) Process 39

1.8 Direct Application of Phosphate Rock 40

1.9 Dissolution and Selective Leaching 41

CHAPTER 2 47

2. Chemistry of the Leaching Process 47

CHAPTER 3 52

3. Material and Methods 52

3.1 Sample Collection 53

3.2 Size reduction and sieving 59

3.3 Sample Preparation for Analysis 63

3.4 Selective Leaching with Lactic Acid 69

3.5 Experimental Procedure 75

3.6 CHAMICALS USED 77

3.7 Instrumental Techniques Used for Analysis 78

3.7.1 Atomic Absorption Spectrophotometer 78

3.7.2 Absorption Mode 78

3.7.3 Emission Mode 78

3.7.4 Chemical Interference 80

3.7.5 Calibration of Atomic Absorption Spectrophotometer 81

3.8 Scanning Electron Microscopy 83

3.8.1 Energy Dispersive X-ray Analysis 83

CHAPTER 4 86

4. Results and Discussion 86

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4.1 Effect of Reaction Time on the Leaching Process at Different 86

Temperatures

4.2 Effect of Reaction Temperature on Dissolution 93

4.3 Effect of Acid Concentration on P2O5 and CO2 Contents 93

4.4 Effect of Liquid/Solid Ratio on P2O5 and CO2 Contents 94

4.5 Leaching with Lactic acid 108

4.5.1 Effect of Acid Concentration on the Conversion 108

4.5.2 Effect of Liquid-Solid Ratio on the Conversion 109

4.5.3 Effect of Particle Size on the Conversion 110

4.5.4 Effect of Temperature on the Conversion 110

4.5.5 Effect of Reaction Time on P2O5 and CO2 Contents 111

4.6 Kinetic Analysis 139

4.6.1 Leaching with Succinic Acid 139

4.6.2 Leaching with Lactic Acid 151

CHAPTER 5 163

5.1 Recovery of Spent Acid and Economy of Process 163

5.2 Routes for the Acid Recovery 163

(a) Use of Ion Exchange Resins 142

(b) Use of Hydrofluoric Acid 163

(c) Use of Fluorosilicic Acid 164

(d) Use of Sulphuric acid 164

5.3 Economy of the Process 167

Conclusions 169

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References 172

APPENDIX A 189

Appendix B 189

PUBLISHED WORK 191

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LIST OF TABLES Page No.

Table 3.2.1 Mesh size, average size, weight fraction and Weight percent 59

Table 3.2.2 Average size, mass fractions and cumulative mass fraction 61

Table 3.3.1 Analysis of the rock samples 64

Table 3.4.1 Analysis of the rock samples 70

Table 4.1.1 Effect of reaction time on leaching of calcareous 87

(XB %) material at various temperatures

Table 4.2.1 Effect of temperature on selective leaching 96

(XB %) of calcareous material at various acid concentrations

Table 4.3.1 Effect of acid concentration on P2O5 and CO2 Contents. 102

Table 4.4.1 Effect of liquid/solid ratio on P2O5 and CO2 105

contents using 8% acid concentration for 55 min at 55oC

Table 4.5.1 Effect of reaction time on the conversion 112

(XB %) of calcareous material at different acid concentrations

Table 4.5.2.1 Effect of Reaction Time on the Conversion of 118

calcareous material in the rock at Various Liquid/Solid Ratio

Table 4.5.3.1 Effect of reaction time on the conversion 124

using different size particles of the rock 103

Table 4.5.4 1 Effect of reaction time on the conversion 130

(XB) of calcareous material in the rock at various temperatures

Table 4.5.5.1 Effect of reaction time on P2O5 and CO2 contents 136

Table 4.6.1.1 The values of ln k and 1/T 147

Table 4.6.2.1 Effect of rpm on the conversion 153

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Table 4.6.2.2 The values of the selected parameters and the reaction 160

periods calculated from Eq. (4.6.2.4) for these parameters

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LIST OF FIGURES Page No

Fig 3.1.1 Phosphate rock deposit of Shahzad mines 53

Fig 3.1.2 Phosphate rock at Largrarian mines 54

Fig 3.1.3 Phosphate rock deposit mining at Shahzad mines 55

Fig 3.1.4 Phosphate rock deposit of Largrarian mines 56

Fig 3.1.5 Phosphate rock mining of Puttiann mines 57

Fig 3.1.6 Phosphate rock mining at Puttiann mines 58

Fig. 3.2.1 Average particle size and wt% of phosphate rock sample 60

Fig. 3.2.2 Average particle size and cumulative mass fraction of phosphate rock 62

Fig. 3.3.1 Effect of average particle size on P2O5 content 65

Fig. 3.3.2 Effect of average size of particle on AIR content 66

Fig. 3.3.3 Effect of average particle size on LOI content 67

Fig. 3.3.4 Effect of average particle size on CO2 content 68

Fig. 3.4.1 Effect of average particle size on P2O5 content 71

Fig. 3.4.2 Effect of average particle size on CO2 content 72

Fig. 3.4.3 Effect of average particle size on LOI content 73

Fig. 3.4.4 Effect of average particle size on AIR content 74

Fig.3.7.1 AAS Hitachi (A-1800) used for analysis 79

Fig.3.7.2 Correlation curve for calcium nitrate versus absorbance 82

Fig.3.8.1 SEM micrograph of the sample 84

Fig. 3.8.2 EDX spectrum of the sample 85

Fig. 4.1.1 Effect of reaction time on the conversion 88

(XB%) of calcareous material at 40oC

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Fig.4.1.2 Effect of reaction time on the conversion 89








Fig. 4.2.1 Effect of temperature on dissolution (XB) with 6% succinic acid 97

concentration and liquid/solid ratio 5:1 for 30 minutes



Fig. 4.2.3 Effect of temperature on dissolution (XB) with 8% 99

succinic acid concentration and liquid/solid ratio 5:1 for 30 minutes



Fig. 4.2.5 Effect of temperature on dissolution (XB) with 10% succinic acid

concentration and liquid/solid ratio 5:1 for 30 minutes 101

Fig. 4.3.1 Effect of acid concentration on P2O5 content with liquid/solid 103

ratio 7:1 for 55 minutes at 55oC

Fig. 4.3.2 Effect of acid concentration on CO2 content with L/S ratio 7:1 for 104

55 minutes at 55oC

Fig.4.4.1 Effect of liquid/solid ratio on P2O5 contents using acid 106

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concentration 8% for 55 minutes at 55oC

Fig.4.4.2 Effect of liquid/solid ratio on CO2 contents using acid 107

concentration 8% for 55 minutes at 55oC

Fig. 4.5.1 Effect of reaction time on the conversion of calcareous material 113

with 5% lactic acid concentration









Fig.4.5.2.1 Effect of reaction time on conversion with liquid/solid ratio (L/S) 4:1 119

Fig. 4.5.2.2 Effect of reaction time on conversion with liquid/solid ratio (L/S) 5:1 120




Fig. 4.5.3.1 Effect of reaction time on conversion for particle size 0.503 mm 125

Fig.4.5.3.2 Effect of reaction time on conversion for particle size 0.356 mm 126

Fig. 4.5.3.3 Effect of reaction time on conversion for particle size 0.2515 mm. 127



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Fig. 4.5.4.1 Effect of reaction time on the conversion at 25oC 131

Fig. 4.5.4.2 Effect of reaction time on conversion at 35oC 132




Fig. 4.5.5.1 Effect of reaction time on P2O5 contents 137

Fig. 4.5.5.2 Effect of reaction time on CO2 contents 138

Fig. 4.6.1.1 Time versus 1-(1-XB)1/3 at 40 oC 141

Fig. 4.6.1.2 Time versus 1-(1-XB)1/3 at 50 oC 142

Fig. 4.6.1.3 Time versus 1-(1-XB)1/3 at 60oC 143

Fig. 4.6.1.4 Time versus 1-(1-XB)1/3 at 70oC. 144

Fig. 4.6.1.5 Time versus 1-(1-XB)1/3 at 80oC 145

Fig. 4.6.1.6 Arrhenius plot of ln k vs 1/T 148

Fig. 4.6.1.7 Agreement between experimental and calculated conversion values 150

Fig. 4.6.2.1 Effect of rpm on the conversion at specific reaction conditions 154

Fig. 4.6.2.2 Agreement between experimental and calculated conversion values 157

Fig. 5.2.1 Schematic diagram suggested for the recovery of spent acid 166

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ABSTRACT

In the present thesis, the dilute solutions of succinic acid have been used to

investigate the selective leaching of calcareous material in low-grade phosphate rock. The

effective parameters on the dissolution rate are reaction temperature, particle size, acid

concentration and liquid-solid ratio. The results indicate that succinic acid can be used to

selectively dissolve the calcareous material in low-grade phosphate rock as it improves

the P2O5 content of the rock and makes it viable as a feed to an acidulation plant.

Using the known size particles of the sample, acid concentration and liquid-solid

ratio, the influence of different reaction temperatures has been studied in order to

elucidate the leaching kinetics of calcareous material in the rock. The results show that

the leaching rate of calcareous material increases with increasing temperature. A kinetic

model has been suggested to describe the selective leaching process of calcareous

material analyzing the kinetic data. The selective leaching curves have been evaluated in

order to check the validity of shrinking core models for liquid-solid systems. The

experimental data have been tested by graphical and statistical methods and it is found

that the leaching of calcareous material in the rock is controlled by the chemical reaction

i.e. tex RT/92.6463/1 1047.1)1(1 . The apparent activation energy of the dissolution

process has been found to be 64.92 kJ mol-1 over the reaction temperature range from 313

to 353 K. Such a value of activation energy indicates that the process is a chemically

controlled reaction and agrees with the values obtained in the similar research of fluid-

solid reaction systems. The agreement between the experimental conversion and the

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values calculated from the suggested empirical equation has been tested, and which is

found to be very good.

For a comparison study, lactic acid has also been used to study the selective

leaching kinetics of calcareous phosphate rock. The effect of acid concentration,

liquid/solid ratio, particle size and temperature has been studied in order to expound the

leaching kinetics of calcareous material in the rock. It has been found that the leaching

rate of calcareous material increases with increasing the acid concentration, liquid/solid

ratio and temperature and decreasing particle size. A semi-empirical model has been

suggested to illustrate the selective dissolution of calcareous material analyzing the

experimental data. The selective leaching curves have been evaluated to test the validity

of kinetic models for liquid-solid systems. The kinetic data are analyzed by graphical and

statistical methods and it has been found that the dissolution of calcareous material in the

rock is controlled by chemical reaction i.e.

teDSLC RT/62.42954737.0627.1753.13/1 )/(1.19)1(1 .

The analysis of the obtained results reveals that the applicability of the suggested

model is good and it can work within a certain range for the choice of adjustable

parameter values depending on the degree of selective leaching. The results show that the

controlling step of the overall process of the heterogeneous reaction is a chemical change.

According to the analyzed results, it is recommended that the parameter values for the

optimum selective leaching rate are C = 8% v/v, L/S = 7 cm3g-1, T = 318 K, SS = 350

min-1 and D = 0.1255 mm.

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CHAPTER NO 1

1. INTRODUCTION

Naturally occurring deposits of tricalcium phosphate which are generally

combined with Calcium fluoride CaCO3 (PO4)2.CaF2 are called phosphate rocks.

Compositions of these rocks varies from deposit to deposit, and are the most important

source for phosphate fertilizers and phosphorous-based chemicals, which are used to supply

food and feed for mankind and animals. There is no substitute of these phosphate rocks as a

raw material for the production of phosphate fertilizers in the world at this time. As the world

population continues to increase, so does the demand for phosphate day by day in the world

due to the consumption of high-grade phosphate rock deposits. The future demand of

phosphate for food and feed grade may be expected to be fulfilling from the low-grade

phosphate rock deposits as an alternative source for the production of phosphate fertilizers

and various phosphorus based chemicals in the world. On the other hand, a number of

deposits of the world need commercial exploitation regarding improvements in mining as

well as beneficiation technology.

In the past, some researches [1-12] have reviewed the status of phosphate rocks

and their beneficiation techniques and various methods. The total phosphate content, or

grade, of phosphate rock is commonly reported as phosphorous pentoxide, P2O5. It also may

be expressed as calcium phosphate, Ca3 (PO

4)

2, which has been referred to traditionally as

'bone phosphate of lime' (BPL), reminiscent of earlier times when animal bones were

consumed widely in fertilizer manufacture (%P2O

5 × 2.1853 = %BPL; %BPL × 0.4576 =

%P2O

5). Marketable phosphate rock, usually beneficiated or as a concentrate, ranges from

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about 60 to 80% BPL (i.e. between 28 and 39% P2O

5); material that enters international trade

generally contains more than 65% BPL or 30% P2O

5 [8].

Prospects for supply and trade cannot reasonably be addressed without some

estimation of demand: the three go hand in hand in any commodity business, though their

relative importance may change. The demand for phosphate rock and prospects for industry

cannot be viewed in isolation from the prospects of the phosphate fertilizer sector as a whole.

Of the world's annual P2O5consumption of 33 000,000 t about 88% for fertilizers, 4% of

animal-feed grade and about 8% was used by manufacturing industry [7]. World

consumption of P2O5 for non-fertilizer uses was about 2 500, 000 t /year and, in addition,

about 500, 000 t/year of P2O5 equivalent was consumed as elemental phosphorus [9].

The bulk of the world's phosphate deposits that are located in sedimentary

horizons also contain considerable amount of carbonate impurities associated with apatite

matrix. Since these are also the most difficult type of deposit to treat, initial exploitation was

restrained to the limited number of deposits in which carbonate was so friable as to be largely

removable in a fine fraction. The increasing demand for phosphate and depletion of the more

amenable reserves, however, have steadily increased the pressure for treatment of these

difficult sedimentary phosphate-carbonate deposits and probably no other mineral processing

topic has attracted as much research in the past [6]. Most of the phosphate rocks are of

sedimentary origin with low-grade phosphate elements along with higher levels of various

impurities. The first challenge in phosphate industrial sector is the reduction of impurities

from low-grade phosphate rocks so that these rocks may be used for the production of

phosphate fertilizers to meet the increasing demand for phosphate and to cover the depletion

of more amenable reserves in the world [12].

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Due to complex raw materials along with varying composition from source to

source, these phosphate rocks from different deposits may be expected to behave differently

in acidulation processes, which are the heart of all phosphate fertilizers industries. Phosphate

rocks are primarily composed of the apatite group in association with a wide assortment of

minerals, mainly fluorides, carbonates, clays, quartz, silicates and metal oxides. These

deposits fall naturally into two major classes (i) igneous origin; (ii) sedimentary origin. In

both of these major natural types of deposits a very important subdivision can be recognized

on the basis of whether the gangue is generally siliceous in character or the ore contains

considerable amount of such carbonate minerals as calcite and dolomite [6].

Due to the adverse effect of impurities [13, 14] in phosphoric acid production,

these impurities must be removed or reduced to the lowest possible levels. The beneficiation

method, which must be employed to achieve an efficient separation, may depend on the type

and nature of impurities present in phosphate rocks. For beneficiation of low-grade

phosphate rocks, a number of techniques may be described as follows:

1.1 Flotation Method

Separation technique for silica based and calcareous gangue material from low

grade phosphate rock by using different types of floating reagents (amine and fatty acids) is

called flotation. For over 50 years flotation has been the most commonly used method for

upgrading phosphate ores both of igneous and sedimentary origin [15-28]. In the former, the

phosphate occurs as high-grade crystalline apatite, and concentrate specifications frequently

call for grades as high as 37-38% P2O5. In sedimentary deposits, however, the phosphate may

be present in a wide variety of forms, ranging from unconsolidated bone fragments to quite

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well crystallized secondary phosphate minerals. The actual phosphate content of these phases

is variable but generally lower than that of igneous apatite, and concentrate specifications are

similarly lower, 28-30% P2O5 being quite well recognized as a lower limit [6, 15].

1.1.1 Siliceous Ores

A large sedimentary deposit with high silica content was mined in Florida at the

rate of one third of the annual world phosphate production. In the beneficiation of Florida

phosphate rock, flotation plays a predominant role because it is the most economical way to

separate sand and other impurities contained in the matrix. Typically, the matrix is washed

and deslimed at 150 meshes. Material finer than 150 meshes is pumped to clay settling

ponds. An estimate of 20-30% of the phosphate (contained in the matrix) is lost in this way.

The rock coarser than 150 mesh can be screened into separate sizes (-31/4+14 mesh) which

may be high or low in phosphate content. Washed rock (-14 mesh+150 mesh) is sized into a

fine usually 35150 mesh, which are treated in separate circuits. A two-stage flotation

process is generally used in these circuits, where the feed is subjected to ‘rougher’ flotation

(fatty acids and fuel oils are used as collectors) to separate the phosphates from most of the

sand [15]. The rougher concentrate is scrubbed by sulphuric acid to remove the fatty acids

and oils. Scrubbed material is then subjected to ‘cleaner’ flotation (amine collectors are used)

to float the sand. This stage of flotation is sensitive to impurities in water, so fresh water is

used instead of recycled water; the fatty acid circuit uses recycled water from the clay ponds.

Flotation of phosphates from the fine feed (35 150 mesh) presents very few

difficulties and recoveries in excess of 90% are commonly achieved using conventional

mechanical cells. On the other hand, recovery of the coarse phosphate feed is more difficult

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and flotation by itself normally yields recoveries of just 60% or even less. In the past,

hammer mills were used for size reduction of the coarse feed. However, due to high

maintenance costs and loss of phosphates as fines (generated during milling) their use has

been discontinued. The industry, however, has taken other approaches to avoid the problem

of low floatability of coarse particles. For example, the use of gravitational devices such as

spirals, tables, launders sluices and belt conveyors modified to perform a "skin flotation" of

the reagentized pulp. Although a variable degree of success is obtained with these methods,

they have to be normally supplemented by scavenger flotation. In addition, some of them

require excessive maintenance; have low capacity or high operating costs. So performance is

less than satisfactory, and in certain cases use has been discontinued.

The exact reasons that explain the low recoveries of coarse particles in

conventional flotation machines are not clear. There are various factors affecting the

behavior of such particles. For instance, the low floatability of large particles could be related

to the extra weight per unit surface area that has to be lifted to the surface (usually under

highly turbulent conditions) and then transferred and maintained in the froth layer. Factors

such as density of the solid, turbulence, stability and height of the froth layer, tenacity of the

particle-bubble attachment, depth of the water column, viscosity of the froth layer, and other

variables that can indirectly influence these factors, determine the floatability of coarse

particles.

Another important disadvantage of the existing beneficiation technology is its

inability to separate impurity minerals containing MgO. Florida phosphate ores contain low

levels of magnesium impurities. However, it is expected that the current deposits of lower

MgO content will be depleted in the next few years. Unfortunately, the future phosphate

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reserves, located south of the present mining district, have a severe limitation in the form of

high MgO content and would require innovations in the beneficiation technology. The

difference in the consumption of reagents can be attributed to the fine size of the flotation

feed since desliming is done with 30-40 micron particles. Nevertheless, the flotation

separation of this ore is much simple (one stage-anionic flotation) because the high

phosphate content (30-31% P2O5) of Senegalese [15].

1.1.2 Dolomite Ores

Commercial phosphate rocks should not contain more than 8% carbonates (about

3.5% CO2) in order to be economical [16, 17]. The presence of free carbonates in the

phosphate rocks usually requires additional acidulent (sulphuric acid) during the manufacture

of phosphoric acid and super phosphates by the wet process. The carbon dioxide produced

during acidulation, causes more foaming and results in production of smaller size gypsum

crystals [16] that may blind the downstream phosphogypsum filters and a low quality

phosphoric acid is produced [16]. The carbon dioxide also contaminates the equipment and

pollutes the atmosphere. An antifoaming can be added, but if the CO2 content exceeds a

certain limit, it will have negative effects on the economy of the whole process [14].

In igneous ores, carbonate frequently occurs as calcite rather than as dolomite. In

some of these ores, it was found that apatite is somewhat more floatable than calcite [15].

Calcite and dolomite, both of these carbonate minerals tend to respond in a similar manner to

phosphate minerals when exposed to the fatty acid-type collectors that are traditionally used

for phosphate flotation. The consequences of the dilution of the phosphate concentrate by

them are, however, some what different: calcite will consume part of the sulfuric acid used

25

for super phosphate manufacture, but dolomite is more objectionable since magnesium in

quantity renders the phosphoric acid unacceptably viscous [6]. The froth flotation to remove

phosphate from the middling [18,19,20] has also been studied, but the phosphate values

could not be economically recovered from the fine fractions of the ore slurry.

The beneficiation of phosphate ores containing dolomitic impurities is

complicated because of the similarities in the chemical behavior of minerals present [21-24].

Much research has been carried out to try to reduce the calcium carbonate content of low-

grade phosphate rocks by flotation [21-23]. Hignett et al. [25] claimed that flotation could

work best on ores containing well-crystallized carbonates. When the ore contained soft and

chalky carbonates the flotation results were less satisfactory. This was the case for East

Mediterranean and North Africa ores where the carbonates crystals are intergrown in such a

way that the phosphate is not liberated by crushing until the size of the material is too fine for

flotation [26,27].

Research efforts to find details of the mechanism of interaction between ions in

the liquid phase and the solid surfaces involved are difficult due to the presence of impurities

[22]. A small level of some impurities in the ore has a marked effect on the surface properties

of calcite and apatite as well as the ionic reactions in the liquid phase. It appears like that a

CaO: P2O5 ratio in the flotation product must be less than 1.6 [28].

A number of depressants have been investigated [29-32] in order to increase the

selectivity of separation using the traditional collectors. Examples of these depressants

include: pearl starch (Brazil), HF, H2SiF6, sulfonates, tannin-based depressant "Totanim"

(South Africa), and starch-ethoxylated carbonic acid (USSR). All of these depressants are

used with the traditional fatty-acid collectors together with added hydrocarbons such as oily

26

gas-tars or kerosene. It is worth mentioning that, to obtain high flotation efficiency, favorable

froth modifications are produced by the use of such reagents as nonyl phenyl polyglycol

ether and ethoxylated alkyl phenol [1].

It is possible that gangue minerals, such as quartz and pyrite, can easily be

separated by flotation. However, the separation of calcite from collophane is a major problem

because; both of these minerals have similar flotation characteristics [30]. Flotation of

carbonate rich ores shows some difficulties because the fatty acids and their derivatives used

as flotation collectors have very poor selectivity for carbonates and phosphate minerals [24].

1.1.3 Highly Selective Collectors for Apatite in Igneous Ores

The emphasis in flotation research has changed from the use of depressants to the

use of more selective collectors, replacing the fatty acid-based reagents. For example,

amphoteric sarcosine has been used successfully in Finland [31] to float apatite from calcite.

Another collector, amino-carboxylic acid reagent (Lilaflot OS-100), has been developed in

Sweden and resulted in excellent separation of apatite from calcite, both in Sweden and

Russian ores. Warren Springs Laboratories (UK) have developed a new reagent of the

immino bis methylene phosphoric acid class, which has given good selectivity in floating

apatite from dolomite of igneous ores [32].

27

1.1.4 Depressants for Apatite in Sedimentary Ores

There is a natural selectivity margin favoring apatite in igneous phosphate-

carbonate ores, however, this is reversed and tends to operate in favour of carbonate

minerals. Consequently, when fatty acid flotation separation of sedimentary ores was

adopted, the search focused not on carbonate, but on possible phosphate depressant systems.

Examples include the use of orthophosphoric acid in the Karatou process [33, 34].

The published literature indicates that a pH of 6.0 is achieved by using 6.0 kg

H3PO4/ton of ore. A synthetic fatty acid (C10-C16) is added at a dose of 0.3 kg/ton. After

carbonate flotation, the pH is raised to 7.6-8 using 1.5 kg/ton of sodium hydroxide and tall oil

(1.7 kg/ton) is added to float phosphates, with sodium silicate at 0.5 kg/ton being used to

depress the silica. The reported data [33,34] indicate that only 75% of the phosphate can be

recovered at a grade of 28% P2O5 from a feed of 22.5% P2O5. In addition, the concentrate

contains 1.3% MgO which is higher than the requirements of a phosphoric acid plant (less

than 1.0%). Moreover, the cost of reagents is excessive to make the process economically

attractive.

Aluminum sulphate and Rochelle salt (sodium potassium tartrate)

(E.N.S.C.Nancy- O.C.P. Morocco Process) were used in this process; the selectivity of

dolomite separation was reported to be very good with over 80% phosphate recovery [29,

35]. However, the presence of clays was found to reduce the efficiency of separation due to

imperfect desliming. Thus, good desliming is a prerequisite to achieve the successful results.

The use of silicofluoride either as sodium salt or hydrofluosilic acid was also reported to be

effective as a depressant for phosphate minerals in their separation from the carbonates

gangue [36, 37]. Addition of alkyl phosphoric acids alone or in combination with HF was

28

found to be effective for apatite depressants as reported by TVA researchers [38, 39].

However, the process was found to be ineffective in processing undeslimed feeds or high

MgO ores, because high reagent doses were needed to achieve separation which rendered the

process uneconomic.

Water soluble salts of tripolyphosphates and hexametaphosphates were also used

by IMC's (International Minerals Company, U.S.A.) researchers to depress phosphate

minerals [40,41]. Dolomite was floated with the sodium salt of sulfonated oleic acid in the

pH range of 5.5 to 6.0. Even though this process has been tested on a pilot plant scale, IMC

does not prefer its commercial application rather than building a heavy media separation

plant. The reasons for not using the above-mentioned process are not known. Nevertheless,

one may speculate that the cost of reagents as well as the increase of phosphate concentration

in recycled water may depress apatite flotation in other circuits dealing with low MgO feeds.

Also, the presence of high levels of sodium ions in process water stored in clay settling ponds

could be an objectionable since this can lead to clay swelling.

1.1.5 Selective Collectors for Carbonates in Sedimentary Ores

Using novel collectors for igneous apatite-carbonate ores, much research work

has been carried out regarding the possible substitution of the traditional fatty acids. For

example, the reagents developed by French researchers [42,43] involving

alkalylaminopropionic and propionic acids showed good separation of MgO from Tunisian

ore. These are amphoteric reagents that are used to float dolomite first at pH 7 then at pH 5.0.

Even, though the types and rates of these reagents consumption are not reported. These types

29

of reagents are quite expensive. It would be worth testing their efficiency in separating

carbonate gangue from the other carbonatic deposits in the world.

In another process, the same workers suggested the bulk flotation of phosphate

and carbonate away from silica by using a phosphoric ester-fuel oil mixture under

moderately basic conditions. H2SO4 then depressed phosphate subsequently at pH 6.0 and

carbonate was refloated with a small addition of collector.

1.2 Other Flotation Separation Techniques

In the IMC process [42,43] in which dense media removes over size phosphate

pebbles by Dynawhrilpool to reject a dolomite-rich lighter fraction is followed by the

grinding of the sink pebbles and treatment by using flotation in conjunction with the original

fines. The classic two-stage flotation system was modified by addition of a third stage to float

phosphate by an amine-kerosene mixture at slightly acidic pH. The U.S. Bureau of Mines

process consists of scrubbing and desliming at 150-mesh to remove MgO-rich slimes. Fatty

acid-fuel oil was used to float apatite and dolomite at pH 9.5 followed by two or three cleaner

stages using sodium silicate as a dolomite depressant. The process did not show promising

results for removing dolomite to less than 1.0% MgO in the concentrates in all situations.

Some other researchers [44,45] studied two and three stages flotation techniques.

Moudgil et al. [44] carried out a process for separation of dolomite from south Florida

phosphate by first conditioning at pH 10.0 with a fatty acid, followed by reconditioning at a

pH around 4.0 using either HNO3 or H

2SO

4 as pH modifiers. The reconditioning pH and

collector concentrations were claimed to be important parameters for flotation selectivity.

30

The currently used commercial process of double flotation [48] is not enough

sufficient for reducing the dolomitic impurity level to less than 1.0 weight percent MgO in

the concentrate, as required by the phosphate industrial sector. Moudgil and Chanchani

[44,46] and Ince [47] carried out some fundamental studies in the development of two

processes to remove the dolomite impurities from apatite. Extension of these fundamental

studies to beneficiate the natural ores on a bench scale was also investigated [49]. However, a

systematic optimization of the major parameters has not been attempted and no guidelines

are available for processing such complex ores. Regarding three stage flotation scheme,

Zhong et al. [45] studied an optimization process for some variables using various chemicals

and flotation reagents.

At the University of Florida, some researchers [47,50,51] have investigated that

sodium chloride acts as an apatite depressant during flotation at acidic pH values with oleic

acid, while dolomite flotation is not affected. However, it was claimed that dolomite particles

larger than 48 meshes were difficult to float and they tended to remain in the sink fraction.

The trends observed during Hallimond cell tests of single and mixed minerals, were

confirmed with mixtures of minerals and a natural ore at bench scale level. At the bench

scale studies, the MgO content of the mineral mixture was reduced from 5 to 1% along with

90% recovery of P2O5 in the sink fraction. However, further testing on pilot scale may be

required to answer some important questions regarding the practicality of the process (e.g.

pH, reagent addition control within specified narrow ranges and the cost of the reagents). The

residual concentration of NaCl in the process water is also one of the considerable

parameters, which can cause environmental problems.

31

Another selective flotation technique for processing dolomitic and calcareous-

siliceous phosphate ores, without the use of a depressant, was developed by Hanna et al,

[52,53]. The process used standard flotation equipment and conventional fatty acid collectors

for carbonate/phosphate separation. It involved a unique procedure of applying conventional

fatty acids as collectors for selectively floating the carbonate minerals under slightly acidic

pH of 4-6. Reagent requirements for the process were modest, being of the order of about

1.5-1.75 kg/t (3.3-3.8 lb/t) oleic acid, 0.66 kg/t (1.3 lb/t) H2SO

4 and 0.5 kg/t (1.1 lb/t) sodium

silicate, for both carbonate and phosphate steps.

Another new method of flotation [32] involved the treatment of the ore

suspension with copper and zinc ions followed by addition of an alkali sulphide solution to

cover the surface of the apatite particles with copper sulphide and zinc sulphide which in turn

are floated by using xanthate type collectors. The economic feasibility of a flotation process

is highly depends on the amount and price of flotation reagents, flotation equipment and the

amount of water needed in the process. For grinding phosphate rock, the energy used was

about 25-35 MJ/ton and the process water needed was around 13 m3/ton while the average

values for ore flotation units were 62.5 MJ per tone and 5.3 m3 per tone for energy and used

water, respectively [54]. For flotation of high carbonate containing sedimentary phosphate

rocks, it has been a challenging need to find the reliable and economical methods. Due to the

changes in impurities, it appears like that each rock has been treated as a special case and

until a reliable and general method is devised accumulation of data and experience are

required.

Another significant development in phosphate flotation was the application of the

flotaire cell [55], which could process rock with larger particle sizes reducing the cost of

32

grinding. It also showed a more favorable grade-recovery relationship. Column flotation has

also received an increased interest [56-61], especially in cleaning flotation, so application to

phosphate flotation should be studied.

33

1.3 Magnetic Separation Techniques

The test work carried out by the USSR researchers indicated that the use of two-

stage dense medium cycloning operation could result in the production of a fine phosphate

sink product of an acceptable grade. It gave an intermediate-density product, which could be

combined with fines from dense medium separation requiring grinding and carbonate

flotation only.

Low-intensity magnetic separation technique was carried out to treat the igneous

apatite deposits (USSR), which contained higher amounts of ferromagnetic minerals, such as

magnetite or titanomagnetite. Latter on, such methods were also extended to the use of high-

intensity magnetic separation to affect the rejection of complex iron bearing silicate minerals

from various new development areas. Magnetic treatment of the +100 mesh fraction resulted

in non-magnetic apatite concentrates (less than 37% P2O5) fines (-100 mesh), which were

then treated by flotation method.

High-intensity magnetic separation results were found to be promising in

upgrading Egyptian ores containing dolomite, silica and significant amounts of pyrite,

organic and water-soluble alkali. After careful dry size reduction, the ore was wetted and

classified to remove fines reducing the alkali content. A low-temperature roast at 650oC

removed organics and converted the pyrites into magnetically separable oxides. Advantage

was then taken of the high sand temperatures to eliminate silica and carbonates in high-

intensity separations at 200 and 800oC. During cooling, the phosphate particles regained

moisture films and reached the surface much faster than the carbonates, thus allowing the

second separation. Pyrite was eliminated by induced-roll magnetic separation process.

34

The selective magnetic coating methods [62-66] were also used to treat the

carbonate-phosphate rocks. The principle (developed at Warren Spring Laboratory, U.K), of

the process was the selective adsorption of fine-grained magnetic material on to mineral

surfaces in a mixed pulp, which rendered the coated grains amenable to recover by

conventional high-intensity magnetic separation. The adsorption of fine particles on to

specific minerals in a pulp was established during the course of a research project concerned

with the phenomenon of slime coating in froth flotation [63].

It was claimed, for example, that at a pH of 8, hematite slimes would adsorb

selectively on to calcite in the presence of apatite, causing depression of the carbonate along

with an increase in apatite grade from 58 to 89% in the concentrate decreasing the overall

recovery of phosphate contents. The effects of pH and dispersants on the zeta-potentials of

calcite, dolomite and apatite were described elsewhere [64,65]. The control of various

parameters could lead to successful separations of such mixtures as phosphate-carbonate,

fluorite-quartz, metallic Pb-Cu [64], scheelite-calcite, slime cassiterite-quartz and mixed bulk

sulphides. By using a mineralogical-scale wet magnetic separator, after treatment at pH 11

with colloidal magnetite, a calcite-apatite mixture was separated at apatite recovery of 85.3%

and grade 99.3% [66,67].

1.4 Calcination Method

Calcination has been proposed in the past to treat carbonate bearing sedimentary

phosphate ores followed by slaking and rejection of the calcined oxide by desliming and

classification operations [68-79]. In the flash calcination method, thermal beneficiation was

carried out to see the effect of temperature, time and particle size for low-grade calcareous

35

phosphate rock. Under various parameter conditions the rock was heated and quenched

immediately. The concentrates obtained from the ore assaying 16% P2O

5 content after

desliming were enriched to a product containing 31-34% P2O

5 without using any expensive

chemicals or energy required in conventional flotation and calcination methods [69].

The high cost of energy required for calcination has always acted against the

implementation of such proposals. Depending on the efficiency of the process, calcination

may lead to almost complete disposal of the carbonates present in the phosphate rock.

However, three hours heating at 1000oC resulted in the decomposition of about 90% of the

calcareous and dolomitic gague materials depending on the size and nature of the raw

phosphate rock [69].

Beneficiation by calcination is one of the better-known processes. It is based on

the dissociation of the calcium carbonate gangue materials by thermal energy. In this case,

carbon dioxide is removed as a gas from the calcination kiln, while the calcium oxide and

most of the magnesium oxide are removed after quenching and desliming of the solids

leaving the kiln. By quenching the hot calcined rock in water, the major impurities usually

pass into the slime phase, which is then mechanically separated from the coarser high grade

'sand' concentrate fraction [70].

The essence of the calcination process may comprises crushing, sieving, washing (to reduce

clay and chlorides); classification, calcination at about 950oC, hydration of the lime,

separation of the lime, and finally further washing and drying [27,70]. Dust collection is

essential. On the other hand, calcination has the following disadvantages:

36

1. It may tend to decrease the solubility and reactivity of the calcined phosphate

rocks in the liquid phase of the acid reagent during the manufacture of phosphoric acid by the

wet process [71,72]. In other words, the calcined phosphate requires more retention time in

order to get the same solubility as that of non-calcined phosphate of the same grade. This

might be due to the changes in the chemical and physical nature of the phosphate rock,

caused by the high temperatures (above 900oC) used in the calcination process. It has been

found that chlorine, for example, is transformed after calcination into a much less soluble

form [73].

2. Calcination needs complete decomposition of the calcite and dolomite present

in the phosphate rock, otherwise hard unaltered particles of calcite and dolomite remaining

may dilute the phosphate contained in the concentrate, thus lowering the grade and increasing

potential acid consumption if the product were to be used for subsequent fertilizer

manufacture [70]. This means either more time will be needed for calcination at temperatures

between 900-950oC, or higher temperatures (that might cause sintering and fusion) would be

required. In the latter case, less slime may be formed and poorer rejection of impurities will

be attained resultantly.

3. The calcination and desliming processes usually result in concentration of silica

in the final concentrate product to more than the 3-4% desired for wet process phosphoric

acid [16]. The removal of the silica is essential and can add an extra cost to the process.

4. Upon hot quenching, the magnesium oxide may be transformed into an

insoluble magnesium hydroxide. The Mg content causes a number of problems during

fertilizer manufacturing [74].

37

5. Although repeated attrition, scrubbing and desliming processes result in higher

concentrate grades, the percent P2O5 recovery may decrease [70].

It is interesting to mention that China proposed a calcination plant using an

abundant supply of cheap coal. The interesting feature of the process was the proposed re-

conversion of the milk of lime and magnesium to solid carbonates by reaction with kiln gases

in a carbonation tower. The resulting solid carbonates were easily dumped reducing the

pollution caused by the tailings. Control of the pressure and CO2 concentration during the

carbonation stage was also permitted the production of a magnesium carbonate by-product

[75]. However, more research work may help to study the kinetics of calcination method

regarding various complex ores and more energy efficient processes [76,77] and advances in

the understanding of fluidized bed behavior [78,79] can also contribute to the

competitiveness of this method.

1.5 Electrostatic Upgrading

Phosphate beneficiation with electrostatic methods [80-85] is gaining in interest

especially in those areas characterized by water shortage where many important ores are

being mined. In these cases, upgrading of run-of-mine ore is often confined to dry

classification whereby the leaner fine fraction is rejected to waste deposits still containing a

considerable amount of phosphate matter. The process is based upon selective turbocharging

of the various mineral components followed by separation in a free-fall chamber. The main

features of the Turbocharger are represented by a reasonably high throughput capacity, a

good efficiency in the finest size range, considerable versatility, low energy consumption

even if heating is required, a modular arrangement and easy operation control, eventually

38

resulting in a relatively low cost of treatment per ton of ore. Results show that the

concentrates can be obtained with fairly good recoveries especially for the ore having

favorable liberation characteristics; and unfortunately the electrostatic separation of pyrite

from the phosphate phase is quite inefficient and thus the concentrate is polluted to a

significant extent [85].

1.6 Bioleaching of Phosphate Ores

The literature survey shows that various researchers [86-92] have studied the

bioleaching of low-grade phosphate rock. Due to a number of unit operations involved and a

relatively higher cost for production of phosphate fertilizers in phosphate industries,

biotechnology may be considered as an economical area in some cases of bioleaching

process. Since Winogradsky suggested the concept of autothrophy, various bioleaching

studies have been investigated [87]. Several microbial species may grow at extreme

conditions, including high acidity and low organic matter concentrations [88]. This

environment, through its high selectivity, may present a number of microbial species. The

predominating species present in these environments are those from the genus Thiobacillus,

especially T. ferrooxidans and T. thiooxidans [89]. Using ferrous iron, T. ferrooxidans may

reduce sulphur as an energy source and T. thiooxidans can only use sulphur-containing

compounds [90]. The presence of bacterial species, originally present in the pyritiferous

concentrate, was investigated using a 9K culture medium [91]. The microbiological

production of acid solutions from sulphide-containing minerals may be considered as an

equivalent technology to solubilize phosphate rocks for the production of phosphate

fertilizers in the fertilizer industrial sector [92]. The metabolic activity of these

39

microorganisms can be accelerated by a series of various factors that must be controlled to

achieve maximum biological action. Control of temperature, pH and the amount of oxygen

supplied, are the main required factors for the maintenance of an ideal leaching process;

otherwise the acid production may not be enough to convert the high phosphate content

soluble and neutralize the carbonate material present in the rocks [92].

Kokal [93] emphasized that mineralogical characterization is a significant factor

regarding the bioleaching of minerals. The low grade and poorly available sources of

phosphorus may be utilized using phosphate solublizaing bacteria. The solubelization of

phosphatic ores using microorganisms is a significant area of research regarding

biofertilization. T he literature indicates that the main bioleaching mechanism for solubilizing

calcium phosphate is related with the acidulation of the medium through biosynthesis and

production of various organic acids [94-101] as well as bioleaching due to being economical

and environmental considerations Delvasto, P.et al [102].

1.7 Partial Acidulation of Phosphate Ores

A fully acidulated phosphate rock is generally treated with the stoichiometric

amount of acid required to completely convert the insoluble phosphate mineral (apatite) to

the water-soluble form monocalcium phosphate monohydrate. On the other hand, a partially

acidulated [103-110] phosphate rock (PAPR), is generally treated with only a portion of the

stoichiometric amount of acid (expressed as a percentage, for example 50% PAPR).

The amount of water-soluble phosphate in a PAPR will therefore vary according

to the degree of acidulation, and a PAPR will contain a mixture of water- and citrate-soluble

40

phosphates [109]. The water-insoluble portion behaves according to the properties of the

source rock and may provide a residual phosphorus source that will gradually dissolve [108].

Partially acidulated phosphate rock has found a favorable response especially in

developing countries where, in recent years, use has increased. Sulphuric and phosphoric

acids are the most common acids used for partial acidulation. For economic reasons,

however, the use of sulphuric acid for acidulation is the more usual choice [109]. The

presence of iron and aluminum give rise to reaction products and decrease phosphate

availability [110]. Because of the unpredictable nature of the acidulation of phosphate rock,

mainly due to the complexity of the reaction and the influence of associated compounds in

the rock, it is necessary to conduct thorough test work on a rock if it is being considered for

partial acidulation, particularly since altering the process variables may significantly alter the

nature of the product. Two processes were expansively investigated for the manufacture of

sulphuric acid-based PAPR [108]: a run-of-pile process; and single-step

acidulation/granulation.

1.7.1 Run-of-Pile (ROP) Process for PAPR

The ROP process is the simpler of the two processes. In this process ground

phosphate rock, sulphuric acid and water (if needed) are fed continuously to a mixer, which

is usually of the pug mill type but may vary according to the properties of the feed rock.

Retention time in the mixer is relatively short, 30-60s, in order to avoid subsequent handling

problems, such as excessive agglomeration and caking of the material involved.

Continuous operation is generally preferable to batch operation, which has been

found to result both in unsatisfactory mixing between the acid and rock and in poor product

41

handling characteristics. It is possible to add other required materials to the mixer to either

improve the acidulation process or provide additional nutrients to the soil. It is necessary to

ventilate the mixer, den, and conveyor system and product storage area so that the fumes

(mainly fluorine gas) may be removed efficiently. These are than removed by scrubbing

system [108].

1.7.2 Single-Step Acidulation/Granulation (SSAG) Process

The single-step acidulation/granulation (SSAG) process is also easy and simple to

handle. The feeding system of materials to the process is similar to those for the run-of-pile

process (phosphate rock, acid and water) but include, additionally, recycled material - the

product screen undersize together with some product.

Depending upon the raw material and operating conditions the amount of recycle

in the feed is normally equal to a recycle/product ratio of about 2, which is generally found to

be enough to maintain acceptable granulation. Individual feeds are fed continuously and

directly to a granulator, which is usually of the rotary-drum type. The acid and water are

sprayed through separate inlets onto a rolling bed of phosphate rock in the granulator. The

amounts of acid and water are kept enough to achieve the required degree of acidulation and

granulation of the material involved. The residence time of the feed in the granulator is

usually in the range of 5 to 8 minutes. To maintain the granulation consistency, undersize

material, together with some product, may be recycled to the granulator. Exhaust gas from

the drier is first passed through a dust collector, in which most of the dust is removed, and

then through a wet scrubber to remove the remaining dust particles and any fluorine gas

[108].

42

1.8 Direct Application of Phosphate Rock

Direct use represents one of the simplest applications of phosphate rock to acid

soils. Technologically the method is very simple, after mining the ore is ground, screened

and applied directly to the soil without any further treatment generally. Finely ground

phosphate rock was successively used as a phosphate fertilizer for direct application to the

soil in many countries of the world [111-120], mainly where acid soils are characterized by a

high phosphorus fixation capacity due to the extreme phosphorus deficiency [111]. In the

acid soils the apatite dissolution conditions are favorable.

The choice of the binder can be important when the micronutrients are added to

the soil. Granulated product made with a soluble-salt binder break down and disperses in soil

to give a performance almost as effective as finely ground rock. However, considerable

variation in the apatite composition may be found in any one deposit, so extensive agronomic

analysis will be essential to assess a phosphate rock as a potential fertilizer for direct

application [112]. For agronomical effective the rock should be ground, typically to between

100 and 200 mesh (Tyler). The handling characteristics associated with this finely ground

dusty material can be improved by granulating the ground rock into larger separate particles

[112].

The International Fertilizer Development Center (IFDC), suggested a method to

granulate ground phosphate rock using a soluble salt binder. The results showed that the

process could further be improved in order to achieve a product with good mechanical

properties without losing its agronomic effectiveness [112,115]. The mini-granulation

process developed by IFDC was consisted of finely ground phosphate rock into spherical

granules using a binder. A number of binders were suggested, including soluble salts (such as

43

urea, potassium chloride, magnesium sulphate and magnesium chloride), mineral acids (such

as sulphuric acid or phosphoric acid) along with some organic materials [116]. Mini-granular

ground rock (48-150 mesh) was found to be promising regarding handling characteristics and

ensuring as effective a crop response as non-granular rock (-100 mesh). Larger granules (6-

14 mesh) did not show an improved effectiveness when incorporated into soil. However,

these granules were found to be considerably effective for surface application to grassland

[112,117].

Using a number of phosphate rock samples from South America, It was found that the

reactivity of a rock was dependent on the origin of the rock [119]. However, it is not simply

the origin of the rock that governs its reactivity, a number of associated fundamental factors

may also be involved significantly, including crystalline structure, amount of carbonate

present in the apatite matrix, association of other nutrients and physical properties such as

surface area, particle size and porosity [120].

1.9 Dissolution and Selective Leaching

A number of researchers studied acidulation and leaching of different phosphate

rocks [121-126]. Attempts to leach phosphate values directly from the low grade matrix,

thereby avoiding beneficiation losses, could result in phosphate extractions as high as 90%

when H2SO

4 was used as the leaching agent [121]. However, two disadvantages to this

approach were observed: crude acids produced by direct acidulation were not of enough

purity to be considered good fertilizer precursors [121,126], and the reaction of H2SO

4 with

the rock generated an insoluble residue of a very small average particle size, resulting in a

slow and difficult separation of the acid from the residue [121]. Performing a similar

44

approach, some other researchers acidulated the rock matrix in the presence of organic

solvents (non-aqueous), the organic compounds that could serve to remove hydrophilic

impurities in situ [127-131]. The results of these studies were found to be promising to

improve the acid purity but did not address the guidelines regarding the solid-liquid

separation problems. Warren Springs Laboratories, U.K, using H2SO

4 along with a careful

control of pH 3.0, suggested the selective leaching of carbonate impurities present in

phosphate rocks. The suggested process could be used to purify high-grade sedimentary ores

containing limited amounts of dolomite impurities. Although the process was technically

feasible, but any practical implementation could depend largely on the economic aspects of

the overall leaching process.

The dissolution study of various ores in inorganic acids has been investigated in

detail [132-137]. Liang et al. [135] used sulfuric acid to produce TiOSO4 and the results

showed that the acid concentration of 15.4 M could achieve an appreciable level of

dissolution of the ore. The results also indicated that the leaching kinetics could be described

by shrinking core model along with estimated activation energy of 72.6kJmol-1. The leaching

of hematite during the course of dissolution was also observed depending on the

concentration of the acid. At higher concentration of the acid the dissolution of iron was

found to be relatively low than the aluminum.

Ashraf et al. [136] used dilute phosphoric acid solution to study the chemical

kinetics of the reaction involved. Using unreacted core model, the results showed that the

value of mass transfer coefficient increased with an increase in particle radius. For example,

its value increased from 7.67 x10-6 to 8.07 x10-6 m/sec as the average size of the particle was

increased from 1.675 x10-4 to 16.55x10-4 m. However, the value of reaction rate constant did

45

not appear to be a strong function of the particle size as the change was found to be very

small.

Zafar et al. [137] used dilute hydrochloric acid solution to dissolve calcareous

material in low-grade phosphate rock. The results indicated that the acid concentration that

gave the promising dissolution level was found to depend on the liquid/solid (vol./wt. basis)

ratios used. It was found that the higher concentrations of the acid favor the dissolution of

more phosphate element rather than that of carbonates in the sample. The temperature

increased sharply during the first 8 minutes and then with a rapid fall it was found to be

constant after about 12 minutes of the dissolution time. The dilute hydrochloric acid could be

used to remove/reduce the calcareous material from low-grade phosphate rock as it rendered

P2O

5 content to marketable and industrially acceptable level. If higher/considerable amount

of phosphate is dissolved during the leaching of calcareous material, the milk of lime,

Ca(OH)2 can be used to easily recover the dissolved apatite by the controlling parameter of a

specified pH range of the recovery process [137].

For dissolution of various ores, the use of organic acids has been increased in the

recent past [138-148]. Abu-Eishah et al. [138], Sadeddin et al. [139] and Abu-Eishah et al.

[140] conducted the studies using dilute acetic acid solutions to leach out the carbonaceous

and dolomite gangue materials from the low-grade phosphate rocks. The published data did

not show any guidelines about the reaction kinetics of the leaching process. In the acetic acid

leaching process, a somewhat higher liquid/solid ratio was recommended to achieve the

desire grade phosphate rock. A higher liquid/solid ratio may result in a relatively lengthy

process including handling, filtration and regeneration of the used acid by using chemicals or

ion exchange reagents corresponding to the volume of the acetate solution from the leaching

46

process. To eliminate/reduce calcareous material from the low-grade phosphate rock, Zafar et

al. [141] used formic acid solution.. Using this technique, the P2O5% was easily raised by up

to 29% corresponding to a reduction of up to 69% in the calcium carbonate content at a

phosphate recovery of more than 73 wt%.

Economou and Vaimakis [142] used acetic acid to selectively leach the calcareous

material in low-grade phosphate rock from Epirus area (Greece) and the results suggested

that the dissolution process could take place in two steps. The first step was found to be

relatively fast attributing to the dissolution of freely available calcite and/or spaces on the ore

particles. While, in the second one the leaching results were found to depend on the presence

of calcite in the interspaces between phosphopeloids. Using dilute acetic acid solutions,

Economou et al. [143] investigated the mechanism of selective leaching of calcite in low-

grade phosphate ore, and the analysis of the results indicated the model equation of a double

exponential decay fitting the experimental data. The kinetic analysis results showed that the

controlling step of calcite dissolution in the low-grade phosphate rock was a chemical

change. On the basis of some assumptions, the experimental results also concluded that the

decomposition of carbonic acid (H2CO3) into carbon dioxide (CO2) and water (H2O) was

most likely to be the rate-controlling step of the overall mechanism. In another similar work

[144] based on the different pH regions, it was suggested that the mass transfer had a

significant role in the rate-determining step of the overall reaction. Depending on the pH

values, two different reaction mechanisms were claimed along with two different model

equations fitting the corresponding experimental data. However, it was again found that the

decomposition of carbonic acid (H2CO3) into carbon dioxide (CO2) and water (H2O) could be

the rate-controlling step of at least two pH regions, 3.96-4.95 and 4.95-5.50.

47

Fredd and Fogler [145] investigated the kinetics of calcite leaching in acetic acid

solutions. The analysis of the data showed that, below about pH 2.9, the dissolution of calcite

was affected by the transport of both reactants and products. However, above about pH 3.7,

the rate of leaching was controlled significantly by the kinetics of the surface reaction.

Demir et al. [148] and Oral et al. [150] investigated the leaching kinetics of

magnesite ores using citric and acetic acids respectively. The experimental data, in both of

the cases of leaching agents, indicated the leaching process as surface reaction control. Using

citric acid, the activation energy of the leaching process was found to be 61.35kJ mol-1. On

the other hand, the activation energy of the same process using acetic acid was reported to be

78.40 kJ mol-1. The reason for the gap of activation energy using these two organic acids was

not addressed. However, it may be attributed to the difference in chemical nature of these

acids or the type and nature of various chemical species present in the ore samples.

Organic acids may be used for the selective leaching of carbonaceous material in

low-grade phosphate rocks. Using organic acids, lower losses of phosphate element may be

expected in dissolution processes, as the leaching capability of organic acids is relatively

weak as compared to inorganic acids. Apatite reserves are located in the North East of the

country, e.g. Kakul, Lagarban, Oatkanala, Batkanala, East and South phosphorites. The P2O5

content of the rock varies from deposit to deposit depending on the type and nature of the

impurities. The rock of these areas contains an objectionable amount of calcareous material;

therefore, it can’t be used directly for acidulation processes unless its P2O5 content is

increased up to an industrially acceptable level. No proper guidelines are available in

literature regarding the selective leaching kinetics of calcareous phosphate rock using organic

48

acids. The present work aimed to study the possibility for using succinic acid to selectively

dissolve calcareous material in the low-grade calcareous phosphate rock.

49

CHAPTER 2

2. CHEMISTRY OF THE LEACHING PROCESS

Calcareous and dolomite impurities present in low-grade phosphate rock can be

reduced or removed by selective leaching process. The type and nature of impurities in low-

grade phosphate rock varies from deposit to deposit through out the world. The type and

level of impurities may vary with in the same deposit as well. Therefore, every phosphate

rock deposit may be considered as special case with respect to beneficiation technique and

technology. For selective dissolution of calcareous and dolomite minerals both the nature and

concentration of a leaching agent are the important factors so that the phosphate element

present in the rock should not be lost due to the decomposition by reaction with the leaching

agent.

As mentioned before the dissolution study of various ores in inorganic acids has

been investigated in detail. However, the studies concerning the solubility in organic acids

are quite limited. Only a few researchers have used dilute organic acids for the leaching of

carbonaceous and dolomite gangue materials in low-grade phosphate rocks. The literature

indicates that the use of organic acids is increased in the recent years. The dissolving ability

of organic acids is relatively weak, however, these weak acids show high selectivity and may

be used as leaching agents for low-grade calcareous phosphate rocks. On the other hand, in

scale up studies with inorganic acids, high CO2 pressure and froth forming produced due to

fast dissolution may lead to some risks. The organic acids may be a promising extracting

agent as the leaching is carried out at intermediate acidic conditions (pH 3-5) and the

degradation of these acids is biologically easy [152]. However, in the industrial processes at

50

relatively higher temperatures organic acids may cause a little corrosion effect [153]. The use

of organic acids at relatively high temperature may be limited because of low boiling

temperatures and their decomposition [148,150].

Calcareous gangue materials may be reduced by leaching the low-grade

phosphate rock with dilute organic acids such as citric acid, acetic acid, formic acid, succinic

acid, etc. depending on various conditions of the leaching process as well as nature and type

of the phosphate ores. Any strong acid when mixed with ground phosphate rocks will react

with the main constituents of the phosphate and change them to other compounds. Therefore,

strong organic acids cannot be used to selectively remove calcareous material in low-grade

phosphate rocks as they may dissolve the phosphate apatite along with the leaching of

calcareous materials. The strong acids may not react, for example, with pure calcium

carbonate due to the large polarity of O-H bond present in these acids. However, in dilute

solutions, water molecules can tend to decrease the effect of polarity of the O-H bond

resulting in selective leaching of calcareous material in the rock due to higher degree of

ionization. The selected leaching agent should not attack the phosphate element. After the

leaching process, the liquid phase should be easily separable for regeneration and recycling.

On this basis, dilute succinic acid is found to be one of the promising leaching

agents for this purpose. In this case, dilute succinic acid has been used to selectively leach the

carbonatic phosphate rock. The salts of succinic acid are, depending on temperature, soluble

in water and may be separated easily from the leached solid phosphate product by filtration.

These characteristics allow the succinic acid molecules to react with calcium carbonates, but

not with the tricalcium phosphate itself appreciably. The dilute acid may be less destructive if

some of the produced solution is added to the reaction solution, or recycled to the process.

51

No physical or chemical changes in the phosphate rock, or considerable losses of its

tricalcium phosphate content, have been observed under lower concentrations of succinic

acid about up to 6-9 %.

The reaction between succinic acid and calcium and/or magnesium carbonates can

be written as follows:

CaCO3+C4H6O4Ca (C4H4O4) +CO

2+H

2O (2.1)

MgCO3+C4H6O4Mg (C4H4O4) +CO

2+H

2O (2.2)

The CO2 is evolved as a gas, while the calcium and magnesium succinate by-

products are soluble in the leaching solution. Reaction (2.1) and (2.2) may include the

formation of instable carbonic acid, which is then after soon decomposed into CO2 and H

2O;

and reaction for the other impurities depending on the nature and composition of the raw

phosphate. These two main reactions represent in fact a lumping of a larger number of steps.

The simplest detailed mechanism necessary for the normal understanding of this system may

be given as follows:

(a) Ionization process of HOOCCH2CH2COOH may be described by the equation:

C4H6O42H++C4H4O4

2- (2.3)

(b) Diffusion of H+ ions through the liquid to the exposed surface of the rock particle.

(c) H+ ions attack on the particles of the calcareous and dolomitic gangue material in the

rock.

nH++CaCO

3H

2CO

3+(n-2)H

++Ca

2+ (2.4)

(n-2)H++MgCO

3H

2CO

3+ (n-4) H

++Mg

2+ (2.5)

52

The H+ ions taking part in these reactions may 'belong' to the succinic acid as well as to

the carbonic acid, which can be formed in the leach slurry.

(d) Ionization of H2CO

3: When carbon dioxide dissolves in water, most of it is present as

CO2 molecules rather than H

2CO

3 molecules. However, if a small amount of carbonic acid is

formed, the ionization, which is fast, process and can be described by the following

equations:

CO2+H

2OH

2CO

3 (2.6)

H2CO

3H

++HCO

3

-

(2.7)

H++HCO

3

-2H

++CO

3

2- (2.8)

2H++CO

3

2-CO

2+H2O (2.9)

(e) Diffusion of products from the reaction sites to the bulk of the liquid.

(f) Reaction between Ca2+

, Mg2+

and C4H4O42-

Ca2+

+C4H4O42-Ca (C4H4O4) (2.10)

Mg2+

+C4H4O42-Mg (C4H4O4) (2.11)

The formation of calcium and magnesium succinate will depend on the various factors,

such as concentration of the acid, reaction time, temperature, nature and grain size of the raw

phosphate rock used for the selective dissolution process. The selective leaching process can

be simplified using the following general equations:

MCO3(s) +H2Y (aq)MY(s/aq)+CO2+H2O (2.12)

M3 (PO4)2+3H2Y3MY(s/aq) +2H3PO4 (2.13)

Where M=Ca2+and/or Mg2+, Y=H2C4O42-(succinate)

53

The selectivity should be such that the second reaction does not take place, while

the first reaction goes to completion or equilibrium depending on solubility products

(Ksp=8.35, 7.46 for CaCO3 and MgCO3 respectively) and acidity constants for succinic acid

(pK1= 4.21, pK2=5.64 at 25 oC) and CO2 (pK1=6.35, pK2=10.33 for H2CO3). The values of

these constants will depend on the reaction temperature and the kinetics may control the

overall success of the leaching process. The CO2 produced during the reaction process can be

separated and under normal conditions the equilibrium for reaction (2.12) lies far to the right

and which may be considered as an irreversible reaction.

The thermodynamic calculations are important for deep understanding the

chemical nature of leaching and leaching reactions related to the environmental conditions

such as temperature, acid concentration, solid particle size and liquid/ solid ratio of the

leaching process. For these reactions, the quantitative analysis is given in chapter 4 under

result and discussion.

54

CHAPTER 3

3. MATRRIAL AND METHODS

3.1 Sample Collection

Apatite reserves are located in the North East of the country, Hazara

Administrative Division of Pakistan. These reserves may be categorized as Kakul, Lagarban,

East phosphorite, South phosphorite, Oatkanala and Batkanala phosphorite. The rock

phosphate of all these areas is generally grey to dark grey, hard and compact. However, the

phosphate rock in the Lagarban area is mostly dolomitic in nature as opposed to the cherty

base of the Kakul phosphate rock. Lagarban rock is oolitic, the matrix and cement being

composed mainly of collophane with some quartz, haematite, limonite and chalcedony. It

also contains a few veins of secondary phosphate [154,155]. The sample, ranging in weight

from 2-3 kg, was collected from Shahzad and Lagrarian mines as in Figs. 3.1.1-3.1.6.

55

Fig 3.1.1 Phosphate rock deposit of shahzad mines.

56

Fig 3.1.2 Phosphate rock mining at shahzad mines.

57

Fig 3.1.3 Phosphate rock deposit of Largrarian mines.

58

Fig 3.1.4 Phosphate rock mining at Largrarian mines.

59

Fig 3.1.5 Phosphate rock deposit of Puttiann mines.

60

Fig 3.1.6 Phosphate rock mining at Puttiann mines.

61

3.2 Size Reduction and Sieving

Phosphate ore collected from the above mentioned area was in the form of lumps

of different weights and grain size. The sample was crushed in jaw crusher to reduce the size.

After crushing, ball mill was used for further grinding to get fine particle sizes. Using U. S.

Tyler Standard Sieves, various size fractions of the sample were collected for analysis. The

fraction of each particle was separated, weighed and percentage of each weight fraction was

calculated as shown in Table 3.2.1 and Fig. 3.2.1. The analysis was also carried out regarding

average size, mass fractions and cumulative mass fraction as shown in Table 3.2.2 and Fig.

3.2.2.

Table 3.2.1 Mesh size, average size, weight fraction and weight percent.

Mesh

No

Size

(mm)

Average size

(mm)

Weight fraction Weight percent

-24+32 # 0.701+0.495 0.598 293.0 20.99

-32+42 # -0.495+0.351 0.423 263.8 18.91

-42+60 # -0.351+0.246 0.299 233.0 16.70

-60+80 # -0.246+0.175 0.210 189.0 13.54

-80+115 # -0.175+0.125 0.15 187.5 13.44

-115 # -0.125 0.125 229.0 16.41

62

12

13

14

15

16

17

18

19

20

21

22

0 0.1 0.2 0.3 0.4 0.5 0.6 0.7

Average size (mm)

Wt %

of r

ock

Fig. 3.2.1 Average particle size and wt% of phosphate rock sample

63

Table 3.2.2 Average size, mass fractions and cumulative mass fraction

Average size (mm) Mass fraction (wi/wt) umulative mass fraction.

0.598 0.210 0.21

0.423 0.189 0.40

0.299 0.167 0.57

0.210 0.136 0.70

0.15 0.134 0.84

0.125 0.165 1.00

64

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

1

1.1

0.1 0.2 0.3 0.4 0.5 0.6 0.7

Average size (mm)

Cum

ula

tive m

ass frac

tion

Fig. 3.2.2 Average particle size and cumulative mass fraction of phosphate rock

65

3.3 Sample Preparation for Analysis

5.0 gm sample of the ground phosphate rock was placed in an electric furnace at

about 110 oC to remove the moisture content before the further analysis of the sample. The

fire-clay crucible for electric furnace was washed, cleaned, preheated (to dry at 110oC), and

then cooled to the room temperature. These sample fractions were analyzed for the main

conditions, namely, P2O

5 content, acid insoluble residue (AIR), loss on ignition (LOI) and

CO2 content as shown in Table 3.3.1.

Acid insoluble Residue has been defined as the amount of the residue that

remained unreacted in the sample, after being treated with a standard HNO3/HCl solution and

ignited at 950C. The loss on ignition has been defined as the decrease in weight of the

sample after being ignited from 550 to 950C. It also refers to the amount of CO2 in the

sample, either before or after the leaching process. The dissolved fraction of the calcareous

material during the leaching process was calculated by using the following equation 3.3.1:

sampleoriginalincalciumofamountTotal

solutiontheincalciumofAmount exp (3.3.1)

And the recovery of P2O5 content was calculated by using the following equation 3.3.2.

100%

%covRe

52

5252

SampleOriginalinOPSampleOriginalofWt

SampleLeachedinOPSampleLeachedofWteryOP

(3.3.2)

Conventional [156,157] as well as instrumental analysis techniques such as

Scanning Electron Micrographs (SEM), Energy Dispersive X-ray Analysis (EDX) and

Atomic Absorption Spectrophotometer (AAS) have been used for analysis. The analysis of

66

the sample is given as shown in Table 3.3.1. The effect of average particle size on P2O5, AIR,

LOI and CO2 contents has been studied as shown in Figs. 3.3.1-3.3.4.

Table 3.3.1 Analysis of the rock samples

Sample (mesh) Size (mm) P2O5 (%) AIR (%) LOI (%) CO2 (%)

-24+32 -0.701+0.495 16.12 2.43 29.97 27.89

-32+42 -0.495+0.351 16.88 2.67 29.17 26.97

-42+60 -0.351+0.246 17.73 2.75 28.13 25.71

-60+80 -0.246+0.175 18.69 2.79 26.93 24.57

-80+115 -0.175+0.125 19.31 2.81 25.87 23.63

67

15.5

16

16.5

17

17.5

18

18.5

19

19.5

20

0.1 0.2 0.3 0.4 0.5 0.6 0.7

Average size (mm)

P2O

5%

Fig. 3.3.1 Effect of average particle size on P2O5 content

68

2.4

2.45

2.5

2.55

2.6

2.65

2.7

2.75

2.8

2.85

0.1 0.2 0.3 0.4 0.5 0.6 0.7

Average size (mm)

AIR

%

Fig. 3.3.2 Effect of average size of particle on AIR content

69

25.5

26

26.5

27

27.5

28

28.5

29

29.5

30

30.5

0.1 0.2 0.3 0.4 0.5 0.6 0.7

Average size (mm)

LOI%

Fig. 3.3.3 Effect of average particle size on LOI content

70

23

23.5

24

24.5

25

25.5

26

26.5

27

27.5

28

28.5

0.1 0.2 0.3 0.4 0.5 0.6 0.7

Average size (mm)

CO

2%

Fig. 3.3.4 Effect of average particle size on CO2 content

71

3.4 Selective Leaching with Lactic Acid

As mentioned before in chapter 3 that two different samples collected from

Shahzad and Lagrarian mines. To study the effect of lactic acid on the selective leaching of

calcareous material in low-grade phosphate rock the Lagrarian sample was used. It is also

known as Puttiann mine, which is present in Hazara Administrative Division (district Abbot

Abad). The sample was crushed and sieved for analysis. Various fractions of different

particle size were collected using US Tayler standard sieves for further analysis and kinetic

study as shown in Table 3.4.1. The effect of average particle size on P2O5, CO2, LOI and AIR

contents has also been given in Figs. 3.4.1-3.4.4.

72

Table 3.4.1 Analysis of the rock samples

Size (mm) P2O5% CO2% LOI% AIR%

0.503 15.89 28.29 30.74 2.13

0.356 16.23 27.75 30.12 2.29

0.2515 17.32 26.87 29.11 2.63

0.1775 19.67 24.67 26.83 2.87

0.1255 22.39 20.13 22.13 2.97

73

15

16

17

18

19

20

21

22

23

0.1 0.2 0.3 0.4 0.5 0.6

Average particle size (mm)

P2O

5%

Fig. 3.4.1 Effect of average particle size on P2O5 content

74

15

17

19

21

23

25

27

29

31

33

35

0.1 0.2 0.3 0.4 0.5 0.6


CO

2 %

Fig. 3.4.2 Effect of average particle size on CO2 content

75

20

22

24

26

28

30

32

0.1 0.2 0.3 0.4 0.5 0.6


LO

I%

Fig. 3.4.3 Effect of average particle size on LOI content

76

2

2.1

2.2

2.3

2.4

2.5

2.6

2.7

2.8

2.9

3

0.1 0.2 0.3 0.4 0.5 0.6


AIR

%

Fig. 3.4.4 Effect of average particle size on AIR content

77

3.5 Experimental Procedure

To study the reaction kinetic, the particle size fraction (-80+115# (-0.175+0.125

mm) was used in a well mixed spherical glass batch reactor (500 ml), which was equipped

with a mechanical stirrer having a digital controller unit and a constant temperature bath. For

each run, a specific concentration of succinic acid (8%) with liquid/solid (L/S) ratio of 7:1

was slowly fed into the reactor vessel containing 5.0 gm of the rock sample. After

completion of each reaction, the reaction vessel was immediately placed in an ice bath to stop

the reaction. The leach slurry was separated by filtration after the leaching process. The solid

and liquid phases were then weighed and analyzed. The Ca2+ content in the leach solution

was determined by (AAS) A-1800. Hitachi. The solid phase after the leaching process was

also analyzed by using calcinations method at 950C to confirm the results for the conversion

of calcareous material and degree of beneficiation of low-grade phosphate rock. The spent

succinic acid from the liquid phase (containing calcium succinate) was recovered by

sulphuric acid. After treating the calcium succinate with sulphuric acid, insoluble calcium

sulphate was separated from the recovered succinic acid by filtration as shown below.

Ca (C4H4O4)+H2SO4CaSO4+C4H6O4 (3.5.1)

To see the effect of different parameters (which were not studied in the case of

succinic acid leaching) i.e. effect of acid concentration, liquid/solid ratio and particle size,

and the sample with particle size fraction of -100+150 mesh (-0.147+0.104 mm) was treated

with lactic acid. Experimental procedure was adopted as mentioned before in the case of

succinic acid. A number of experiments were carried out by using different acid

concentration, liquid/solid ratio and various size particles of calcareous material at different

reaction temperatures with a constant stirring rate of 350 rpm.

78

For a number of reaction run, the reaction vessel was placed in an ice bath to stop

the reaction, after heating at different temperatures. The stop point of temperature may a little

bit different at various levels. However, small variation in the stop point of temperature is not

expected to influence the experimental results significantly.

79

3.6 CHAMICALS USED

Chemical Name Formula Mol Wt %age Purity Company

Ammonium

Molybdate

(NH4)2MoO4 1235.86g/mol Extra pure MERK

Succinic Acid C4H6O4 118g/mol 99.9% MERK

Sulphuric Acid H2SO4 98.08g/mol 97% Fluka Riedal

Nitric Acid HNO3 63.01g/mol 65% Riedal-ed Hasen

Hypochloric Acid HClO4 11.58 gm Extra pure BDH

Lactic Acid C3H6O3 90.08 gm 90% MERK

Ammonium

Hydroxide

NH4OH 35g/mol 35% MERK

Sodium Hydroxide NaOH 40g/mol 97% Fluka

Phenolphthalein C20H4O4 318.32 gm PH=8.3-10 Fluka

Calcium Nitrate Ca(NO3)2.4H2O 236.15g/mol Extra pure Fluka

80

3.7 INSTRUMENTAL TECHNIQUES USED FOR ANALYSIS

3.7.1 Atomic Absorption Spectrophotometer

In order to find out the concentration of calcium in the aqueous phase, an atomic

absorption spectrophotometer (AAS) A-1800 Hitachi (Fig. 3.8.1) was used to study mass

transfer and reaction kinetics in the present leaching process. Metals are estimated by the

atomic absorption spectrophotometer based on two different modes; 1 Absorption mode. 2.

Emission mode.

3.7.2 Absorption Mode

In absorption mode, the flame of the atomic absorption spectrophotometer

converts the incoming aerosol into atomic vapor, which can then absorb light from the

primary light source (the hollow cathode lamp). The amount of radiation absorbed by the

metal atoms is measured to find out the concentration of metal in the solution.

3.7.3 Emission Mode

This technique performs the flame to do two jobs; 1. The flame converts the

aerosol into an atomic vapor; 2. Thermally activated atoms are elevated to an excited

electronic state. These atoms emit radiation when they return to the ground state. The

instrument detects the radiation. Then the intensity of radiation is related to the concentration

of the element in solution.

81

Fig.3.7.1 AAS (A-1800) Hitachi used for analysis.

82

3.7.4 Chemical Interference

Chemical interference may affect the efficiency of production of neutral atoms in

the flame causing effects on both absorption and emission in a similar manner [158]. One of

the most general types of chemical interference is the formation of refractory compounds

with the test element, usually by an anion in the aspirated solution [159] resulting in a

decreased signal. For example, phosphate may react with calcium ion to produce calcium

phosphate in air-acetylene flame. In such a case, the determination of calcium cannot be

considered as a reliable measure due to the interference resulting in a decreased signal.

Adding an appropriate releasing agent may generally minimize chemical

inferences. These agents either compete for the interfering substance or displace it from the

test element. For example, a sufficiently high concentration (about 1% depending on the

actual phosphate levels) of strontium or lanthanum chloride may be added to the sample

solution to eliminate phosphate interference with calcium absorption. The strontium or

lanthanum will preferentially react with the phosphate and prevent its reaction with the

calcium or high concentration of EDTA may be added to form a chelate with the calcium and

prevent its reaction with phosphate. Because addition of an external reagent may sometimes

change the total composition, it should also be added to the standard solutions prepared for

the test metal. The use of high temperature flames can frequently eliminate chemical

interferences. Phosphate interference on calcium, for example, does not occur in the nitrous

oxide-acetylene flame [160].

83

3.7.5 Calibration of Atomic Absorption Spectrophotometer

The calibration curve procedure was adopted to determine the concentration of

the concerned element in the present work. This method is generally selected where

interferences are either absent or are fully controlled. A series of standard solutions of

calcium nitrate (standard solution for the calibration of the atomic absorption

spectrophotometer, analytical grade, supplied by B. D. H) in the range of 5 to 30 ppm were

prepared with distilled water. The atomic absorption spectrophotometer was calibrated using

these standard solutions and plot of metal concentration versus absorbance for each standard

was determined.

The calibration curve (Fig. 3.8.2) for the calcium was found linear over this range

of the standard solutions used. The determination of calcium concentration in aqueous

solution was carried out at a wavelength of 422.7 nm in absorption mode using a hollow

cathode lamp specific for calcium. Adjustment of the burner head relative to the light path of

the instrument was necessary to obtain maximum absorption. The lamp current and silt width

were set at 15 mA and 0.7 nm respectively. Flow rates of acetylene 18 and air at 55 l/min

gave a persistent non-luminous flame. For the analysis of calcium, the solutions were diluted

so as to fall within the range of 5 to 30 ppm.

84

Absorbance

0

0.2

0.4

0.6

0.8

1

0 5 10 15 20 25 30 35Calcium(ppm)

Abs

orba

nce

Fig.3.7.2 Correlation curve for Calcium nitrate versus Absorbance.

85

3.8 Scanning Electron Microscopy

Scanning electron microscopy (SEM) is used for inspecting topographies of

specimens at very high magnifications using the equipment, which is known as scanning

electron microscope. In the present case, the scanning electron microscopy has been used to

characterize morphological features of the sample as shown in Fig. 3.9.1. SEM micrograph

of sample explains many prominent characteristics features. Both dark and bright colored

morphological features are evident. The surface is extremely rough with numerous

disclamations and serrated features mostly in the form of cuts. Most of the pore mouths are

blind. However, some pore drilled into the matrix can be seen. The micrograph reveals

calcareous and siliceous inclusions. The black region embedded in the bright calcareous or

matrix evident the presence of phosphorus in the sample.

3.8.1 Energy Dispersive X-ray Analysis

A SEM may be equipped with an EDX analysis system to enable it to carry out

compositional analysis of samples. EDX Analysis stands for Energy Dispersive X-ray

Analysis. It is sometimes may be referred to also as EDS or EDAX analysis. The EDX

analysis system works as an integrated feature of a scanning electron microscope (SEM), and

cannot work on its own without SEM. EDX analysis is useful in identifying materials and

contaminants, as well as estimating their relative concentrations on the surface of the sample.

The higher a peak in a spectrum, the more concentrated the element is in the sample. Energy

Dispersive X-ray Analysis (EDX) of the sample is shown in Fig. 3.9.2. The EDX of the

sample evaluates the existence of calcareous and siliceous nature of the material under study.

It also indicates the presence of small amount of iron within the matrix of the apatite.

86

Fig.3.8.1 SEM micrograph of the sample

87

Fig. 3.8.2 EDX spectrum of the sample

88

CHAPTER 4

4. RESULTS AND DISCUSSION

4.1 Effect of Reaction Time on the Leaching Process at Different Temperatures

The selective leaching of calcareous material in low-grade phosphate rock may

depend on the various parameters such as reaction time, temperature, acid concentration,

liquid/solid ratio, particle size and stirring speed of the reaction mixture. A number of

experiments have been carried out to see the effect of reaction time on the dissolution of

calcareous material in low-grade phosphate rock at various temperatures as shown in Table

4.1.1 and Figs. 4.1.1-4.1.5. The experimental results show that the degree of selective

leaching of calcareous material increases with an increase in the reaction time depending on

the reaction temperature. According to the experimental results the optimum leaching time

has been found to be between 45 to 55 minutes at 60oC using the acid concentration of 8%

and a liquid/solid ratio of 7:1.

As the temperature of reaction is increased, the conversion of calcareous material

increases with a decrease in reaction time required to reach the equilibrium. This indicates

that the higher rates as well as higher solubility at higher temperatures. However, the results

show that the rate of dissolution is not appreciable after 60-70oC. This situation may be

attributed due to the establishment of equilibrium or the formation of a solid product, which

prevents further leaching of the calcareous material in the rock. It shows that higher

temperatures may tend to decrease the solubility of succinate product along with the

contamination of CO2 gas stream with water and succinic acid vapors, which are vaporized

during the reaction relatively at higher temperature.

89

Table 4.1.1 Effect of reaction time on leaching of calcareous (XB %) material at various

temperatures.

Time (min) 40oC 50oC 60oC 70oC 80oC

5 3.0 7.0 11.0 15.0 19.0

15 9.0 19.0 27.0 37.0 56.1

25 12.0 27.0 37.0 64.21 83.36

35 13.0 37.0 56.0 77.3 96.41

45 21.0 36.39 65.0 81.48 98.94

55 17.73 38.59 73.0 97.0 -

65 17.21 34.15 73.0 97.0 -

90

0

5

10

15

20

25

0 10 20 30 40 50 60 70

Leaching reaction time(min)

XB(%

)

Fig. 4.1.1 Effect of reaction time on the conversion (XB %) of calcareous material at 40oC

91

0

5

10

15

20

25

30

35

40

45

0 10 20 30 40 50 60 70


XB

(%)

Fig.4.1.2 Effect of reaction time on the conversion (XB %) of calcareous material at 50oC

92

10

15

20

25

30

35

40

45

50

55

60

65

70

75

0 5 10 15 20 25 30 35 40 45 50 55 60 65 70


XB(%

)

Fig. 4.1.3 Effect of reaction time on the conversion XB % of calcareous material at 60oC

93

0

10

20

30

40

50

60

70

80

90

100

0 10 20 30 40 50 60 70


XB(%

)

Fig. 4.1.4 Effect of reaction time on the conversion XB (%) of calcareous material at 70 oC

94

15

25

35

45

55

65

75

85

95

105

0 10 20 30 40 50 60 70

leaching reaction time(mni)

XB

%

Fig. 4.1.5 Effect of reaction time on the conversion XB (%) of calcareous material at 80oC

95

4.2 Effect of Reaction Temperature on Dissolution

To see the effect of temperature on selective leaching of calcareous material at

various acid concentrations, a number of experiments have been performed with stirring

speed of 350 rpm using liquid-solid ratio (L/S) 7:1 as shown in Table 4.2.1 and Figs 4.2.1-

4.2.5. The results indicate that the selective leaching process is sensitive to reaction

temperature depending on the acid concentration. The results show that the rate of dissolution

increases with an increase in the acid concentration. However, the dissolution rate of

calcareous material in low-grade phosphate rock is not much higher after 8% concentration

the leaching agent used.

In the leaching of colemanite ore with acetic acid [161], the results showed that as

the acid concentration in the leaching process was increased, the rate of formation of the

product increased by attaining the saturation value near the solid particle along with the

appearance of a sparingly soluble solid film layer around the particle resulting in a decrease

in the leaching process. However, in the present case of selective leaching the acid

concentration has not been increased more than 7-8%. It has been observed separately that by

increasing the acid concentration to about 9-10% and above, the acid starts to dissolve the

phosphate element itself appearing as losses in the filtrate solution. Therefore, the acid

concentration has not been increased more than 7-8% in the present case of selective

leaching.

4.3 Effect of Acid Concentration on P2O5 and CO2 Contents

The concentration of succinic acid can play an important role to selectively

dissolve the calcareous material in a low-grade phosphate rock. An increase in the

96

concentration of the acid is expected to increase the P2O5% content of a low-grade phosphate

rock, but the higher concentration of succinic acid can effect the selective leaching of the

calcareous material and the mechanism of the reaction kinetic. A number of experiments

have been carried to see the effect of acid concentration on the P2O5 and CO2 Contents as

shown in Table 4.3.1 and in Fig 4.3.1.

The experimental results show that the P2O5 content increases with an increase in

the acid concentration. The acid concentration that gives the most promising leaching results

has been found to depend on the liquid/solid ratio used in the leaching process. An acid

concentration of 8% with a liquid/solid ratio of 7:1 appears to be the most promising. As

mentioned before that when the succinic acid concentration increases to 9-10% and above,

the acid starts to considerably dissolve the apatite in the phosphate rock, which indicates that

there should be an optimum acid concentration to increase the degree of beneficiation by

avoiding an attack on apatite in the rock.

The results show that as the acid concentration is increased up to 9%, the increase

in the P2O5 content is not significant, which may due to an increased effect of polarity of the

O-H bond present in the acid molecules. On the other hand, the CO2% decreases

correspondingly with an increase in the acid concentration up to a certain limit and after that

the decrease in CO2% content is not much higher.

4.4 Effect of Liquid/Solid Ratio on P2O5 and CO2 Contents

A number of experiments haven been carried out to see the effect of liquid/solid

ratio (vol./wt. basis), on the degree of beneficiation of low-grade calcareous phosphate rock

as shown in Table 4.4.1 and Figs. 4.4.1-4.4.2. An increase in the liquid/solid ratio (vol./wt.

97

basis) may be expected to decrease the CO2% content resulting in an increased degree of

beneficiation of a low-grade calcareous phosphate rock. In the leaching process, the

liquid/solid ratio has been varied from 4:1 to 9:1. The experimental results show that P2O5

content increases as the liquid/solid ratio increases in the leaching process along with

corresponding reduction in CO2% content. The results indicate an appreciable increase in the

P2O5 content with liquid/solid ratio from 4:1 to 7:1, after which the degree of beneficiation is

not much higher.

On the other hand, after the liquid/solid ratio of 7:1 the corresponding reduction in

CO2% content is not appreciable. Therefore, the liquid/solid ratio of 7:1 has been found to be

promising depending on the reaction conditions in the leaching process. Depending on the

reaction conditions, the results indicate that both of the parameters (acid concentration and

liquid/solid ratio) should be optimized in order to increase the efficiency of the leaching

process reducing the risk of apatite dissolution.

The summery of the optimal parameters regarding calcareous phosphate rock and

succinic acid are given below.

Optimum time required for completion the reaction =55min

Optimum liquid/solid required for completion the reaction =7:1

Optimum acid concentration required for completion the reaction =8%

Optimum temperature required for completion the reaction =60oC

Optimum rpm required for completion the reaction =350

98

Table 4.2.1 Effect of temperature on selective leaching XB % of calcareous material at

various acid concentrations

Tem. (OC) Acid 6% Acid 7% Acid 8% Acid 9% Acid 10%

40 1.88 4.70 7.50 11.0 17.0

50 4.89 11.88 18.90 25.0 28.70

60 6.12 16.87 27.65 35.67 39.87

70 9.88 26.78 41.26 44.0 53.78

80 14.56 31.26 49.87 55.0 59.87

99

0

0.02

0.04

0.06

0.08

0.1

0.12

0.14

0.16

35 40 45 50 55 60 65 70 75 80 85

Leaching reaction temperature oC

Fra

ctio

n (xB

) dis

solv

ed

Fig. 4.2.1 Effect of temperature on dissolution (xB) with 6% succinic acid concentration and

liquid/solid ratio 5:1 for 30 min

100

0

0.05

0.1

0.15

0.2

0.25

0.3

0.35

35 40 45 50 55 60 65 70 75 80 85


Fra

ctio

n (

xB)

diss

olv

ed

Fig. 4.2.2 Effect of temperature on dissolution (XB) with 7% succinic acid concentration and


101

0

0.1

0.2

0.3

0.4

0.5

0.6

35 40 45 50 55 60 65 70 75 80 85


Fra

ctio

n (x B

) dis

solv

ed



102

0

0.1

0.2

0.3

0.4

0.5

0.6

35 40 45 50 55 60 65 70 75 80 85


Fra

ctio

n (x

B)

dis

solv

ed



103

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

35 40 45 50 55 60 65 70 75 80 85


Fra

ctio

n (xB

) dis

solv

ed

Fig. 4.2.5 Effect of temperature on dissolution (XB) with 10% succinic acid concentration

and liquid/solid ratio 5:1 for 30 min

104

Table 4.3.1 Effect of acid concentration on P2O5 and CO2 Contents

Concentration (%) P2O5% CO2%

2 19.9 23.1

3 22.5 17.6

4 26.2 11.7

5 27.9 9.0

6 29.1 7.4

7 30.9 5.5

8 31.5 5.1

9 31.9 5.0

105

15

17

19

21

23

25

27

29

31

33

0 1 2 3 4 5 6 7 8 9 10

Leaching agent concentration(%)

P2 O

5 %

Fig. 4.3.1 Effect of acid concentration on P2O5 content with liquid/solid ratio 7:1 for 55 min

at 55 oC

106

3

5

7

9

11

13

15

17

19

21

23

25

0 1 2 3 4 5 6 7 8 9 10

Leaching agent concentration (%)

CO

2 %

Fig. 4.3.2 Effect of acid concentration on CO2 content with L/S ratio 7:1 for 55 min at 55oC

107

Table 4.4.1 Effect of liquid/solid ratio on P2O5 and CO2 contents using 8% acid concentration

for 55 min at 55oC

L/S Ratio P2O5% CO2%

4 27.9 9.0

5 29.1 7.4

6 30.9 5.5

7 31.5 5.1

8 32.5 5.1

9 32.7 4.9

108

27

28

29

30

31

32

33

34

35

3 4 5 6 7 8 9 10

Liquid/solid ratiox

P2 O

5 %

Fig.4.4.1 Effect of liquid/solid ratio on P2O5 contents using acid concentration 8% for 55 min

at 55oC.

109

4

5

6

7

8

9

10

3 4 5 6 7 8 9 10

Liquid/solid ratio

CO

2%

Fig.4.4.2 Effect of liquid/solid ratio on CO2 contents using acid concentration 8% for 55 min

at 55oC.

110

4.5 Leaching with Lactic acid

Lactic acid has also been used to see the effect of various parameters on the

selective leaching of low-grade phosphate rock with various reaction conditions such as acid

concentration, liquid/solid ratio, particle size, stirring speed and temperature. The succinic

acid can be used as a leaching agent as it dissolves the calcareous material in low-grade

phosphate rock and improves the P2O5 content up to an industrially acceptable level.

However, the use of this acid is limited by its solubility at lower temperatures. The succinic

acid does not appear to work well below about 37oC.

As mentioned before, the formic and acetic acids can be used for selective

leaching of calcareous material, but they may cause higher corrosion effects on the

equipment along with greater tendency of attack on the phosphate element as well.

Lactic acid, being relatively weak and having more organic character, may be

expected to show less corrosion effects on the equipment along with minimal risk of leaching

of the phosphate element itself as compared to formic and acetic acids. On the other hand,

lactic acid is relatively stronger than succinic acid and may be expected to efficiently

dissolve the calcareous material in low-grade phosphate rock. Therefore, the possibility of

lactic acid to selectively dissolve the calcareous material in low-grade phosphate rock has

also been studied.

4.5.1 Effect of Acid Concentration on the Conversion

In the case lactic acid, the selective dissolution of calcareous material in low-

grade phosphate rock is also expected to depend on the various parameters such as reaction

time, temperature, acid concentration, liquid/solid ratio, particle size and stirring speed of the

111

reaction mixture. A number of experiments have been carried out to see the effect of acid

concentration on the selective leaching of calcareous material in the rock as shown Table

4.5.1 and in Figs. 4.5.1-4.5.5. The experimental results show that at relatively higher acid

concentration, the leaching rate of carbonaceous material is not much higher, a situation that

may be attributed to the increase in the polarity of the acid O-H bond.

At relatively higher acid concentration the number of hydrogen ions decreases in

the dissolution medium as the water amount decreases more and more in the medium. The

results indicate that, in the case of selective leaching for calcareous material in the phosphate

rock, the acid concentration should not be increased more than the required level to avoid the

acid attack on phosphate element during the leaching process as already has been described

in the use of succinic acid.

4.5.2 Effect of Liquid-Solid Ratio on the Conversion

To see the effect of liquid/solid ratio on the dissolution process, a number of

experimental have been carried out as shown in Table 4.5.2.1 and Figs. 4.5.2.1-4.5.2.5. The

experimental results show that the degree of dissolution increases with an increase in the

liquid/solid ratio, a situation that indicates that the total number of acid ions is increased due

to the increase the volume of the medium at the same concentration of the acid. It has been

found in a separate experiment that with lower acid concentration, the value of liquid/solid

ratio considerably increases to achieve the required degree of leaching. However, higher

liquid/solid ratio is resulted in a relatively lengthy process including handling, filtration and

regeneration of the consumed acid.

112

4.5.3 Effect of Particle Size on the Conversion

A number of experimental have been carried out to see the effect of particle size

ratio on the dissolution of calcareous material as shown in Table 4.5.3.1 and Figs. 4.5.3.1-

4.5.3.5. The experimental results show that the efficiency of the leaching process increases

with a decrease in the particle size. This can be expected because the surface area per unit

weight of sample increases as the sample is reduced in its size.

On the other hand, the liberation of more calcareous material from the apatite

matrix can be expected as the degree of size reduction is increased more and more. The

efficiency of the leaching process may be increased by further size reduction and fine

grinding of the phosphate feed, but the problems related to handling, filtration and marketing

of fine phosphates, may not allow grinding to exceed a certain limit. On the other hand,

higher amount of energy required for grinding and size reduction may cause an extra cost to

the process.

4.5.4 Effect of Temperature on the Conversion

At various reaction temperatures, a number of experiments have been carried out

to see the effect of reaction time on the dissolution of calcareous material at the known

parameter conditions as shown Table 4.5.4.1. The leaching rate curves have been shown in

Figs. 4.5.4.1-4.5.4.5. The results indicate that, by increasing the temperature, the dissolution

rate of calcareous material in the phosphate rock increases with a decrease in reaction time

required to reach the equilibrium.

The experimental results also show that after about 45-50C the dissolution rate

of calcareous material in the sample is not much higher, a situation which indicates that

113

higher temperatures may tend to decrease the solubility of lactate along with the

contamination of CO2 gas with water and lactic acid vapors.

4.5.5 Effect of Reaction Time on P2O5 and CO2 Contents

Various experiments have been carried out to see the effect of reaction time on

P2O5 and CO2 contents depending on acid concentration, liquid/solid ratio, particle size and

temperature as shown in Table 4.5.5.1 and Figs. 4.5.5.1-4.5.5.2. The results show that lactic

acid may be used to reduce calcium carbonate in low-grade calcareous phosphate rock as it

increases the degree of beneficiation making it viable as a feed to an industrial acidulation

process. For example, at 45 min the P2O

5 content has been raised up to about 35% along with

corresponding reduction in CO2 content up to about 70% of the sample at a phosphate

recovery of more than 75%.

The summery of the optimal parameters regarding calcareous phosphate rock and lactic acid

are given below.

Optimum time required for completion the reaction= 45min

Optimum liquid/solid required for completion the reaction =7:1

Optimum acid concentration required for completion the reaction=8%

Optimum particle required for completion the reaction =0.1255mm

Optimum temperature required for completion the reaction =45-50 oC

Optimum rpm required for completion the reaction =350

114

Table 4.5.1 Effect of reaction time on the conversion XB% of calcareous material at

different acid concentrations

Time (min) C=5% C=6% C=7% C=8% C=9%

5 0.05 0.07 0.10 0.13 0.16

10 0.09 0.14 0.17 0.22 0.25

15 0.13 0.19 0.24 0.28 0.33

20 0.17 0.23 0.29 0.35 0.39

25 0.20 0.28 0.36 0.41 0.50

35 0.31 0.37 0.44 0.55 0.59

45 0.35 0.45 0.55 0.64 0.69

55 0.39 0.51 0.59 0.72 0.83

115

0

0.1

0.2

0.3

0.4

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction time (min)

L/S=7, D=0.1255mm, T=45 OC , SS=350 rpm

a e

xp

Fig. 4.5.1 Effect of reaction time on the conversion of calcareous material with 5% lactic acid

concentration

116

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

1

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction time (min)

L/S=7, D=0.1255mm, T=45oC, SS=350 rpm

a e

xp


concentration

117

0

0.2

0.4

0.6

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction Time (min) C8% L/S=7 T=45 OC SS=350rpm

ae

xp


concentration

118

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction time(min)

L/S=7, D=0.1255mm, T=45 OC SS=350rpm

ae

xp


concentration

119

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

0 5 10 15 20 25 30 35 40 45 50 55 60

Reactiom time(min)

L/S=7, D=0.1255mm, T=45OC SS=350rpm

ae

xp


concentration

120

Table 4.5.2.1 Effect of Reaction Time on the Conversion of calcareous material in the rock at

Various Liquid/Solid Ratio

Time (min) L/S=4 L/S=5 L/S=6 L/S=7 L/S=8

5 0.03 0.05 0.08 0.1 0.13

10 0.08 0.09 0.15 0.18 0.25

15 0.13 0.16 0.21 0.27 0.31

20 0.17 0.23 0.28 0.33 0.41

25 0.21 0.28 0.35 0.41 0.49

35 0.31 0.39 0.49 0.56 0.65

45 0.38 0.49 0.63 0.71 0.80

55 0.41 0.55 0.69 0.79 0.91

121

0

0.1

0.2

0.3

0.4

0.5

0.6

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction time(min)C=8%D=0.1255MM.T=450C,SS=350rpm

a e

xp.

Fig.4.5.2.1 Effect of reaction time on the conversion with liquid/solid ratio (L/S) 4:1

122

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction time(min)

C=8%, D=0.1255mm, T=450C.SS=350rpm

aex

p.

Fig. 4.5.2.2 Effect of reaction time on the conversion with liquid/solid ratio (L/S) 5:1

123

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction time (min )

C=8%, D=.1255mm, T=45oC, SS=350rpm

a e

xp


124

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

1

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction time(min)

C=8%, D=0.1255mm.T=450C SS=350rpm

ae

xp.


125

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

1

0 10 20 30 40 50 60

Reaction Time (min)

C=8%, D=0.1255mm.T=450C SS=350rpm

ae

xp.


126

Table 4.5.3.1 Effect of reaction time on the conversion using different size particles of the

rock

Time (min) 0.503 (mm) 0.356 (mm) 0.2515 (mm) 0.1775 (mm) 0.1255 (mm)

5 0.02 0.04 0.07 0.1 0.12

10 0.05 0.1 0.12 0.16 0.21

15 0.08 0.15 0.19 0.24 0.29

20 0.12 0.2 0.26 0.31 0.41

25 0.17 0.27 0.31 0.38 0.45

35 0.26 0.38 0.45 0.56 0.67

45 0.34 0.46 0.57 0.65 0.79

55 0.36 0.49 0.61 0.76 0.87

127

0

0.1

0.2

0.3

0.4

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction time(min)

C=8%, L/S=7, T=45OC, SS=350rpm

aexp

Fig. 4.5.3.1 Effect of reaction time on the conversion for particle size 0.503 mm

128

Fig.4.5.3.2 Effect of reaction time on the conversion for particle size 0.356 mm

0

0.1

0.2

0.3

0.4

0.5

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction time(min)

C=8%, L/S=7,T=45OC, SS=350rpm

aexp

129

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction Time(mm)

C=8%, L/S=7, T=45OC, SS=350rpm

ae

xp


130

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction tiime(min)

C=8%, L/S=7, T=45OC, SS=350rpm

ae

xp


131

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction time(min)

C=8%, L/S=7, T=45OC, SS=350rpm

ae

xp


132

Table 4.5.4 1 Effect of reaction time on the conversion (XB) of calcareous material in the

rock at various temperatures

Time (min) Tem. 25oC Tem. 35oC Tem. 45oC Tem. 55oC Tem. 65oC

5 0.03 0.09 0.13 0.15 0.19

10 0.06 0.13 0.19 0.25 0.35

15 0.09 0.18 0.29 0.35 0.49

20 0.13 0.22 0.36 0.46 0.65

25 0.15 0.27 0.43 0.57 0.82

35 0.19 0.36 0.61 0.79 0.99

45 0.22 0.45 0.77 0.98 -

55 0.27 0.55 0.93 - -

133

0

0.1

0.2

0.3

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction Time(min)C=8%L/S=7D=0.1255mm SS=350rpm

a e

xp.

Fig. 4.5.4.1 Effect of reaction time on the conversion at 25oC

134

0

0.1

0.2

0.3

0.4

0.5

0.6

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction Time(min)C==8%L/S=7D=0.1255mm SS=350rpm

a e

xp.


135

0

0.2

0.4

0.6

0.8

1

0 10 20 30 40 50 60

Reaction time(min)C=8%L/S=7D=0.1255mm SS=350rpm

a e

xp.


136

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

1

0 10 20 30 40 50 60

Reaction time(min)C=8%L/S=7D=0.1255mmSS=350rpm

a e

xp.


137


0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

1

0 10 20 30 40 50 60

Reaction time(min) C =8% D=,1255mm rpm= 350

a e

xp

138

Table 4.5.5.1 Effect of reaction time on P2O5 and CO2 contents

Time (min) P2O5% CO2%

0 22.39 20.13

5 25.41 16.81

10 26.69 15.48

15 28.23 13.71

20 30.24 11.49

25 31.49 7.95

35 33.46 7.95

45 35.27 5.96

55 37.49 3.52

139

22

24

26

28

30

32

34

36

38

40

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction time(min) C=8%, L/S=7, D=0.1255mm, T=45oC, SS=350rpm

P2 O

5 %

Fig. 4.5.5.1 Effect of reaction time on P2O5 contents

140

3

5

7

9

11

13

15

17

19

21

0 5 10 15 20 25 30 35 40 45 50 55 60

Reaction time(min)

C=8% L/S=7, D=0.1255mm, T=45oC, SS=350rpm

CO

2 %

Fig. 4.5.5.2 Effect of reaction time on CO2 contents

141

4.6 Kinetic Analysis

4.6.1 Leaching with succinic acid

Fluid-solid heterogeneous reactions are of a great interest for industrial

applications regarding a number of chemical and hydrometallurgical processes. A successful

and efficient performance of chemical reactors for these processes is generally based on

kinetic data. In order to determine the kinetic parameters and rate-controlling step for the

selective leaching of calcareous material in the phosphate rock using succinic acid, the

experimental data may be analyzed according to the heterogeneous reaction models

[162,163]. According to the model, the reaction between a fluid and a solid can be written as:

F(fluid)+bS(solid)Products (4.6.1.1)

The rate of reaction between calcareous material particle and the dissolution agent

may be controlled by one of the following steps:

Diffusion through the fluid film, diffusion through the ash/product layer or the chemical

reaction at the surface of the particle involved.

Let the time of completion of the dissolution process be k*, the fractional

conversion of calcareous material be and at any time t the integrated equations for liquid-

solid heterogeneous reactions are:

For film diffusion control,

11*kt (4.6.1.2)

For chemical reaction control,

3/111* kt (4.6.1.3)

and for ash or product layer diffusion control,

142

)1(2131* 3/2 kt (4.6.1.4)

The value of *k will depend on various reaction conditions according to the

kinetic models. For example, according to the chemical reaction controlled model, eq.

(4.6.1.3) *k is:

AsoB CbKRk /* (4.6.1.5)

where *k stands for the time required for complete leaching (min), B is the

molar density of the solid reactant (mol m-3), oR is the size of the solid particle (m), b is the

stoichiometric coefficient of the solid reactant, sk is the surface reaction rate constant (m

min-1) and AC is the leaching agent concentration (mol m-3). To determine the kinetic

parameters and rate-controlling step for selective leaching of calcareous material in the

phosphate rock sample, the experimental data have been analyzed on the basis of liquid-solid

heterogeneous reaction models.

The validity of the experimental data into the integral rate has been tested by

statistical and graphical methods. The kinetic analysis results for the dissolution of

calcareous material has been found to be promising with a chemically controlled reaction and

the integral rate expression is determined to follow the following kinetic model:

kt 3/1)1(1 (4.6.1.6)

The eq. (4.6.1.6) has been found to give the best straight lines in comparison with

other kinetic models tested. At various reaction temperatures, using a known size range of the

rock particles and the acid concentration with a specific liquid/solid ratio, the results have

been plotted between the tvsX B .)1(13/1

as shown in Figs. 4.6.1.1-4.6.1.5. Where XB=,

is the fraction of the reacted calcareous material in the rock sample.

143

0

0.01

0.02

0.03

0.04

0.05

0.06

0.07

0.08

0 20 40 60

Leaching reaction time (min)

1-(1

-XB)1/

3

Fig. 4.6.1.1 Time versus 1-(1-XB )1/3 at 40oC

144

0

0.02

0.04

0.06

0.08

0.1

0.12

0.14

0.16

0.18

0 20 40 60


1-(1

-XB)1/

3

Fig. 4.6.1.2 Time versus 1-(1-XB) 1/3 at 50 oC

145

0

0.05

0.1

0.15

0.2

0.25

0.3

0.35

0.4

0 20 40 60


1-(1

-XB)1/

3

Fig. 4.6.1.3 Time versus 1-(1-XB) 1/3 at 60oC

146

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0 20 40 60


1-(1

-XB)1/

3


147

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

1

0 20 40 60


1-(1

-XB)1/

3


148

The values of the apparent rate constants, k, have been determined from the slopes

of the straight lines in Figs 4.6.1.1-4.6.1.5. Using the Arrhenius law,

RTEaekk / (4.6.1.7)

A plot of ln k versus 1/T should result in a straight line with a slope of -Ea/RT and

an intercept of ln k* if the experimental data are fitted well by the Arrhenius equation. For

each value of the reaction temperature, the data have been plotted between ln k and 1/T as

shown in Table 4.6.1.1 and Fig. 4.6.1.6. From Fig. 4.6.1.6, the following values have been

calculated:

Ea = 64.92 kJ mol-1

k=1.47106 s-1

149

Table 4.6.1.1 The values of ln k and 1/T.

1/T (K) ln k

0.003195 -10.8198

0.003096 -9.97248

0.003 -9.1303

0.002915 -8.4844

0.002833 -8.04199

150

-12

-11

-10

-9

-8

-7

-6

0.0028 0.0029 0.003 0.0031 0.0032 0.0033

1/T (K)

ln k

(s-1

)

Fig. 4.6.1.6 Arrhenius plot of ln k vs 1/T

151

The value of activation energy shows that the process is a chemically controlled

reaction and is consistent with the values found in the similar research work of liquid-solid

reaction systems [148,164,150]. The activation energy indicates a relatively low value at

higher temperatures, a situation that may be attributed to the mixed chemically-diffusion

controlled behavior of the leaching process. However, at lower temperatures the leaching

mechanism of the low-grade phosphate rock appears to be chemically controlled process,

which agrees with the similar findings reported in literature [165].

The value of activation energy in the dissolution process has been characterized to

determine the controlling step. The activation energy of a diffusion controlled process has

been typically found from 4-12 kJ mol-1, while for a chemically controlled process it has

been reported usually greater than 40 kJmol-1 [155]. The experimental results show that the

dissolution of calcareous material in low grade phosphate rock with succinic acid is

controlled by chemical reaction and thus, Eq. (4.6.1.6) may be written as:

teX RTB

/92.6463/1 1047.1)1(1 (4.6.1.8)

To assess the applicability of the suggested kinetic model, the theoretical values

of the conversion (Xcalc) have been calculated using Eq. (4.6.1.8). To test agreement between

the experimental conversion and the values calculated from the suggested empirical equation,

the graph of Xexpl versus Xcalc has been plotted as shown in Fig. 4.6.1.7. It is has been

observed that the agreement between the experimental and calculated values is good as the

data in the scatter diagram show a positive trend to cluster around the regression line, which

indicates the existence of a good relationship between the variables.

152

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

1

0 0.2 0.4 0.6 0.8 1

Xexp

Xca

l

Fig. 4.6.1.7 Agreement between experimental and calculated conversion values

153

4.6.2 Leaching with lactic acid

To determine the values of kinetic parameters and rate-controlling step for

selective leaching of calcareous material in the phosphate rock sample, the experimental data

have been analyzed on the basis of fluid-solid heterogeneous reaction models as given in the

case of succinic acid. Using statistical and graphical methods, the validity of the

experimental data into the integral rate has been tested. The analysis results for the leaching

of calcareous material with lactic acid have been found to be consistent with a chemically

controlled reaction as well. Thus, the integral rate expression has been determined to follow

the following kinetic model:

kt 3/1)1(1 (4.6.1.6)

After determine the values of the reaction rate constant, k, the Arrhenius equation

may be used to calculate the activation energy and thus the kinetic model:

tek RTEa /3/1)1(1 (4.6.2.1)

To explain the selective dissolution of calcareous material in the rock, the above

kinetic model, Eq. 4.6.2.1 is not expected to hold good for different parameter conditions as

its applicability is limited by the specific values (acid concentration, liquid/solid ratio and

particle size) used in the leaching process. The reaction parameters can affect the rate of

dissolution of calcareous material. These parameters may be included into the above-

mentioned model. Using succinic acid, these parameters have not been included and the

selective leaching kinetic model is based on the Arrhenius equation only. However, to see the

effect of different parametric conditions on the kinetic model, a semi empirical model has

been developed by statistical multiple regression. To establish this model representing the

154

selective leaching process, it has been accepted that the reaction rate constant k may be given

as follows:

RTEcbao

aeDSLCkk /)/( (4.6.2.2)

Combining Eqs. (4.6.2.1) and (4.6.2.2) it follows that:

teDSLCk RTEcbao

a /3/1 )/()1(1 (4.6.2.3)

Where is the conversion fraction, C the acid concentration, L/S the liquid/solid

ratio, D the particle size, R, ko, a, b, c constants, T the reaction temperature, Ea the activation

energy, t the reaction time for the leaching process.

The stirring speed has been omitted from multiple regression analysis as the

results in Table 4.6.2.1 and Fig. 4.6.2.1 show that the leaching of calcareous material in the

rock increases slowly with an increase up to 350 rpm. Further increase in the stirring rate

does not show a considerable effect on the dissolution of calcareous material. This means

that the diffusion through the fluid film cannot be considered as a rate-controlling step.

155

Table 4.6.2.1 Effect of rpm on the conversion

rpm Conversion ()

200 0.47

250 0.54

300 0.62

350 0.67

400 0.68

450 0.69

500 0.70

156

0.4

0.45

0.5

0.55

0.6

0.65

0.7

0.75

200 280 360 440 520

rpm C=8%, L/S=7:1 D=.1255mm, T-45oC

aexp

Fig. 4.6.2.1 Effect of rpm on the conversion at specific reaction conditions

157

For statistical calculations by simultaneous multiple regression a computer

program has been used giving the following values for the constants:

ko=19.1

a=1.753

b=1.627

c= -0.737

Ea=42964.62 J mol-1

Such a value of activation energy, for the selective leaching of calcareous

material, shows that the process is a chemically controlled reaction, which is consistent with

the values found in the fluid-solid reaction systems [136,148,150]. The results show a change

in the leaching phenomena of calcareous material after about 318 K, a situation that may be

attributed to the mixed chemically-diffusion controlled behaviour of the reaction process.

However, at relatively low temperatures the selective dissolution of calcareous material

shows to be chemically controlled process, which is consistent with the similar findings

given in the literature [165]. Depending on the type and nature of the solid materials as well

as strength of the leaching agents, a number of calcite dissolution mechanism studies are

reported in literature [143-145,166-172]. At moderate pH values, most of the studies have

shown that the overall rate of leaching is controlled by a chemical change.

For the selective leaching of calcareous material in the rock with acetic acid

[144], the investigations show that, in the pH range from 2.37 to 4, a chemical change has

been found to be the rate-controlling step for the overall process. Regarding the strength of

the acid used in the selective dissolution, the findings of the present work are also supported

158

by the similar study [145], which claims that the calcareous material dissolution above about

pH 3.7 has been significantly controlled by the kinetics of the surface reaction.

On the other hand, the kinetic analysis regarding the production rate of calcite

leaching in the leach solution shows that the diffusion-controlled mechanism is not the

principal mechanism in the dissolution process. The mechanism of calcite dissolution for the

rate-determining slowest step may need further investigations; however, the results indicate

that the rate-controlling step of the overall process of the heterogeneous reaction system is a

chemical change. Thus, using the values determined by simultaneous multiple regression

analysis the Eq. (4.6.2.3) may be given as:

teDSLC RT/62.42954737.0627.1753.13/1 )/(1.19)1(1 (4.6.2.4)

To test the agreement between the experimental conversion and the values

calculated from the semi empirical model, the graph of exp versus cal has been plotted as

shown in Fig. 4.6.2.2.

159

0

0.2

0.4

0.6

0.8

1

0 0.1 0.2 0.3 0.4 0.5 0.6 0.7 0.8 0.9 1

exp

ca

l

Fig. 4.6.2.2 Agreement between experimental and calculated conversion values

160

The results show that the agreement between the experimental and calculated

values is very good with a correlation coefficient of 0.983 and standard deviation of

0.04156. On the other hand, the data in the scatter diagram indicates a positive trend to

cluster around the regression line, a situation that can be attributed to the existence of a good

relationship between the variables involved. The value of the standard deviation of regression

or standard error of estimate of cal on exp indicates that the degree of scatter of the observed

values about the regression line is not higher. By the statistical analysis with 157 numbers of

experimental data, a relative mean square of errors of 0.1296 has been calculated by Eq.

4.6.2.5:

2/1

12

2exp

)(

1

N

i cal

cal

NER

(4.6.2.5)

The value of relative mean square of errors shows that, in probabilistic models,

such a value of the random errors is not significant. It is interesting to test the applicability of

the semi-empirical model by using the various adjustable parameter conditions for the

dissolution of calcite in low-grade phosphate rock. For 70% dissolution of calcareous

material in the rock, using the various parameter values in the semi-empirical model, the

reaction times have been calculated as shown in Table 4.6.2.2.

The comparison between case 1 and 2 as shown in Table 4.6.2.2 shows that when

the acid concentration decreases from 8 to 7%, the reaction period increases from 46.94 to

59.32 min. Although the increase in the reaction period is about 1.26-fold, however, it shows

that 26.4% more amount of the material can be handled. Regarding the cases 1 and 3 in Table

4.6.2.2, when the acid concentration decreases from 9 to 8%, the reaction period increases

161

only about 8.76 min, which indicates a relatively poor efficiency of the dissolution process as

compared to the above case.

The comparison between case 1 and 4 as shown in Table 4.6.2.2, when the liquid-

solid ratio increases from 6:1 to 7:1, the decrease in the reaction time has been found to be

about 13.38 min. While, regarding the cases 1 and 5 in Table 4.6.2.2, when the liquid-solid

ratio increases from 7:1 to 8:1, the reaction time decreases only about 9.17 min, which shows

that the value of liquid-solid ratio of 7:1 is more beneficial than the other liquid-solid values

used in the leaching process.

162

Table 4.6.2.2 The values of the selected parameters and the reaction periods calculated from

Eq. (4.6.2.4) for these parameters.

Case

Studied

Conc.

(% v/v)

L/S ratio

(cm3 g-1)

Dia.

(mm)

Tem.

(K)

Dissolution

(%)

Time

(min)

1 8 7 0.1255 318 70 46.94

2 7 7 0.1255 318 70 59.32

3 9 7 0.1255 318 70 38.18

4 8 6 0.1255 318 70 60.32

5 8 8 0.1255 318 70 37.77

6 7 8 0.1255 318 70 47.74

7 8 7 0.1775 318 70 60.61

8 8 7 0.1255 328 70 28.61

9 8 7 0.1255 308 70 79.55

163

On the other hand, regarding the cases 1 and 6 in Table 4.6.2.2, the interchange of

concentration and liquid-solid ratio does not indicates any considerable effect on the reaction

time, a situation that shows that these two parameters equally compete for the selective

leaching of calcareous material in the rock. For the cases 1 and 7 in Table 4.6.2.2, when the

particle size decreases from 0.1775 to 0.1255 mm, the reaction period also decreases from

60.61 to 46.94 min, a situation that may be attributed to the fact that the size reduction causes

an easier liberation of the calcareous material from the apatite matrix.

Comparing the two cases 1 and 8 in Table 4.6.2.2, when the reaction temperature

increases from 318 to 328 K, the reaction period decreases from 46.94 to 28.61 min with an

increase in the amount of handled material about 39%. On the other hand, for the cases 1 and

9 in Table 4.6.2.2 when the reaction temperature increases from 308 to 318 K, the reaction

time decreases from 79.55 to 46.94 min with an increase in the amount of handled material

about 41%. The comparison between these two cases shows that to study with 318 K reaction

temperature is more beneficial than the other temperatures. As a result, 8% v/v acid

concentration, 7:1 liquid-solid ratio cm3 g-1, 0.1255 mm particle size, 350 min-1 stirring speed

and 318 K reaction temperature may be recommended as the parameter values, because these

values show the optimum selective dissolution rate (case 1).

To increase the rate of dissolution of calcareous material in the rock using lactic

acid, the results show that the reaction period may be decreased, by increasing the acid

concentration and/or liquid solid ratio as long as the considerable leaching of phosphate

element is not appeared in the leach slurry. Further reduction in particle size may be expected

to improve the efficiency of the dissolution process depending on the type and nature of low-

grade calcareous phosphate rock. Based on the required degree of selective dissolution, the

164

study may be extended to a pilot scale work to beneficiate low-grade calcareous phosphate

rock. Depending on the type and nature of phosphate rock, the design parameters may need

some further investigations at a commercial scale; however, the method may be of value to

others confronted with the need to use typical indigenous phosphate rocks of various

countries in the world.

165

CHAPTER 5

5.1 RECOVERY OF SPENT ACID AND ECONOMY OF PROCESS

The economy of the leaching process may depend on the price of leaching agent

and the cost of its recovery. Succinic acid may be recovered from the calcium succinate

solution by a number of routs as reported in literature [141].

It is estimated that succinic acid did not lost during performing the practical as

long as the succinate solution is recycled to the leaching process for dilution of fresh succinic

acid. However, after the leaching process succinic acid can be recovered from the spent

liquid using any strong acid, providing that an insoluble salt is formed in order to separate the

recovered succinic acid by simple methods such as filtration. For the recovery of succinic

acid, some methods may also be used such as ion exchange, hydrofluoric acid, fluorosilicic

acid.

5.2 Routes for the Acid Recovery

A number of routs, which may be applied to recover the spent acid, are discussed below:

(a) Use of Ion Exchange Resins

Succinic acid from the succinate solution may be recovered by using ion

exchange methods. Treatment of the bed to recover succinic acid from succinate solution

may be written as:

Ca(C4H4O4)+2HRC4H6O4+CaR2 (5.2.1)

Under suitable conditions of the bed, an ion exchange process may recover more than 95%

succinic acid from the succinate solution.

166

Hydrochloric acid may be used to regenerate the bed as shown in equation (5.2.2):

CaR2+2HClCaCl

2+2HR (5.2.2)

(b) Using Hydrofluoric Acid

Succinic acid may be regenerated by another route using succinate solution with

the hydrofluoric acid produced during the manufacture of phosphoric acid or super

phosphates from fluoroapatite.

Ca(C4H4O4)+2HFC4H6O4+CaF2 (5.2.3)

(c) Use of Fluorosilicic Acid

The fluorosilicic acid separated from the phosphoric acid plant or its fresh stock

may be used to treat the succinate solution, which produces succinic acid and insoluble

calcium fluorosilicate salt.

Ca(C4H4O4)+H2SiF6C4H6O4+CaSiF

6 (5.2.4)

Calcium fluorosilicate salt is insoluble in cold water and thus the spent acid may be separated

for recycle.

(d) Use of Sulphuric Acid

Succinic acid may be recovered by the reaction of calcium succinate solution with

any strong acid, provided that an insoluble salt is formed in order to separate the recovered

succinic acid by filtration. The use of sulphuric acid has been found to be the most attractive

method for recovering succinic acid from calcium succinate solution. It potentially converts

the calcium succinate into succinic acid and calcium sulphate. The reaction between

sulphuric acid and calcium succinate can be written as:

Ca(C4H4O4)+H2SO

4+2H

2O+C4H6O4 +CaSO

4.2H

2O (5.2.5)

167

Calcium sulphate produced during the reaction can be separated by filtration.

Stoichiometrically, 98 parts of H2SO4 are needed per 44 parts of CO2 removed from the

sample. The spent organic acid may be regenerated for further use and recycling process.

However, the regeneration needs further studies and details regarding process and designing

parameters. Therefore its extended study is suggested for future. Schematic diagram

suggested for the recovery of the spent acid has been shown in Fig. 5.2.1.

168

Fig. 5.2.1 Schematic diagram suggested for the recovery of spent acid.

169

5.3 Economy of the Process

The economy of the leaching process depends on a number of parameters such as

availability of the rock, demand and application of the product, price of leaching agent and

the cost of its recovery from the spent liquor. Succinic acid may be recovered from the spent

liquor, for example, using sulphuric acid as described in the previous section. Loss of

succinic acid in a closed circuit plant may be expected low. The recovered succinic acid as

well as the washing solution may be recycled to the leaching process decreasing the amount

of the fresh acid. This means that a small amount of the fresh acid may be needed to make-up

the leaching agent concentration in the process.

The suggested leaching process for the beneficiation of low-grade phosphate rock

may eliminate/reduce some of the problems, which are faced during manufacture of

phosphoric acid, a number of phosphate rock based chemicals and phosphate fertilizers. The

dilute solution of the acid is expected to have minimal corrosion effects on the equipments

along with the selective leaching of calcareous material without any considerable amount of

dissolution of phosphate element itself. To increase the overall economy of the suggested

process, it can directly be attached with an acidulation plant in order to eliminate/reduce the

downstream processing steps like drying, handling and packing of the product after the

leaching and beneficiation of the rock.

The filtrate solution (spent liquor) containing succinate salt may be evaporated to

produce crystalline calcium succinate. The calcium succinate may be used directly for other

industrial purposes. Another alternative is to market the calcium succinate and CO2 gas

produced during the leaching process. The by-products, CO2 and calcium sulphate may be

used or sold to other industries to meet the cost of commercial succinic acid and its recovery

170

by sulphuric acid. It is expected that the marketing of the CO2 gas product for other

industries may pay back some of the operating costs of the leaching process.

171

Conclusions

The analysis of the sample indicates that loss on ignition (LOI), insoluble residue

(AIR), P2O5 and CO2 contents depend on the particle size as well as the type and nature of

calcareous phosphate rock. The dissolution results show that succinic acid can be used as

leaching agent in order to eliminate/reduce the calcareous material in low-grade phosphate

rock as it improves the degree of beneficiation up to marketable and industrially acceptable

level. Using -0.175+0.125 mm size fraction, an acid concentration of 8% with liquid/solid

ratio of 7:1 is found to be optimum for the selective dissolution of calcareous material in the

rock.

Analysis of the kinetic data by various kinetic models shows a chemically

controlled process with mixed chemical-diffusion control relatively at higher temperatures.

The activation energy for the leaching process has been found to be 64.92 kJ mol-l, which

agrees with a chemically controlled reaction. For the selective dissolution of calcareous

material in the rock, the applicability of the suggested kinetic model, i.e.,

te RT/92.6463/1 1047.1)1(1 has been tested by the agreement between the experimental

and calculated conversion values. The analysis results show that the agreement between

experimental and calculated values of the conversion is good, however, it has been found that

succinic acid is not a successful leaching agent below about 37 oC due to its limited

solubility.

172

Lactic acid can also be used to selectively leach the calcareous material present in

the phosphate rock as it improves the P2O5 content of the sample making it viable as a feed to

an acidulation plant. For the leaching rate of calcareous material, the process parameters such

as acid concentration, liquid-solid ratio, particle size and reaction temperature have been

optimized using the stirring speed of 350 rpm.

Analysis of the kinetic data by different kinetic models shows that the selective

dissolution of calcareous material in the phosphate rock with lactic acid is also a chemically

controlled process. The value of activation energy has been found to be 42.96 kJmol-1, which

is consistent with a chemically controlled reaction. Regarding the values of activation energy,

the comparison between succinic and lactic acid shows that the rate of dissolution of

calcareous material is relatively higher with lactic acid, which can be expected as the lactic

acid is stronger than the succinic acid.

The applicability of the suggested semi-empirical kinetic model, i.e.,

teDSLC RT/62.42954737.0627.1753.13/1 )/(1.19)1(1 has been tested and the results

show that the calcareous fraction reacted can be estimated with a relative mean square of

errors of 0.1296. The agreement between experimental and calculated values of the

conversion also indicates a good relationship between the variables involved. The analysis

results show that the parameter values for the optimum selective leaching rate are C = 8%

v/v, L/S = 7 cm3g-1, T = 318 K, SS = 350 min-1 and D = 0.1255 mm.

173

The economy of the leaching process mainly depends on the price of the leaching

agent and the cost of its recovery. Both of the leaching agents (succinic and lactic acid) may

be recovered from their soluble salts (calcium succinate and calcium lactate) solutions by a

number of routs. However, sulphuric acid is found potentially most important for recovering

the spent acid from the leach slurry.

Regeneration of organic leaching agents is crucial for further application and

recycling process. Therefore it is suggested to extend study in the future, regarding

regeneration process conditions as well as the process design parameters.

174

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Appendix A

List of Symbols

MCP Monocalcium Phosphate

DCP Dicalcium Phosphate

TCP Tricalcium Phosphate

BPL Bone Phosphate of Lime

TPL Total Phosphate of Lime

NP Nitrophos

DAP Diammonium Phosphate

MAP Monoammonium Phosphate

SSP Single Superphosphate, or

NSP Normal Superphosphate (also known as superphosphate)

TSP Triple Superphosphate (also known as concentrated superphosphate)

AIR Acid Insoluble Residue

LOI Loss on Ignition

=XB dissolved fraction of Ca++ (–)

cxp experimental dissolved fraction of Ca++ (–)

cal predicted fraction of Ca++ (–)

C acid concentration (% v/v)

L/S liquid/solid ratio (g-1cm3)

D average diameter of particle (mm)

t reaction time (min)

T reaction temperature (K)

192

Ea activation energy (Jmol-1)

SS stirring speed (min-1)

N number of experimental data

193

Appendix B

Published Work

REACTION KINETICS AND MASS TRANSFER...

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