SCHOOL OF ENGINEERING AND TECHNOLOGY DERPARTMENT OF CHEMICAL AND PROCESS SYSTEMS ENGINEERING DESIGN OF A 1

SCHOOL OF ENGINEERING AND TECHNOLOGY
DERPARTMENT OF CHEMICAL AND PROCESS SYSTEMS ENGINEERING
DESIGN OF A 1.6 TPD POTASSIUM DICHROMATE MANUFACTURING PROCESS BY ELECTROCHEMICAL OXIDATIVE DECOMPOSITION OF CHROMITE SLAG AND LOW-GRADE CHROMITE IN A CONCENTRATED ALKALINE SYSTEM.

BY
NYARADZO NDUNA H140122J
SUPERVISED
Mr L KADZUNGURA
THIS CAPSTONE DESIGN WAS SUBMITTED TO THE HARARE INSTITUTE OF TECHNOLOGY IN PARTIAL FULFILMENT OF THE BACHELOR OF TECHNOLOGY (HONOURS) DEGREE IN CHEMICAL AND PROCESS SYSTEMS ENGINEERING

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2018
COPYRIGHTSAll rights reserved. No part of this Capstone project may be reproduced, stored in any retrieval system, or transmitted in any form or by any means, electronic, mechanical, photocopying, recording or otherwise from scholarly purpose, without the prior written permission of the author or of Harare Institute of Technology on behalf of the author.

DECLARATIONI, Nduna Nyaradzo hereby declare that this work has not been accepted in substance for any degree and is not being concurrently submitted in candidature for any degree. The work will not be submitted to another University in the awarding of a degree.
Student’s Signature: ………………………………. Date ………………………….
(Nduna Nyaradzo)
Supervisor’s Signature: ……………………………. Date ………………………….
(Mr. L Kadzungura)

DEDICATIONThis project is dedicated to my mother for her faithful support and encouragement during the design of this project.

ACKNOWLEDGEMENTSI would like to thank God, first and foremost for the abounding mercies during the course of this capstone design.

Secondly, I would like to thank my supervisor Mr. L Kadzungura for his guidance throughout the project. Thirdly I would like to thank my mother for her encouragement and support. Special appreciation goes to all my friends and relatives for the encouragement and technical support.
Lastly my sincere gratitude goes to all lecturers in the Chemical and Process Systems Engineering Department and Metallurgical team at ZIMASCO and The Department of Metallurgy in Harare without whom it would have been impossible for me to acquire the knowledge and samples to undertake this project.

ABSTRACTChromite processing is an important metallurgical industry in Zimbabwe, done primarily by ZIMASCO, where the chromite ore is processed to get Ferro-chrome. Zimbabwe produces over 50 000 tonnes of slag annually and already has dumps of over 2 million tonnes in the Kwekwe area. The aim of this project is to carry out a technical and economic feasibility study of setting up a potassium dichromate manufacturing plant from electric arc furnace slag and low-grade chromite ore by electrochemical oxidative decomposition. The design of major equipment and construction of a prototype was done. Experimental work carried out showed that at 60 wt. % alkali, 21.0125g of Cr2O3 gin 100g of a mixture of ore and slag, gave a recovery of 89.67%. The experimental data was used to do a mass and energy balance, showing an output of 1.6 tonnes per day of potassium dichromate. A HAZOP analysis was done for the major equipment to ensure equipment operability and safety. The project requires a total capital investment of US$ 1 259 823.85, total manufacturing cost is US$ 691 860.035 and the selling price of the product is US$ 2000/tonne. The projected sales are considered viable with a 3.02 years payback period and a rate of return of 33% along with a breakeven point of US$ 332 519.

Keywords: electrochemical oxidative decomposition, electric arc furnace slag, potassium dichromate.

TABLE OF CONTENTS TOC o “1-3” h z u COPYRIGHTS PAGEREF _Toc515841774 h iDECLARATION PAGEREF _Toc515841775 h iiDEDICATION PAGEREF _Toc515841776 h iiiACKNOWLEDGEMENTS PAGEREF _Toc515841777 h ivABSTRACT PAGEREF _Toc515841778 h vTABLE OF CONTENTS PAGEREF _Toc515841779 h viLIST OF FIGURES PAGEREF _Toc515841780 h xiLIST OF TABLES PAGEREF _Toc515841781 h xiiLIST OF ABBREVIATIONS PAGEREF _Toc515841782 h xivCHAPTER ONE: INTRODUCTION PAGEREF _Toc515841783 h 11.0Background PAGEREF _Toc515841784 h 11.1 Introduction PAGEREF _Toc515841785 h 21.2 Problem statement PAGEREF _Toc515841786 h 21.3 Justification PAGEREF _Toc515841787 h 31.4 Aim PAGEREF _Toc515841788 h 31.5 Objectives PAGEREF _Toc515841789 h 31.6 Hypothesis PAGEREF _Toc515841790 h 41.7 Scope of study PAGEREF _Toc515841791 h 41.8 Market analysis and demand PAGEREF _Toc515841792 h 4CHAPTER TWO: LITERATURE REVIEW PAGEREF _Toc515841793 h 52.0 Introduction PAGEREF _Toc515841794 h 52.1 Potassium dichromate PAGEREF _Toc515841795 h 52.1.1 Properties PAGEREF _Toc515841796 h 62.1.2 Uses of potassium dichromate PAGEREF _Toc515841797 h 62.1.3 Safety PAGEREF _Toc515841798 h 72.1.4 Manufacturing processes PAGEREF _Toc515841799 h 72.2 Geology and mineralogy of chromite PAGEREF _Toc515841800 h 92.3 Proposed manufacturing method PAGEREF _Toc515841801 h 112.3.1 Advantages of proposed process PAGEREF _Toc515841802 h 112.3.2 Proposed method PAGEREF _Toc515841803 h 12CHAPTER THREE: RESEARCH METHODOLOGY PAGEREF _Toc515841804 h 213.0 Introduction PAGEREF _Toc515841805 h 213.1 Sources of data PAGEREF _Toc515841806 h 213.1.1 Primary sources PAGEREF _Toc515841807 h 213.1.2 Secondary sources PAGEREF _Toc515841808 h 213.2 Data collection and presentation PAGEREF _Toc515841809 h 213.3 Experimental design PAGEREF _Toc515841810 h 223.3.1 Experiment 1 PAGEREF _Toc515841811 h 223.3.2 Experimental procedure to get potassium dichromate PAGEREF _Toc515841812 h 223.3.3 Experiment 2 PAGEREF _Toc515841813 h 243.3.4 Experiment 3 PAGEREF _Toc515841814 h 253.3.5 Experiment 4 PAGEREF _Toc515841815 h 253.3.6 Experiment 5 PAGEREF _Toc515841816 h 263.3.7 Experiment 6 PAGEREF _Toc515841817 h 26CHAPTER FOUR: EXPERIMENTAL RESULTS AND ANALYSIS PAGEREF _Toc515841818 h 284.0 Introduction PAGEREF _Toc515841819 h 284.1 Experimental results and analysis PAGEREF _Toc515841820 h 284.1.1 Experiment 1 PAGEREF _Toc515841821 h 284.1.2 Experiment 2 PAGEREF _Toc515841822 h 294.1.3 Experiment 3 PAGEREF _Toc515841823 h 304.1.4 Experiment 4 PAGEREF _Toc515841824 h 314.1.5 Experiment 5 PAGEREF _Toc515841825 h 324.1.6 Experiment 6 PAGEREF _Toc515841826 h 334.2 Conclusion PAGEREF _Toc515841827 h 33CHAPTER FIVE: PROCESS DESIGN PAGEREF _Toc515841828 h 355.0 Introduction PAGEREF _Toc515841829 h 355.1 Process description PAGEREF _Toc515841830 h 355.1.1 Crushing PAGEREF _Toc515841831 h 355.1.2 Milling and classification PAGEREF _Toc515841832 h 355.1.3 Electrolysis PAGEREF _Toc515841833 h 365.1.4 Dilution PAGEREF _Toc515841834 h 365.1.5 Filtration PAGEREF _Toc515841835 h 375.1.6 Leaching PAGEREF _Toc515841836 h 375.1.7 Filtration PAGEREF _Toc515841837 h 375.1.8 Evaporation PAGEREF _Toc515841838 h 375.1.9 Drying PAGEREF _Toc515841839 h 375.2 Block flow diagram PAGEREF _Toc515841840 h 385.4 Material balances PAGEREF _Toc515841841 h 415.4.1 Background data PAGEREF _Toc515841842 h 415.4.2 Mass balance around jaw crusher PAGEREF _Toc515841843 h 425.4.3 Mass balance around ball mill and cyclone 1 PAGEREF _Toc515841844 h 435.4.4 Mass balance around ball mill and cyclone 2 PAGEREF _Toc515841845 h 455.4.5 Mass balance around electrolytic reactor PAGEREF _Toc515841846 h 475.4.6 Mass balance around dilution tank PAGEREF _Toc515841847 h 485.4.7 Mass balance around pressure filter PAGEREF _Toc515841848 h 495.4.8 Mass balance around leaching tank PAGEREF _Toc515841849 h 505.4.9 Mass balance around pressure filter PAGEREF _Toc515841850 h 515.4.10 Mass balance around evaporator PAGEREF _Toc515841851 h 525.4.11 Mass balance around dryer PAGEREF _Toc515841852 h 535.4.12 Conclusion PAGEREF _Toc515841853 h 545.5 Energy balances PAGEREF _Toc515841854 h 555.5.1 Energy balance around jaw crusher PAGEREF _Toc515841855 h 555.5.2 Energy balance around ball mill 1 PAGEREF _Toc515841856 h 555.5.3 Energy balance around ball mill 2 PAGEREF _Toc515841857 h 565.5.4 Energy balance around reactor PAGEREF _Toc515841858 h 565.5.5 Energy balance around forced circulation evaporator PAGEREF _Toc515841859 h 585.5.6 Energy balance around dryer PAGEREF _Toc515841860 h 60CHAPTER SIX: EQUIPMENT DESIGN PAGEREF _Toc515841861 h 646.0 Introduction PAGEREF _Toc515841862 h 646.1 Forced circulation evaporator PAGEREF _Toc515841863 h 646.1.1 Introduction PAGEREF _Toc515841864 h 646.1.2 Selection criteria for heat exchanger PAGEREF _Toc515841865 h 646.1.3 Types of heat exchangers considered PAGEREF _Toc515841866 h 646.1.4 Flow arrangement PAGEREF _Toc515841867 h 656.1.5 Assumptions PAGEREF _Toc515841868 h 656.1.6 Calculation of Heat exchanger design PAGEREF _Toc515841869 h 666.1.7 Chemical engineering design PAGEREF _Toc515841870 h 666.1.8 Thermal design PAGEREF _Toc515841871 h 676.1.9 Mechanical design PAGEREF _Toc515841872 h 756.2 Ball mill PAGEREF _Toc515841873 h 806.2.1 Parameters design PAGEREF _Toc515841874 h 816.2.2 Mechanical design PAGEREF _Toc515841875 h 84CHAPTER SEVEN: PROCESS CONTROL AND HAZOP ANALYSIS PAGEREF _Toc515841876 h 887.0 Introduction PAGEREF _Toc515841877 h 887.1 Alarms, safety, trips and interlocks PAGEREF _Toc515841878 h 887.2 Process control around heat exchanger PAGEREF _Toc515841879 h 887.2.1 Flow rate control PAGEREF _Toc515841880 h 897.2.2 Temperature control PAGEREF _Toc515841881 h 897.3 Process control around ball mill PAGEREF _Toc515841882 h 907.3.1 Level control PAGEREF _Toc515841883 h 907.3.2 Temperature control PAGEREF _Toc515841884 h 907.4 Hazop PAGEREF _Toc515841885 h 917.4.1 Hazop analysis of heat exchanger PAGEREF _Toc515841886 h 937.4.2 Hazop analysis of ball mill PAGEREF _Toc515841887 h 94CHAPTER EIGHT: SITE SELECTION AND PLANT LAYOUT PAGEREF _Toc515841888 h 958.0 Introduction PAGEREF _Toc515841889 h 958.1 Site selection PAGEREF _Toc515841890 h 958.1.1 Factors affecting site selection PAGEREF _Toc515841891 h 958.1.2 Suitable sites PAGEREF _Toc515841892 h 958.1.3 Decision matrix PAGEREF _Toc515841893 h 968.2 Plant layout PAGEREF _Toc515841894 h 968.2.1 Maneuvering space PAGEREF _Toc515841895 h 978.2.2 Plant layout PAGEREF _Toc515841896 h 99CHAPTER NINE: ENVIRONMENTAL IMPACT ASSESMENT PAGEREF _Toc515841897 h 1009.0 Introduction PAGEREF _Toc515841898 h 1009.1 Environmental legislation PAGEREF _Toc515841899 h 1009.2 Environmental impacts and mitigation PAGEREF _Toc515841900 h 1019.2.1 Site clearance and preparation PAGEREF _Toc515841901 h 1019.2.2 Construction impacts PAGEREF _Toc515841902 h 1019.2.3 Operation impacts PAGEREF _Toc515841903 h 1039.2.4 Decommissioning impacts PAGEREF _Toc515841904 h 104CHAPTER TEN: ECONOMIC ANALYSIS PAGEREF _Toc515841905 h 10610.0 Introduction PAGEREF _Toc515841906 h 10610.1 Estimation of capital investment PAGEREF _Toc515841907 h 10610.1.1 Direct costs PAGEREF _Toc515841908 h 10610.1.2 Indirect cost PAGEREF _Toc515841909 h 11010.1.3 Working capital PAGEREF _Toc515841910 h 11110.1.4 Total capital investment PAGEREF _Toc515841911 h 11110.2 Estimation of total production cost PAGEREF _Toc515841912 h 11110.2.1 Fixed charges PAGEREF _Toc515841913 h 11110.2.2 Variable costs PAGEREF _Toc515841914 h 11210.2.3 Plant overheads PAGEREF _Toc515841915 h 11310.3 Production cost analysis PAGEREF _Toc515841916 h 11310.4 Profitability analysis PAGEREF _Toc515841917 h 11310.4.1 Profit margin PAGEREF _Toc515841918 h 11310.4.2 Payback period PAGEREF _Toc515841919 h 11410.4.3 Rate of return PAGEREF _Toc515841920 h 11410.4.4 Break-even point PAGEREF _Toc515841921 h 11510.4.5 Net present value PAGEREF _Toc515841922 h 115CHAPTER ELEVEN: CONCLUSION AND RECOMMENDATIONS PAGEREF _Toc515841923 h 11711.0 Conclusion PAGEREF _Toc515841924 h 11711.1 Recommendations PAGEREF _Toc515841925 h 117APPENDIX PAGEREF _Toc515841926 h 119REFERENCES PAGEREF _Toc515841927 h 122
LIST OF FIGURES
Fig 2.1 Chemical structure of potassium dichromate5
Fig 2.2 Potassium dichromate appearance6
Fig 2.3 Ferrochrome10
Fig 2.4 Single toggle jaw crusher13
Fig 2.5 Tumbling mill15
Fig 2.6 Grate and overflow mill discharge16
Fig 2.7 Hydrocyclone17
Fig 2.8 Pressure filtration18
Fig 2.9 Forced circulation evaporator20
Fig 4.1 Conversion at different alkali concentrations29
Fig 4.2Conversion at varying temperatures30
Fig 4.3 Conversion at different values of current31
Fig 4.4 Effect of alkali to ore mass ratio32
Fig 4.5 Effect of residence time on recovery33
Fig 5.1 Process flow diagram39
Fig 5.2 Piping and instrumentation diagram62
Fig 7.1 Process control around heat exchanger 90
Fig 7.2 Process control around ball mill 91
Fig 8.1 Plant layout99
LIST OF TABLES
Table 2.1 Properties of potassium dichromate6
Table 2.2 Composition of slag and low-grade chromite11
Table 4.1 Composition of low grade chromite ore28
Table 4.2 Composition of slag28
Table 4.3 Effect of varying alkali concentration29
Table 4.4 Effect of varying temperature30
Table 4.5 Effect of varying current31
Table 4.6 Effect of alkali to ore mass ratio32
Table 4.7 Effect of residence time33
Table 5.1 KOH solubility in water36
Table 5.2 Parameters and data for steam59
Table 5.3 Operating parameters59
Table 6.1 Experimental data for design 67
Table 6.2 Ball mill design parameters 84
Table 6.3 Ball mill selected parameters 84
Table 7.1 Hazop key 92
Table 7.2 Hazop analysis of heat exchanger 93
Table 7.3 Hazop analysis of ball mill 94
Table 8.1 Decision matrix for site selection 96
Table 10.1 Bill of quantities of a ball mill106
Table 10.2 Bill of quantities forced circulation evaporator107
Table 10.3 Purchased equipment cost108
Table 10.4 Estimation of physical plant cost109
Table 10.5 Estimation of fixed capital109
Table 10.6 Direct cost110
Table 10.7 Indirect cost110
Table 10.8 Plant overheads113
Table 10.9 Net present value cash flow116
LIST OF ABBREVIATIONSCOPRChromite Ore Processing Residue
LPOLiquid Phase Oxidation
SMSSub Molten Salt System
ROSReactive Oxygen Species
ORROxygen Reduction Reaction
EAFElectric Arc Furnace
ROSReactive oxygen species
DAQData acquisition device
CHAPTER ONE: INTRODUCTIONBackground
Pollutions from the mining industry are no doubt a threat to Zimbabwe’s environment. Contaminating our water bodies, air and harming our soils it is no surprise that most industries are looking into more environmentally friendly ways of running processing plants. This increased interest in the environment has also seen a spike in research for more sustainable ways that are more resource efficient and have optimal processes. In the traditional chrome processing industry, hazardous chromium containing residues (Sun et al., 2009a) are produced which have both biological and medical negative effects on nature. An example is calcium chromate which is carcinogenic.

Production of ferrochrome, an alloy of iron and chromium containing 50-70% chromium by weight by carbothermic reduction of chromite produces a large amount of slag. During the processing, a part of these metals gets oxidized and these oxides end up in the slag or dust. Apart from the adverse economic factor accompanying the metal losses due to oxidation, these slags pose a serious environmental threat as these metals finally end up in slag dumps only to be leached by acid rains over years, however they are also a rich source of chromate with can be oxidised to a dichromate.

Thus, it is highly convenient to utilize the slag as a secondary source of raw material. In this context, a process that can be used both for slags containing these metals as well as for low grade ores would be attractive as the high-grade ore availability is shrinking on a global perspective.

Traditionally, the metal recovering processes of metals can be classified into the following categories: pyro metallurgy, hydrometallurgy, solvent extraction and ion-exchange. By applying mineral processing technologies, such as crushing, grinding, magnetic separation, eddy current separation, flotation and so on, leaching or roasting, it is possible to recover metals such as Fe, Cr, Cu, Al, Pb, Zn, Co, Ni, Nb, Ta, Au, and Ag etc2.

EAF stainless steel slags contain high content of Cr. The final EAF slag from high alloy steel processing in Sweden contains, on the average, 2-3 weight percent Cr. Consequently, it is necessary to treat the slags prior to landfilling or other applications.

1.1 IntroductionThe chrome industry is an important mining industry providing both basic and intermediate products for the metallurgical, refractory and engineering industries. It is however a major source of pollution. The common way of processing chromite ore is the use of a kiln or rotary furnace after the addition of limestone and dolomite at around 1373K. This is characterized by low resource utilisation and energy efficiency, producing hazardous residues (NRC Committee., et al 1974). Due to poor mass transfer efficiency common with the roasting process, up to 75% of chromium is recovered in the stages that follow afterwards, which are leaching stages. This highlights a wastage in resources. Large quantities of chromite ore processing residue around 2.5-3 tonnes per tonne of chromate produced (Wang et al., 2013), chromite dust and waste gases are discharged.

Health effects of chromium include skin rashes, ulcers, respiratory problems kidney and liver damage to name a few. This is if has been ingested or inhaled. Its effects on the environment are that it affects the soil’s natural compositions. It also damage the gills of fish in water sources.

With an increasing interest in sustainable ways of production in the nation, the need to process chromite slag optimally and efficiently is no exception. This paper proposes such a solution by introducing an electrochemical field to enhance chromite decomposition in a potassium hydroxide (KOH) sub-molten salt medium (SMS), under relatively moderate reaction conditions (Patil et al., 2006; Wang et al., 2010). With over 2 million tonnes of chromium containing slag lying in dumps in Kwekwe’ s ZIMASCO refinery and over 50000 tonnes produced annually, the health hazard it poses is clear as these metals end up being leached into the soil by acid rain. The need for value addition and environmental protection opens the way for an economically feasible way to make use of this slag, in this case to manufacture potassium dichromate
1.2 Problem statement
Although there has been a number of marked improvements to the conventional roasting method in chromate production such as roasting with less or no calcareous additives(Anthony et al.,2006), the problems associated with high temperature roasting technology which include low overall resource utilisation, high energy consumption and environmental pollution due to formation of COPR still remain unsolved. This supports the initiative to design a suitable more efficient dichromate production plant which will not only reduce overall cost and pollution, but also marginally increase chromate recovery.

1.3 Justification
It is in line with the Zimbabwe Agenda for Sustainable Socio-Economic Transformation (ZIM ASSET). Value addition and Beneficiation.

Helps in tackling pollution caused by COPR since it is now the raw material for the production of potassium dichromate
With an abundant source of chrome, availability of slag and low grade chromite ore is enough to sustain the plant
Creation of jobs
Economic – process will use electric arc furnace slag and low grade dolomite which is cheaper and in abundance. Calcination of chromite at 12000C after with dolomite and limestone, the conventional way of making potassium dichromate is expensive due to the large energy consumption. This coupled with the low utilisation of resources as 75% of chromate is recovered and environmental hazards posed by COPR it is hard to justify this process on an economic basis (Antony et al.,2006 ).

In contrast, the proposed liquid phase oxidation (LPO) technology integrated with electrolysis, has capabilities of occurring at around 1500-3500C. Research has proved it beneficial with low energy efficiencies together with marginally higher conversions at greater than 88%, using a solution of KOH-KNO3-H2O for cases where electrolysis was not involved (Zhang et al., 2009). This project proposes not using non nitrate LPO methods because nitrates are expensive and lead to complicated nitrate separation and recycling procedures. The chromite ore will be oxidised by air or oxygen in pure molten or sub molten alkaline (Hundley et al., 1985; Sun et al., 2007a; Xu et al., 2006).

Socially – value addition and creation of jobs.

Environmentally – helps to tackle pollution on the environment and lessens animals and humans’ exposure to COPR like calcium carbonate that can be carcinogenic.

1.4 Aim
The project aims to develop and design an effective potassium chromate production plant by electrochemical oxidative decomposition of chromite tails in a sub molten KOH solution.
1.5 Objectives
Design a process for manufacture of potassium dichromate from chromite ore tails after processing
Design a suitable electrolytic reactor and forced circulation evaporator that can achieve high recoveries of potassium dichromate
Ascertain the optimum conditions of operation
Find suitable ways to treat the waste after recovery of potassium dichromate
Provide a sustainable source of potassium dichromate in Zimbabwe
1.6 Hypothesis
H0 –The use of electrochemical oxidative decomposition of chromite tails can produce sustainable potassium dichromate.

H1 –The process is not feasible
1.7 Scope of study
The project will critically review the impact of electrochemical oxidative decomposition on chromite tails, with the integration of a sub molten salt system, as a sustainable source of potassium dichromate. The case study will be ZIMASCO tails. Study will cover;
Process design
Mechanical design- electrolytic reactor and leaching tank
Process control
Prototype development
Site selection and plant layout
1.8 Market analysis and demand
Potassium dichromate in todays’ market is valued at US$ 3100/ tonne. Demand has recently marginally improved owing to the growing demand in Asia( Transparency Market Research 2015). Demand is mostly seen in tannery (using around 175 kg a day to treat around 200 hides) industry, detergent manufacture , medicinal and photography.

CHAPTER TWO: LITERATURE REVIEW2.0 Introduction
Zimbabwe is the fourth largest producer of chromite in the world. Mining around 7 300 000 metric tonnes in 2002(Mineral commodity summary 2006). Chromite, also called iron chromium oxide belonging to the spinel group is by far the most industrially important mineral for the production of metallic chrome which is used as an alloying ingredient in stainless steel tools.

ZIMASCO in Zimbabwe is in the business of processing chromite to produce high carbon ferrochrome. The process produces large amounts of slag annually that have chromium containing compounds. It is the purpose of this project to make use of these slag.

2.1 Potassium dichromate
Potassium dichromate, K2Cr2O7, is a common inorganic chemical reagent, most commonly used as an oxidizing agent in various laboratory and industrial applications. As with all hexavalent chromium compounds, it is acutely and chronically harmful to health. It is a crystalline ionic solid with a very bright, red-orange colour. Potassium dichromate occurs naturally as the rare mineral lopezite. It has only been reported as vug fillings in the nitrate deposits of the Atacama Desert of Chile and in the Bushveld igneous complex of South Africa ( Anthony et al.,2003)
Fig 2.1 Chemical structure of potassium dichromate

Fig 2.2 Potassium dichromate appearance

2.1.1 Properties
Table 2.1 Properties of potassium dichromate
Variable Property
Chemical formula K2Cr2O7
Molar mass 294.18g/mole
Appearance Red-orange
Odour Odourless
Density 2.676g/cm3
Melting point 3980C
Boiling point 5000C
Solubility 4.9 g/100ml of water at 00C
13g/100ml of water at 200C
102g/100ml of water at 1000C
Insoluble in alcohol and acetone
2.1.2 Uses of potassium dichromate
Staining and tanning of leather
Medicine (external antiseptic and vertinary medication)
Bleach
Oxidising agent
Depolariser for dry cells
Used as a dye
Photography and photographic screen printing
Construction as it increases density and texture of cement
2.1.3 Safety
In 2005–06, potassium dichromate was the 11th-most-prevalent allergen in medical patch tests (Warshaw et al., 2006)
Potassium dichromate is one of the most common causes of chromium dermatitis, chromium is highly likely to induce sensitization leading to dermatitis, especially of the hand and fore-arms, which is chronic and difficult to treat. Toxicological studies have further illustrated its highly toxic nature. With rabbits and rodents, concentrations as low as 14 mg/kg have shown a 50% fatality rate amongst test groups ( Sigma et al.,2011)Aquatic organisms are especially vulnerable if exposed, and hence responsible disposal according local environmental regulations is advised.

As with other Cr (VI) compounds, potassium dichromate is carcinogenic and should be handled with gloves and appropriate health and safety protection. The compound is also corrosive and exposure may produce severe eye damage or blindness. Human exposure further encompasses impaired fertility, heritable genetic damage and harm to unborn children.

2.1.4 Manufacturing processes
Potassium dichromate has largely been supplanted by the cheaper sodium dichromate but is still used whenever its advantage of being non hygroscopic is important, for example, in the match, firework, film, and photographic industries.

However there three main methods used in the manufacture of potassium dichromate, which are;
Reaction between potassium chloride and sodium dichromate
KCL + NaCr2O7?K2Cr2O7 + 2NaCl
Reaction of potassium chromate (K2CrO4) and an acid
The reaction of potassium chromate and an acid give a dichromate salt.

Roasting of chrome ore
In the traditional chromate production process, chromite is calcinated with sodium carbonate at 1200? in a rotary kiln with the addition of dolomite and limestone, with low-utilization ef?ciency of resources and energy, producing a hazardous residue containing about 5% chromium, e.g. CaCrO4, which is harmful to the environment. Reactions 1–3 occur with the generation of CaO from limestone and dolomite under the condition of traditional process. Once the CaCrO4 is airborne, it behaves as a further threat to all living beings because of its irritative, corrosive, and carcinogenic characters.

2CaO+Cr2O3+ 32O2?2CaCrO4….12CaO+MgO.Cr2O3+ 32O2?2CaCrO4+MgO….22Cao+FeO.Cr2O3+ 74O2?2CaCrO4+12Fe2O3….3To optimize the chromate production process, many alternate methods, such as acid leaching(Li et al., 2011; Vardar et al.,1994), carbon reduction at high temperature (Chakraborty et al.,2010), and liquid-phase oxidation (LPO) (Kashiwas et al., 1974a), have been proposed, among which the LPO technology is regarded as the most promising alternative. In early reports regarding LPO, chromite ore was treated with molten NaOH–NaNO3 medium or highly concentrated KOH–KNO3 aqueous solution under oxidative conditions using air, oxygen or nitrate as the oxidant (Kashiwas et al., 1974b; Sun et al., 2009; Zhang et al., 2010). The main differences between the roasting methods are, the oxidation reactions in these LPO approaches are pseudo-homogeneous or homogeneous, and the mass transfer efficiencies are much higher. Fundamental studies suggest that in the molten medium or highly concentrated alkaline solutions (namely the sub-molten salt medium, SMS), chromite could be oxidized by oxygen, air, nitrate, or the reactive oxygen species (ROS), which could be obtained via the pyrolysis
of hydroxyl or nitrate ions and the oxygen reduction reaction (ORR) process (Chong et al., 2008; Jin et al., 2010; Sun et al.,2009). While ROS could cause mineral lattice distortion or catalytic oxidation of metal sub oxides in the ore, the above characteristics are what contribute to the high reaction efficiency of LPO processes in comparison with the traditional roasting method. For example, in the NaOH–NaNO3 binary molten salt medium, the chromite was efficiently oxidized by sodium nitrate with a chromium yield reaching 95% after reacting for 2 hours at 350? with a liquid-to-ore ratio of 30:1. However, these nitrate-involving LPO approaches are not economic due to the consumption of expensive nitrate, and the complicated nitrate separation/recycling procedure. In this regard, none-nitrate LPO methods have been proposed. In these new LPO processes, the chromite was oxidized by air or oxygen in pure
molten or sub-molten alkaline medium (Hundley et al., 1985; Sun et al., 2007a; Xu et al., 2006; Zheng et al., 2006). Compared with nitrate- involving LPO methods, these new LPO approaches can achieve approximately the same chromium yield with lower cost. Particularly, in the KOH sub-molten salt medium (75 wt. % KOH aqueous solution), the chromite can be efficiently decomposed at 300?.

2.2 Geology and mineralogy of chromite
Chromium ore, or chromite, occurs exclusively in rocks formed by the intrusion and solidification of molten lava or magma which is very rich in the heavy, iron containing minerals such as pyroxenes and olivines. Within these rocks, often referred to as ultramafic igneous rocks, chromium occurs as a chromium spinel, a highly complex mineral made up, in its basic form, of magnesium as MgO and aluminium as Al2O3. However, the magnesium can be substituted in varying proportions by divalent iron, and the aluminium can be substituted, also in varying proportions, by trivalent chromium and trivalent iron. Thus the chromium
spinel may be represented as: (Fe,Mg)O.(Cr,Fe,Al)2O3. Large variations in the total and relative amounts of Cr and Fe in the lattice occur in different deposits. These affect the ore grade not only in terms of the Cr2O3 content but also in the Cr:Fe ratio which determines the chromium content of the ferrochromium produced. The variations also affect the reducibility (relative ease of reduction) of the ore. For example, increasing amounts of magnesium compared with iron in the divalent site will make the spinel more difficult to reduce. Conversely, increasing amounts of iron in the trivalent site, replacing aluminium, will increase the reducibility of the spinel.

The greater the index, the more refractory, or less reducible, the ore. Chromium is the most abundant of the Group V1A family of elements and at an average concentration of nearly 400ppm in the earth’s crust it is the 13th most common element. However, as with all minerals or elements, economic deposits occur only where it has been concentrated in nature. The chromium spinel is a heavy mineral and it concentrates through gravity separation from most of the other molten material in the magma during crystallisation from the cooling magma. Commercial chromite deposits are found mainly in two forms: stratiform seams in basin-like intrusions, often multiple seams through repeated igneous injections, and the more irregular
podiform or lenticular deposits.

Fig 2.3 Ferrochrome
The best known example of a stratiform deposit is the Bushveld Igneous Complex of South Africa. This complex contains most of the world’s chromite reserves. The Great Dyke of Zimbabwe, traversing nearly the length of the country, is very similar and has been linked to the Bushveld in geological history. These two features are well-known also for their important and very large commercial deposits of the platinum-group metals. Other stratiform deposits occur in Madagascar and in the Orissa district of India.Stratiform deposits are generally very large complexes. They can be more than 5,000 metres thick and cover thousands of square kilometres. For example, the largest, the Bushveld, covers an area of 12000 square kilometres.

The podiform deposits are relatively small in comparison and may be shaped as pods, lenses, slabs or other irregular shapes. Many have been extensively altered to serpentine and they are often faulted. They are generally richer in chromium than the stratiform deposits and have higher Cr: Fe ratios. Ore reserves in Kazakhstan are of the podiform type. Podiform ores were
originally highly sought after, especially those from the deposits in Zimbabwe, as the best source of metallurgical grade chromite for high-carbon ferrochromium. These ores also tend to be massive (hard lumpy) ores, as opposed to the softer, more friable ores from the stratiform deposits, and this makes for better electric smelting operation.
There is a third type of chromite deposit but of very limited commercial significance. These are the eluvial deposits that have been formed by weathering of chromite-bearing rock and release of the chromite spinels with subsequent gravity concentration by flowing water. Chromium may also be concentrated in high-iron lateritic deposits containing nickel and there have been attempts to smelt these to produce a chromium-nickel pig iron for subsequent use in the stainless steel industry.

2.3 Proposed manufacturing method
Recently, a novel KOH leaching process according to the principles of green chemistry, in which the recovery of chromium from chromite ore can be raised to above 88% and the residue amount reduced to 0.5 ton per ton product when compared with 2.5 ton in traditional processes, was proposed ( Sun et al., 2009). A demonstration plant was been built in the Henan province of China, which has approached zero emission of chromium residue. The process has been named a sub molten salt method. Attracting interest from related research groups and industries worldwide.

This paper proposes the same technology whilst introducing an electrochemical field to enhance chromite decomposition process. This is based on the successful application of acidic slurry electrolysis (Linge et al., 1995; Wang et al., 2010) and the feasibility of electrochemical oxidation of Cr2O3 in alkaline media, confirmed by previous studies (Jin et al., 2012; Zanello and Raspi et al., 1997).
To be able to get an estimate of the expected composition in EAF slag the Uddeholm tooling and Vietnamese chromite (Xinlei et al., 2010) is shown below;
Table 2.2 Composition of slag and low grade chromite ore
Stages CaO % MgO % Al2O3 % Cr2O3 %
After tapping 41.4 7.91 3.29 3.25
Original 0.12 8.88 11.94 41.40
2.3.1 Advantages of proposed process
Less energy requirements
Justifiable slot current density of 750 A/m2 with the activation energy of decomposition of chromite at 55.63KJ/mole at temperatures of 230?-370? based on the feasibility of electrochemical oxidation of Cr2O3 in alkaline media, confirmed by previous studies (Jin et al., 2012; Zanello and Raspi et al., 1997).

Cheaper since it incorporate the use of non-nitrate LPO methods, chromite ore will be oxidised by air or oxygen in pure molten or sub molten alkaline.

The oxidation reactions in these LPO approaches are pseudo-homogeneous or homogeneous, and the mass transfer efficiencies are much higher compared to roasting.

2.3.2 Proposed method 2.3.1.1 Electrolysis
Electrolysis is a process of breaking down a compound by electricity. An electric current is the flow of charged particles. In solids, substances that conduct electricity are called Conductors. These are mostly metals and graphite. This is because metals and graphite contain free electrons in their structures to carry the charge. The solids which do not conduct electricity are called Insulators.

The electrolysis cell is a battery each pole connected to an electrode and both electrodes are dipped in the liquid to be electrolysed.

The electrode connected to the positive pole is called the anode.

The electrode connected to the negative pole is called the cathode.

There are two types of electrodes, active electrodes and inert electrodes. Active electrodes take place in the process its self. Inert electrodes are just there to conduct the current without interfering.

Inert electrodes can be either graphite or platinum but graphite is more widely used because it’s cheaper. Inert electrodes are always used in electrolysis; active ones are used in electroplating.

Electrolysis separates an ionic compound back to the elements that form it. For example by electrolysis we can obtain sodium and chlorine from sodium chloride.

When the current is turned on, the negative ion in the electrolyte gets attracted to the positive electrode because they are oppositely charged. When this happens, the negative ion loses the electrons it gained from the positive ion during bond formation and becomes an atom. The electrons lost are transferred through the wire in the outer circuit from the anode to the cathode. At the same time, the positive ion from the electrolyte is attracted to the cathode, where it gains the electrons lost by the negative ion and becomes an atom too.

In ionic compounds the positive ion is a metal and it is collected at the cathode. And the negative ion is a non-metal and collected at the anode.

The electrons are transferred from the anode to the cathode through the wires.

The electrolyte is an ionic compound either in its molten or aqueous form. Ionic compounds conduct electricity only when they are in these forms because they contain free mobile ions which can carry the current but they don’t in solid form.

In this case the idea is the use of a stainless steel cathode and anode, thus they work as inert electrodes.

The major anodic reaction for the proposed reaction is;
2FeO.Cr2O3+2OH-+12O2?Fe2O3+CrO72-+H2O2.3.1.2 Crushing
Primary crushers are heavy duty machinery which are used to reduce the size of material, ore in this case, to manageable sizes for the secondary crushers. Types of primary crushers are;
Jaw crusher.

Gyratory crusher.

For the purpose of this project a single toggle jaw crusher was chosen.

Fig 2.4 Single toggle jaw crusher
Jaw crushers have two plates that open and close and open referred to as jaws. One jaw is pivoted and the other one fixed. Material after being fed is nipped and released to fall further into the crushing chamber.

The swing jaw which is suspended on the eccentric shaft moves towards the fixed jaw under the action of the toggle plate. As the eccentric rotates, the swing jaw also moves vertically. This motion allows the rock to be pushed through crushing chamber.

Advantages of a jaw crusher
Lighter.

Since the crushing rate is below 900t/hr( Will’s Mineral Processing Technology et al.2006) it is more economic for Mimosa to use a jaw crusher.

Less expensive to maintain.

2.3.1.3 Milling and classification
Grinding
The second stage of comminution after crushing is grinding which is the final stage of comminution. Particle size reduction is achieved by a combination of abrasion and impact. Grinding is performed in rotating cylindrical vessels known as tumbling mills. The mills contain loose grinding media which are large, hard and heavy relative to the ore particles but small in relation to the mill volume. When the mill is rotated the mixture of the medium, ore and water (mill charge) is intimately mixed. Comminution is by one of three mechanisms ie impact/chipping, compression, and abrasion. This process is usually performed wet although dry process is possible.

Advantage of wet milling;
Lower power consumption.

Higher capacity per unit mill volume.
Wet screening or classification is possible to produce a closed size product.

Environmentally acceptable since dust is eliminated.

Handling and transportation is made simple through the use of pipes and pumps.

Aids longitudinal flow through the mill.

Facilitates addition of chemical reagents to the pulp when their actions on freshly broken surfaces are desire
There are a variety of mill types, for different process materials and resources. Some of the type of mills include;
Autogenous mills – these achieve grinding through the action of the particles on one another. They are low capital cost and are capable of handling wet sticky material, have low manpower requirements and minimal grinding media expense.

Semi autogenous mills- These use a combination of both mill feed and steel balls as the grinding media.

Rod mills- these use steel rods as grinding media.

Ball mills- these use steel or cast iron balls up to 125mm in diameter to effect grinding. The balls used are:
High carbon- high manganese steel with alloying elements
Chilled cast iron
Forged low carbon steel
Stainless steel and cast nickel alloy.

Chromium steel balls
Balls have greater surface area per unit weight as compared to rods, so are used for fine grinding. They can be trunion overflow discharge or grate discharge. Ball mills are normally operated in closed circuit with classifiers for improved operating efficiency.

Fig 2.5 Tumbling mill
There are three basic types of grinding mills ie ball mill, rod mill and autogenous mill. A mill is generally a horizontal cylindrical shell with renewable wearing liners and a charge of grinding medium. The cylinder is supported so as to rotate on its axis on hollow trunions attached to the end walls. Feed material is introduced to the mill continuously through one end and the ground product leaves via the other end although some designs have ports on the cylinder periphery.

The final stage of comminution is done in the ball mill. Steel balls can be used as grinding media. Balls have greater surface area per unit mass. The length of the cylindrical shell is 1 to 1.5 times the diameter. If the ratio goes up to 3 – 5 then the mill is a tube mill. Tube mills are usually used in dry operation to grind materials like gypsum, cement and phosphate. Often pebbles are used as grinding media.
The system of discharge is such that pulp flows through the opening of the grate and is lifted to the trunion opening and this leads to a relatively coarser product.

Steel balls are used as grinding media. Balls have greater surface area per unit mass. The length of the cylindrical shell is 1 to 1.5 times the diameter. If the ratio goes up to 3 – 5 then the mill is a tube mill. Tube mills are usually used in dry operation to grind materials like gypsum, cement and phosphate. Often pebbles are used as grinding media. There are two types of ball mills namely the grate and overflow discharge mill.

The grate discharge has a system of discharge such that pulp flows through the opening of the grate and is lifted to the trunion opening and this leads to a relatively coarser product. The overflow discharge or trunion overflow leads to a finer product relative to grate discharge and energy consumption is 15% less for the same mill size.

Fig 2.6 Grate discharge and overflow discharge
Classification
Classification is a method of separating mixtures of minerals into two or more products on the basis of the velocity with which the grains fall through a fluid medium. CITATION Hei93 l 12297 (Heiskanen, 1993).

When a solid particle falls freely in a vacuum, it is subject to constant acceleration and its velocity increases indefinitely, being independent of size and density. Thus a lump of lead and a feather fall at exactly the same rate.

In a viscous medium, such as air or water, there is resistance to this movement and the value increases with velocity. When equilibrium is attained between the gravitational and fluid resistances forces, the body reaches its terminal velocity and thereafter falls at a uniform rate. The nature of the resistance depends on the velocity of the descent. At low velocities motion is smooth because the layer of fluid in contact with the body moves with it, while the fluid a short distance away is motionless. Between these two positions is a zone of intense shear in the fluid all around the descending particle. Effectively all resistance to motion is due to the shear forces or viscosity of the fluid and is hence called viscous resistance. At high velocities the main resistance is due to the displacement of fluid by the body, and viscous resistance is relatively small; this is known as turbulent resistance.

Fig 2.7 Hydro-cyclone
This is a continuously operating classifying device that utilizes centrifugal force to accelerate the settling rate of particles. It is one of the most important devices in the minerals industry, its main use in mineral processing being as a classifier, which has proved extremely efficient at fine separations.

2.3.1.4 Filtration
Filtration is a process of separating solids from liquid by means of a porous medium which retains the solid but allows the liquid to pass. This process follows after thickening.

Types of FiltersCake filters are most frequently used in mineral processing, where the recovery of large amounts of solids from fairly concentrated slurries is the main requirement. Cake filters may be:
Pressure filters.

Vacuum filters.

Pressure FiltersSeparating solids from a liquid is done by means of driving the pulp at high pressure against a porous material (filter cloth).The solids are captured on the surface of the porous material, and the liquid flows through the material. The solids build up on the medium or cloth to form a cake. As the cake builds up, it acts as a filtration medium in itself, and so the flow rate through the medium decreases. The liquid that passes through the cake and the cloth is termed filtrate. A pressure system is used to create the force necessary to push the filtrate through the cake. Sometimes flocculants are used to aid filtration, normally low molecular weight flocculants, as the higher molecular weight flocculants, used in thickening, tend to hold water and form large particles. A pressure differential across a filter medium plays a major role in determining the rate and efficiency of filtration.

Fig 2.8 Pressure filtration
Pressure differential: The difference between the pressure measured at the driving force side and the pressure measured at the filtrate side
Factors Affecting the Rate of FiltrationThe pressure drop from the feed to the far side of the filter medium. This is achieved in pressure filters by applying a positive pressure at the feed end and in vacuum filters by applying a vacuum to the far side of the medium, the feed side being at atmospheric pressure.

Surface Area The rate of filtration increases with the area of the filtering surface.

Slimes tend to blind the filter medium (the pores of the filter cloth) and this result in a filter cake with high moisture content.

Slurry Density A filter feed density of 1.5 to 1.6 is considered to be adequate for filtration processes. A low feed density to the filter results in a high cake moisture content, at the same cycle times.

Flocculants and coagulants agglomerate the ore particles to form larger particles which are easy to filter. Too much of the reagents cause the cake to become sticky and this upsets the filtration rate.

Filtration rate tend to decrease as the slurry viscosity increases.

2.3.1.5 Forced circulation evaporator
In the evaporation process, concentration of a product is accomplished by boiling out a solvent, generally water. The recovered end product should have an optimum solids content consistent with desired product quality and operating economics. It is a unit operation that is used extensively in processing foods, chemicals, pharmaceuticals, fruit juices, dairy products, paper and pulp, and both malt and grain beverages. Also it is a unit operation which, with the possible exception of distillation, is the most energy intensive.

While the design criteria for evaporators are the same regardless of the industry involved, two questions always exist: is this equipment best suited for the duty, and is the equipment arranged for the most efficient and economical use? As a result, many types of evaporators and many variations in processing techniques have been developed to take into account different product characteristics and operating .parameters.

The forced circulation evaporator was developed for processing liquors which are susceptible to scaling or crystallizing. Liquid is circulated at a high rate through the heat exchanger, boiling being prevented within the unit by virtue of a hydrostatic head maintained above the top tube plate. As the liquid enters the separator where the absolute pressure is slightly less than in the tube bundle, the liquid flashes to form a vapour.

Fig 2.9 Forced circulation evaporator
CHAPTER THREE: RESEARCH METHODOLOGY
3.0 Introduction
In order to achieve the best results and predictions before the actual implementation of the proposed process, various research techniques and instruments have to be applied. This chapter highlights the methods of data collection to be done.

3.1 Sources of data
The gathering of relevant data is the first step in conducting any study and described below are the major sources of data where information to be used in the project will be obtained from.

3.1.1 Primary sources
Surveys and interviews
Case studies
Experiments
3.1.2 Secondary sources
These are sources in which the researcher does not directly link or interact with the author of the information, but uses information stored or written by the author and are as follows:
Internet
Text books
E-books
3.2 Data collection and presentation
Data for use in this project was collected from credible sources to aid the in solving the problem. To satisfy the dictates of the problem, the relevant information for the project was gathered from a number of sources and these are textbooks, journals, records, internet and interviews
The design is made up of both qualitative and quantitative data. In this design, all qualitative data was presented in text format using the recommended styles. Quantitative data refers to that data which constitutes of figures. The following data presentation tools were used to represent the quantitative data, process flow diagram, Material balance, energy balance, process instrumentation and control model fabrication and engineering drawing software.

3.3 Experimental design
3.3.1 Experiment 1TITLE: Mineral analysis of chromite ore samples and slag respectively
AIM:To analyse the percentage composition of Cr2O3, FeO, Al2O3 and MgO in chromite ore and slag respectively.

REAGENTS AND APPARATUS
Laboratory mill
Laboratory sieves of sizes 20µm, 38µm, 45µm, 75µm, and 106µm
XRF (x-ray fluorescence) spectrometer
Laboratory crusher
PROCEDURE
Using a laboratory jaw crusher, chromite ore was crushed and then milled.

Separately the EAF slag was also milled.

Using sieves separately for both samples, a 20g sample each of less than 75µm and less was sent for analysis using an XRF spectrometer.

Results were collected, for further analysis and use during the calculation of percentage conversion
3.3.2 Experimental procedure to get potassium dichromateREAGENTS AND APPARATUS
Electrolytic reactor
Compressed air
Rectifier
Stainless steel electrodes
Magnetic stirrer
Water pump
Temperature controller
DC power supply
Stainless steel reactor
Water bath/heater/burner
KOH
Deionised water
Potassium carbonate
Crucibles
Gloves
Hot air oven
Weighing balance
Beakers
Test tubes
Filter paper
Samples of chromite ore and slag
PROCEDURE
A predetermined amount of potassium hydroxide was added to electrolytic cell along with deionised water.

Heating was done to a desired temperature whilst stirring at 700rpm until homogeneity was achieved.

The chromite ore and slag of 95% passing 75µm was added along with a constant supply of pressurized air at 1L/min, with 5g of potassium dichromate as catalyst.

The power supply was then switched on to 2.5Volts
Water was supplied to the system at 30min time intervals or as needed to compensate the water loses encountered due to electrolysis and evaporation.

The reaction was left to occur for 4hrs whilst collecting 2g of slurry at every 60 min interval.

The 2g sample was mixed in 80ml deionised water at 25? for 5min.

The sample was filtered to obtain a filtrate sample and a tails sample. Both samples were weighed and their masses recorded.

100ml of water at 120? was then added to tails sample for 10min whilst stirring. This was to allow the leaching out of potassium dichromate from tails.

Sample was filtered using filter paper and both filtrate and tails were weighed. Results were then recorded.

The remaining filtrate was placed in a crucible and placed in a hot air oven at 110? for 24hours.

Mass of crystallised potassium dichromate was determined.

The percentage conversion was then established using the equation below;
% conversion=(1-CrrCro)×100Where; Crr=concentration of chromium in residueCro=concentration of chromium in ore and slag3.3.3 Experiment 2
TITLE:Effect of alkali concentration
AIM:To investigate the effect of different alkali concentrations on potassium dichromate recovery.

REAGENTS AND APPARATUS
Electrolytic reactor
KOH at different concentrations of 50 wt. %, 60 wt. % and 70 wt. %
Beakers
Filter paper
Crucibles
Hot air oven
Gloves
Samples of chromite ore and slag
Weighing balance
PROCEDURE
A 100g sample of chromite ore and slag was prepared at a ratio of 3:1 respectively.

500g of alkaline solution were prepared at different concentrations. To obtain optimal reaction conditions, the different concentrations were then examined under standard conditions.

A slot current density of 750 A/m2, alkali to ore mass ratio of 5:1, pressurised air flow at 1L/min, stirring speed of 700 rpm, reaction temperature of 160? and a slag and chromite particle size of less than 200 mesh was used.

Steps from experimental procedure to get potassium dichromate were followed from 1-13.

For representative results, the experiment was repeated thrice for each respective concentration of alkali.

3.3.4 Experiment 3TITLE:Effect of temperature
AIM:To determine the optimal reaction temperature whilst investigating the effect of different temperatures on potassium dichromate recovery.

REAGENTS AND APPARATUS
Temperature controller
Same as apparatus for experimental procedure to get potassium dichromate
HCL
PROCEDURE
Electrodes were cleaned using 1+1 hydrochloric acid and deionised water.

100g of chromite ore and slag was prepared at a ratio of 3:1, 500g of alkaline at 60 wt. %, SCD 750A/m2, air flow at 100L/min, a stirring speed of 700 rpm and a chromite and slag particle size of less than 200 mesh was used.

To optimise reaction conditions, temperatures of 100?, 130?, 160? and 170? were used.

Steps from experimental procedure to get potassium dichromate were followed, whilst heating at the desired temperature.

For each temperature, the experiment was done three times.

3.3.5 Experiment 4
TITLE:Effect of slot current density
AIM:To investigate the effect of slot current density to potassium dichromate recovery.

REAGENTS AND APPARATUS
HCL
Same as experimental procedure to get potassium dichromate
PROCEDURE
Electrodes were cleaned using 1+1 hydrochloric acid and deionised water.

Preparation of 100g of chromite ore and slag was done at a ratio of 3:1, 500g of alkaline solution at 60 wt. %, air flow at 1L/min, stirring speed of 700 rpm, temperature of 150? and an ore and slag particle size of less than 200 mesh was used.

To optimise reaction conditions, SCD values of 200A/m2, 400A/m2, 600A/m2 and 800A/m2 were used separately.

Steps from experimental procedure to get potassium dichromate were repeated using a selected SCD.

For each SCD, the experiment was repeated thrice.

3.3.6 Experiment 5TITLE:Effect of alkali to ore mass ratio
AIM:To investigate the effect of different alkali to ore mass ratios on the percentage of potassium dichromate recovery.

REAGENTS AND APPARATUS
HCL
Same as experimental procedure to get potassium dichromate
PROCEDURE
Electrodes were cleaned using 1+1 hydrochloric acid and deionised water.

Preparation of 100g of chromite ore and slag was done at a ratio of 3:1,speed of 700rpm SCD of 750A/m2, air flow of 1L/min, temperature of 150?, alkali at 60 wt. % and a chromite and slag particle size of less than 200 mesh.

To optimise reaction conditions, alkali to ore mass ratios of 4:1, 5:1, 6:1 and 7:1 were used separately.

Steps from experimental procedure to get potassium dichromate were followed whilst using different amounts of alkali for each ratio.

Experiments were done 3 times for each alkali to ore ratio.

3.3.7 Experiment 6TITLE:Effect of residence time
AIM:To investigate the effect of residence time on percentage recovery of K2Cr2O7
REAGENTS AND APPARATUS
HCL
Same as experimental procedure to get potassium dichromate
PROCEDURE
Electrodes were cleaned using 1+1 hydrochloric acid and deionised water.

Preparation of 100g of chromite ore and slag was done at a ratio of 3:1, using a SCD of 750A/m2, 600g of alkaline at 60 wt. %, air flow of 1L/min, stirring speed of 700 rpm, particle size of chromite and slag of less than 200 mesh and a temperature of 160? was done.

Reaction times of 4hrs, 6hrs and 8hrs were used to investigate the effect of residence time.

Steps from experimental procedure to get potassium dichromate were followed using different residence times.

Experiments were repeated 3 times for each time span.

CHAPTER FOUR: EXPERIMENTAL RESULTS AND ANALYSIS4.0 Introduction
This chapter highlights the link between the experimental data acquired and the future work to be done. This information is important in both the process and equipment design stages of this project.
4.1 Experimental results and analysis
4.1.1 Experiment 1
This was done to analyse the percentage composition of Cr2O3, FeO, Al2O3 and MgO in chromite ore and slag respectively. The use of and X-ray fluorescence spectrometer was employed.

Table 4.1 Composition of low grade chromite ore
Component Cr2O3 FeO MnO2 TiO2
% wt. 24.21 12.76 0.267 0.166
Table 4.2 Composition of EAF slag
Component Cr2O3 FeO MnO2 TiO2
% wt. 11.42 4.03 0.222 0.222
The above results show that there is a considerable amount of chromium in the chromite spinel. With 24.21% in low grade chromite ore and 11.42% for EAF slag.

For a 100g reaction sample of low grade ore and slag at a ratio of 3:1, the following calculations hold true;
% Cr2O3in slag=11.42100×25g=2.855g% Cr2O3in low grade ore=24.21100×75g=18.1575gTotal Cr2O3 before reaction=21.0125g4.1.2 Experiment 2
Done to investigate the effect of different alkali concentrations on potassium dichromate recovery.

The samples of residue at different alkali concentrations were analysed using the XRF spectrometer. The data recorded was then used to get % conversion, using;
% conversion=(1-CrrCro)×100Where; Crr=concentration of chromium in residueCro=concentration of chromium in ore and slagTable 4.3 Effect of varying alkali concentrations on recovery
Concentration 50 wt. % 55 wt.% 60 wt.% 65 wt.%
Cr0 21.0125 21.0125 21.0125 21.0125
Crr 5.799 2.244 1.5906 3.699
% conversion 72.494 89.32 92.43 82.393

Fig 4.1 Percentage conversion at different alkali concentrations
Discussion- The graph, fig 4.1 shows that the extraction rate of chromium exhibited a parabolic trend with the increase in alkali concentration. However, with the increasing alkali concentration, the medium viscosity increases dramatically, and the salting-out effect will increase as well, resulting in a significant decrease in the mass transfer rate and the oxygen solubility. Refer to Appendix 1.

It is clear from Appendix 1 that as alkali concentration was increased, except for the mean activity coefficient of the dissolved oxygen, the dissolved oxygen solubility and the oxygen diffusion coefficient both decrease, highlighting that while a high alkali concentration is more beneficial for the oxidation reactions in terms of thermodynamic consideration, it is not the case if the reaction kinetics are considered. Thus an alkali concentration of 60 wt. % was chosen for the following experiments.

4.1.3 Experiment 3
Done to determine the optimal reaction temperature whilst investigating the effect of different temperatures on potassium dichromate recovery.

The reaction temperature is another important variable that has a considerable impact on the extraction of chromium taking both thermodynamic and kinetics properties into account.

Table 4.4 Effect of different temperatures on conversion
Temperature 100? 130? 160? 190?
Cr0 21.0125 21.0125 21.0125 21.0125
Crr 9.456 4.555 3.950 7.264
% conversion 54.998 78.8 81.2 65.43

Fig 4.2 Percentage conversion at varying temperatures
Discussion- effect of temperature on the extraction of chromium also exhibits parabolic behaviour. This was expected because at very low and high temperatures oxygen dissolution is unfavourable. Thus a reaction temperature of 150 ? was chosen.

4.1.4 Experiment 4This was carried out to investigate the effect of slot current density to potassium dichromate recovery. Current density is a measure of the density of an electric current. It is defined as a vector whose magnitude is the electric current per cross-sectional area. In SI units, the current density is measured in amperes per square metre.

The oxidation of chromite can be greatly enhanced in the electrochemical field due to the direct and indirect electrochemical oxidation of spinel.

Table 4.5 Effect of current on recovery
Current 200A/m2 400A/m2 600A/m2 800A/m2
Cr0 21.0125 21.0125 21.0125 21.0125
Crr 12.116 7.369 4.005 1.6378
% conversion 42.34 64.93 80.94 92.2

Fig 4.3 Conversion at different current densities
Discussion- Fig 4.3 shows that the extraction rate of chromium increased with increasing current density, with the highest decomposition of chromite occurred at the maximum current density of 800 A/m2. This is because high values of current are more favourable for the direct anodic oxidation of spinel as well as the electrochemical synthesis of reactive oxygen species (ROS). During electrolysis, large amounts of nano sized oxygen bubbles are produced due to the anodic oxidation of hydroxyl ions (Moureaux et al., 2013), which increase the oxygen solubility in the medium, thereby intensifying the oxidation reactions. However, due to the severe salting-out effect, oxygen has very low solubility in this highly concentrated alkaline solution at these high temperatures. Therefore, with the further increase of current density to 800 A/m2, the percentage of dissolved oxygen no longer increases, even though the anodic evolution of oxygen becomes much more intense, and the chromium extraction rate only slightly increases, resulting in unnecessary consumption of power. Thus a current density of 750 A/m2 was chosen for the following experiments.

4.1.5 Experiment 5
To investigate the effect of different alkali to ore mass ratios on the percentage of potassium dichromate recovery, was the agenda of this experiment.

Table 4.6 Effect of alkali to ore mass ratio on recovery
Ratio 3:1 4:1 5:1 6:1
Cr0 21.0125 21.0125 21.0125 21.0125
Crr 7.875 4.982 1.780 1.208
% conversion 62.52 76.29 91.53 94.25

Fig 4.4 Effect of alkali to ore mass ratio on conversion
Discussion- The results show that the extraction rate of chromium marginally increases when the alkali to ore ratio increases from 3:1 to 5:1, which is attributed to the fact that the increase of the alkali to ore mass ratio decreases the medium viscosity, favoring the mass transfer of oxygen and ore particles
4.1.6 Experiment 6This was done to investigate the effect of residence time on percentage recovery of K2Cr2O7.

Table 4.7 Effect of residence time on recovery
Time 2 hours 4 hours 6 hours 8 hours
Cr0 21.0125 21.0125 21.0125 21.0125
Crr 7.911 2.177 1.605 1.248
% conversion 62.35 89.67 92.36 94.06

Fig 4.5 Effect of residence time on recovery
Discussion – Extraction rate of chromium increase with residence time as seen by the fig 4.5. However with further increase of time, the recovery rate increases slightly. Thus it would not be economic to use more energy just for a small increase in recovery.

4.2 Conclusion
From the results obtained, the best optimum operating conditions for the electrochemical oxidation of low grade chromite ore and slag is at 60 wt. % alkali, current density of 750 A/m2, ore to slag ratio of 3:1, airflow of 1L/min, stirring speed of 700rpm, a particle size distribution of less than 75µm, reaction temperature of 160? and a residence time of 4 hrs.

At these conditions the expected recovery is 89.67%.

CHAPTER FIVE: PROCESS DESIGN5.0 Introduction
The process design’s purpose is to make clear the process being with the aid of process flow diagrams, material and energy balances. It shows a detail outline of what the process will entail.

5.1 Process description
5.1.1 Crushing
Comminution means the reduction of solid materials from one average particle size to a smaller average particle size. This can be done by crushing, grinding, cutting, which in the case of the proposed project relates to an increase in the rate and percentage of recovery of potassium dichromate as the surface area of contact would have increased.

The type of crusher to be used is a single toggle jaw crusher. The agenda is to reduce the boulder size from around 200-400mm to a third of the initial size. Only the low-grade chromite ore will be processed using a jaw crusher since the particle size of the slag is already small enough to go straight to milling at 0.1-0.2mm.

5.1.2 Milling and classification
The second stage of comminution after crushing is grinding which is the final stage of comminution. Particle size reduction is achieved by a combination of abrasion and impact. Grinding is performed in rotating cylindrical vessels known as tumbling mills. The mills contain loose grinding media which are large, hard and heavy relative to the ore particles but small in relation to the mill volume. When the mill is rotated the mixture of the medium, ore and water (mill charge) is intimately mixed. Comminution occurs by any of the three mechanisms i.e. impact/chipping, compression, and abrasion. This process is usually performed wet although dry process is possible. In this case wet grinding was chosen.

For classification the use a hydro-cyclone will be employed. This is a continuously operating classifying device that utilizes centrifugal force to accelerate the settling rate of particles. It is one of the most important devices in the minerals industry, its main use in mineral processing being as a classifier, which has proved extremely efficient at fine separations. The target being 55% passing 75µm.

5.1.3 Electrolysis
Occurring in a batch reactor with steel electrodes, this process is the major process since it is here that potassium dichromate is formed from the chrome spinel.

There are two major reactions occurring during this stage, one at the cathode and the other at the anode. The electrolyte used in this case is concentrated potassium hydroxide with potassium dichromate as the catalyst. What occurs is the oxidative decomposition of chromite ore.

Cathodic reaction;
2FeO.Cr2O3+2OH-+12O2?Fe2O3+Cr2O72-+H2OAnodic reaction;
O2+e-?O2-These reactive oxygen species (ROS), are stable super oxides that aid in promoting oxidative decomposition of chromite ore.

5.1.4 Dilution
For the purpose of removing and recycling potassium hydroxide, this process is included. Water at 25? is added to the slurry after electrolysis. This is done because potassium hydroxide dissolve in water at this temperature but not as much for potassium dichromate.

Table 5.1 KOH solubility in water
Mass dissolved in g/ml of water Temperature ?
85g/100ml -23.2
97g/100ml 0
121g/100ml 25
138.3g/100ml 50
162.9g/100ml 100
Potassium dichromate at 100? dissolves at 102g/100 ml water and at 20?, it dissolves at 13g/100ml water.

5.1.5 Filtration
Filtration is a process of separating solids from liquid by means of a porous medium which retains the solid but allows the liquid to pass. To separate the filtrate which is rich in potassium hydroxide and the potassium dichromate rich cake, the use of a pressure filter will be employed.

5.1.6 Leaching
Leaching is the preferential selection of one or more constituents of a solid mixture by contact with a liquid solvent. After filtration the filter cake is then conveyed into the leaching tank. Here water at 120? is used as a leachate since the wanted potassium dichromate readily dissolves in water at this temperature.

5.1.7 Filtration
The use of a pressure filter is employed again. In this instance the filtrate is now the wanted substance. The iron rich cake is disposed of in dumps as it is now more environmentally friendly and can be used as raw material in the manufacture of tar or it can be used as a source of iron. The filtrate is rich in potassium hydroxide.

5.1.8 Evaporation
The use of a forced circulation evaporator will be employed. This because it is well suited for crystallizing operations and when the product is toxic or corrosive. The agenda is to retrieve highly concentrated potassium dichromate product which is a mixture of potassium dichromate crystals and moisture.

5.1.9 Drying
This is the final process and its feed comes from the evaporation section. The use of a spray dryer will be employed. The rotary is a type of industrial dryer employed to reduce or minimize the liquid moisture content of the material it is handling by bringing it into direct contact with a heated gas. It has wide applications and some of those are in the mining industry.

5.2 Block flow diagram314325074352153371850499110027051003943350010496554133850015430504413250189088713463395619751088619Comminution
Comminution
39624006795135Drying
00Drying
3552825461010Oxygen
0Oxygen
214185551435KOH and catalyst
00KOH and catalyst
101917551435Low grade ore
00Low grade ore
666751327785H20
00H20
476250527685Slag
00Slag
1714501232959Electrolysis
Dilution
Filtration
Leaching
Filtration
Crystallisation
K2Cr2O7
Evaporation
Residue washing
Fe enriched residue
Water
Recycled KOH
Recycled water
Electrolysis
Dilution
Filtration
Leaching
Filtration
Crystallisation
K2Cr2O7
Evaporation
Residue washing
Fe enriched residue
Water
Recycled KOH
Recycled water

Pipe Colour codes
Colour Meaning
1752601028700 Service air
20383510985500 Non-potable and hot water
2038359779000 Chemicals
25209510477500 Steam
Fig 5.1 Plant process flow diagram with colour codes
5.4 Material balances
The knowledge of material balance is a useful tool in determining the quantity of raw material required and the products produced. The simple material balance acquaints the process designer the maximum yield obtainable in this process while providing a quick check on the profitability of the proposed process.

The general conservation equation for the process system can be written as:
Material out = Material in + Generation – Accumulation………. (1)
Material out = Material in………… (2)
5.4.1 Background data
The agenda is to design a plant producing 1.6 t/day of potassium dichromate. The major reaction occurring is;
2FeO.Cr2O3+2OH-+12O2?Fe2O3+CrO72-+H2OThe ratio of chromite spinel to potassium dichromate is 2:1 respectively. From experimental data the potassium dichromate produced has an average purity of 92.1%. The remainder is accounted to moisture and impurities. Thus the 1.6 tonnes produced will also have the average purity of 92.1%.

Mass of pure potassium dichromate=92.1% of 1600kg Mass of pure potassium dichromate=1473.6kgMrK2Cr2O7=296MrFeO.Cr2O3=224nK2Cr2O7=mMr=1473.6kg296g/mol=4.9784kmolesnFeO.Cr2O3=2×nK2Cr2O7=2×4.9784=9.95675kmolesmFeO.Cr2O3=n×Mr=9.95675kmoles×224=2230.368kgHowever according to experimental data, the chromium content of a combined sample of ore and slag is 21.0125g per 100g of sample. This relates to 21.0125% of chromium before the reaction occurs for any given mass of sample.

Using simple proportion;
Mass of ore for 1.6 tonnes of product=10021.0125×2230.368kg=10614.48kg10614.48kg?10615kgTherefore the mass of combined feed ore before any processing is 10615kg. This is the mass required to produce 1.6 t/day of potassium dichromate.

The ratio of low grade ore to slag is 3:1 from experimental data. This ratio can be used to calculate the separate masses of ore and slag required.

Mass of low grade ore=34×10615kg=7961.25kgday=331.720 kg per hourMass of slag=14×10615kg=2653kgday=110.57 kg per hour5.4.2 Mass balance around jaw crusher
Assuming dust losses are negligible. Since no reaction occurs during this stage the following mass balance equation holds true;
Material out = Material in
Basis =1 hour
Jaw crusher
331.725kg/hr
331.725kg/hr
Jaw crusher
331.725kg/hr
331.725kg/hr

5.4.3 Mass balance around ball mill and cyclone 1The milling section serves to grind the ore to a size sufficiently fine to free the valuable minerals from one another and from the gangue (valueless) minerals. The grinding method used is a closed circuit wet milling operation and this calls for enough water at the mill feed end to achieve the required water quantities suffice for efficient wet milling. Ball mill 1 is for low grade ore. Since there is no reaction occurring the mass balance equation holds true;
Material out = Material in
Basis=1 hour
Mill Feed 0.332t/hr
Overflow 55% solids
Discharge Pump
Ball Mill
Discharge
Sump
Feed Water 30% of solids
Underflow 83% solids
Dilution Water 10% of solids
72% Solids
0.2324t/hr solids
0.047t/hr
0.096t/hr
0.033t/hr
Mill Feed 0.332t/hr
Overflow 55% solids
Discharge Pump
Ball Mill
Discharge
Sump
Feed Water 30% of solids
Underflow 83% solids
Dilution Water 10% of solids
72% Solids
0.2324t/hr solids
0.047t/hr
0.096t/hr
0.033t/hr

Around ball mill
Feed water=30100×0.332thr=0.096 t per hr=0.096m3/hrDilution water=10100×0.332thr=0.033t per hr=0.033m3/hrTotal water=0.129m3Total volume in dischrge sump=0.461 tonnes per hourCyclone feed density=amount of solidstotal cyclone feed×100=0.3320.461×100=72% solidsPulp density=100s100s-xs-1Where s=specific gravity of ore
D=pulp density
x=% solids
Pulp density=100(4.76)100(4.76)-0.724.76-1=1.006Around cyclone
Overall mass balance;
F=U+OWhere,F=Total solids in feed
U=solids in underflow
O=solids in overflow
Component balance using dilution ratios;
fIF=uIU+oIOWhere fI=dilution ratio of feed slurry
uI=dilution ratio of underflow
oI=dilution ratio of overflow
718F=1783U+911O…..i0.332=U+O…..iiSolving simultaneously;
O=0.0996 t per hourU=0.2324 t per hourCirculating load=oI-fIfI-uI×100Circulating load=233.2% From this data we can calculate the water in both underflow and overflow using simple proportion.

Owater=0.08149 m3/hrUwater=0.0476 m3/hrTotal water entering cyclone;
Total water in cyclone=Owater+Uwater=0.129 m3/hr5.4.4 Mass balance around ball mill and cyclone 2Ball mill 2 is for EAF slag. Since there is no reaction occurring the mass balance equation holds true;
Material out = Material in
Basis=1 hour
Mill Feed 0.111t/hr
Overflow 55% solids
Discharge Pump
Ball Mill
Discharge
Sump
Feed Water 30% of solids
Underflow 83% solids
Dilution Water 10% of solids
71.4% Solids
0.0333t/hr
0.0111t/hr
0.0756t/hr
0.015484t/hr
Mill Feed 0.111t/hr
Overflow 55% solids
Discharge Pump
Ball Mill
Discharge
Sump
Feed Water 30% of solids
Underflow 83% solids
Dilution Water 10% of solids
71.4% Solids
0.0333t/hr
0.0111t/hr
0.0756t/hr
0.015484t/hr

Around ball mill
Feed water=30100×0.111thr=0.0333 t per hr=0.0333m3/hrDilution water=10100×0.111thr=0.0111t per hr=0.0111m3/hrTotal water=0.0444m3Total volume in dischrge sump=0.1554 tonnes per hourCyclone feed density=amount of solidstotal cyclone feed×100=0.1110.1554×100=71.4% solidsPulp density=100s100s-xs-1Where s=specific gravity of ore
D=pulp density
x=% solids
Pulp density=100(4.76)100(4.76)-0.7144.76-1=1.006Around cyclone
Overall mass balance;
F=U+OWhere,F=Total solids in feed
U=solids in underflow
O=solids in overflow
Component balance using dilution ratios;
fIF=uIU+oIOWhere fI=dilution ratio of feed slurry
uI=dilution ratio of underflow
oI=dilution ratio of overflow
143357F=1783U+911O…..i0.111=U+O…..iiSolving simultaneously;
O=0.0354 t per hourU=0.0756 t per hourCirculating load=oI-fIfI-uI×100Circulating load=213.4% From this data we can calculate the water in both underflow and overflow using simple proportion.

Owater=0.028964 m3/hrUwater=0.015484 m3/hrTotal water entering cyclone;
Total water in cyclone=Owater+Uwater=0.0444 m3/hr5.4.5 Mass balance around electrolytic reactor
Assuming water loses due to evaporation are negligible since water is added periodically to replace this water. A reaction occurs during electrolysis in the presence of oxygen and a potassium carbonate catalyst. The feed rate into reactor be designed as required to produce 1.6 t/day since there is the presence of surge tanks that act as storage vessels of slurry from the mills. These are the ones that supply the reactor with feed and also act as a buffer to the whole process. The low grade to slag ratio of 3:1 is taken into account.

Material out = Material in + Generation – Accumulation
Basis =4 hours
Reactor
Oxygen
KOH=0.19395t
Solids=0.443t
Water=0.129t
K2Cr2O7=0.2626t
Fe2O3=0.1419t
Water=0.129t
KOH=0.09285t
Inerts=0.1397t
Reactor
Oxygen
KOH=0.19395t
Solids=0.443t
Water=0.129t
K2Cr2O7=0.2626t
Fe2O3=0.1419t
Water=0.129t
KOH=0.09285t
Inerts=0.1397t

Low grade chromite ore;
0.332t solids
0.096m3 water
EAF slag;
0.111t solids
0.0333m3 water
Total solids0.443tonnes
Total water0.1293m3
Reactions occurring
Cathodic reaction;
2FeO.Cr2O3+2OH-+12O2?Fe2O3+Cr2O72-+H2OAnodic reaction;
O2+e-?O2-KOH
Potassium hydroxide is added at 60 wt. %. The total water in the system is 0.1293m3. Using simple proportion;
Total potassium hydroxide required=6040×0.1293=0.19395tonnesAssuming 100% conversion
nFeO.Cr2O3=mMr=443kg224g/mol=1.978kmolesnK2Cr2O7=mMr=0.5nFeO.Cr2O3=0.98884kmolesmK2Cr2O7=n×Mr=0.98884kmoles×296=292.7kgmKOH=n×Mr=1.978kmole×57=112.728kgAt 89.67% conversion
mK2Cr2O7=89.67% of 292.7kg=0.2625tonnesmFe2O3=n×Mr=0.98884kmoles×160=0.1419tmKOH=89.67% of 112.728kg=0.1011tmKOHleft=0.19395t-0.1011t=0.09825tTotal amount of reactants=0.444t+0.1293t+0.19395t=0.76625tMass of inerts=0.76625- mK2Cr2O7+ mFe2O3+mKOH+mwaterMass of inerts=0.1397tOxygen for the reaction is added at a rate of 1L/min continuously. This helps to maximise conversion. However using the reaction equation at 89.67%, 14.18kg of oxygen is required to form a mole of potassium dichromate.

5.4.6 Mass balance around dilution tank
Assuming potassium hydroxide dissolve completely and that dissolution of potassium dichromate is negligible at 25?.

Material out = Material in
Basis=4 hours
`Dilution tank
Dilution water=0.46425t
K2Cr2O7=0.2626t
Fe2O3=0.1419t
Water=0.129t
KOH=0.09285t
Inerts=0.1397t
 
 
 
Water=0.59355t
K2Cr2O7=0.2625t
KOH=0.09285t
Fe2O3 +Inerts=0.2816t
Dilution tank
Dilution water=0.46425t
K2Cr2O7=0.2626t
Fe2O3=0.1419t
Water=0.129t
KOH=0.09285t
Inerts=0.1397t
 
 
 
Water=0.59355t
K2Cr2O7=0.2625t
KOH=0.09285t
Fe2O3 +Inerts=0.2816t

Using experimental data, the ratio of dilution water and potassium hydroxide is 5:1 respectively.

mdilution water=5×0.09285t=0.46425t=0.46425m35.4.7 Mass balance around pressure filter
Assuming negligible solid loses to filtrate stream and that the dissolution of potassium dichromate is negligible. The pressure filter will function with a dewatering efficiency of 95% and 92% of potassium hydroxide dissolves from experimental data.

No reaction occurs in the pressure filter, thus the equation below holds true;
Material out = Material in
Basis=4 hours
Pressure filter
Water=0.59355t
K2Cr2O7=0.2625t
KOH=0.09285t
Fe2O3 +Inerts=0.2816t
 
 
KOH=0.085422t
Water=0.563872t
Water=0.02968t
K2Cr2O7=0.2625t
KOH=0.007428t
Fe2O3 +Inerts=0.2816t
Pressure filter
Water=0.59355t
K2Cr2O7=0.2625t
KOH=0.09285t
Fe2O3 +Inerts=0.2816t
 
 
KOH=0.085422t
Water=0.563872t
Water=0.02968t
K2Cr2O7=0.2625t
KOH=0.007428t
Fe2O3 +Inerts=0.2816t

mKOHfiltrate=92%×0.09285t=0.085422tmKOHcake=0.09285t-0.085422=0.007428tWater in filtrate is 95% of water entering this unit;
mwaterfiltrate=0.95×0.59355m3=0.563872m3mwatercake=0.59355m3-0.563872m3=0.02968m3The filtrate is then taken to an evaporator for concentration after which it is recycled back into the process.

5.4.8 Mass balance around leaching tank
No reaction occurs in the leaching tank. Water at 120? is used as the leachate at a ratio of 4:1 with the potassium dichromate.
As mentioned above no reaction occurs, only dissolution. Thus the equation below holds true;
Material out = Material in
Basis= 4hours
Leaching tank
Water=1.05t
Water=0.02968t
K2Cr2O7=0.2625t
KOH=0.007428t
Fe2O3 +Inerts=0.2816t
 
Water=1.07968t
K2Cr2O7=0.2625t
KOH=0.0.007428t
Fe2O3 +Inerts=0.2816t
Leaching tank
Water=1.05t
Water=0.02968t
K2Cr2O7=0.2625t
KOH=0.007428t
Fe2O3 +Inerts=0.2816t
 
Water=1.07968t
K2Cr2O7=0.2625t
KOH=0.0.007428t
Fe2O3 +Inerts=0.2816t

mwater=4×0.2625t=1.05t=1.05m35.4.9 Mass balance around pressure filter
Assuming 100% dissolution of potassium dichromate recovered during electrolysis. The dewatering efficiency of the pressure filter is set at 95%. No reaction occurs in the pressure filter thus the equation below holds true;
Material out = Material in
Basis= 4hours
Considering that 100% dissolution of potassium dichromate occurs;
Total mass of water and potassium dichromate=1.34218tmfiltrateK2Cr2O7+water=0.95×1.34218t=1.275071tmK2Cr2O7 in filtrate=0.95×0.2625t=0.249375tmwater in filtrate=1.275071t-0.249375t=1.025696t=1.025696m3mwater in cake=1.07968m3-1.025696m3=0.053984m3Pressure filter
Water=1.07968t
K2Cr2O7=0.2625t
KOH=0.007428t
Fe2O3 +Inerts=0.2816t
 
Water=1.025696t
K2Cr2O7=0.249375t
KOH=0.0.007428t
Fe2O3 +Inerts=0.2816t
Water=0.053984t
Fe2O3 +Inerts=0.2816t
Pressure filter
Water=1.07968t
K2Cr2O7=0.2625t
KOH=0.007428t
Fe2O3 +Inerts=0.2816t
 
Water=1.025696t
K2Cr2O7=0.249375t
KOH=0.0.007428t
Fe2O3 +Inerts=0.2816t
Water=0.053984t
Fe2O3 +Inerts=0.2816t

5.4.10 Mass balance around evaporator
With a boiling point of 500?, potassium dichromate does not evaporate at 140?.

For the purpose of this material balance, the evaporative efficiency of the forced circulation evaporator will be assumed to be the same as the efficiency found during the experiment, which is 95%.

Since no reaction occurs the equation below holds true;
Material out = Material in
Basis= 4 hours
mwater that evaporates=0.95×1.025696t=0.974406t=0.974406m3mwater in product stream=1.025696m3-0.974406m3=0.0512905m3Evaporator
Water=1.025696t
K2Cr2O7=0.249375t
KOH=0.007428t
 
Water=0.974406t
Water=0.0512905t
K2Cr2O7=0.249375t
KOH=0.0.007428t
Evaporator
Water=1.025696t
K2Cr2O7=0.249375t
KOH=0.007428t
 
Water=0.974406t
Water=0.0512905t
K2Cr2O7=0.249375t
KOH=0.0.007428t

5.4.11 Mass balance around dryer
Relating the experimental data of potassium dichromate purity of 92.1%, the dryer for this design would have to function at an efficiency of 72.8%.

Material out = Material in
Basis= 4 hours
mwater that evaporates=0.728×0.0512905t=0.0373395t=0.0373395m3mwater in product=0.0512905t-0.0373395t=0.013951t=0.013951m3Dryer
Water=0.0512905t
K2Cr2O7=0.249375t
KOH=0.007428t
Water=0.0373395t
Water=0.013951t
K2Cr2O7=0.249375t
KOH=0.007428t
Dryer
Water=0.0512905t
K2Cr2O7=0.249375t
KOH=0.007428t
Water=0.0373395t
Water=0.013951t
K2Cr2O7=0.249375t
KOH=0.007428t

5.4.12 Conclusion
Total mass of product=0.270754tPurity of potassium dichromate=0.249375t0.270754t×100=92.1%Moisture content in product=0.013951t0.270754t×100=5.2%Potassium dichromate of 92.1% purity and 5.2% moisture is produced. Taking into account that the residence time in the electrolytic reactor is 4hours and proposing mine wise working hours. These are 24 hours a day with different shifts. Thus in a single day the reactor works six times since the reaction time is 4 hours.

Mass of potassium dichromate a day=6×0.270754t=1.62 tonnesThus the target of 1.6 tonnes per day is feasible to reach.

5.5 Energy balances
In chemical and process systems engineering processes energy is usually generated or lost. Material can change its state, new material species can be formed. However the total energy flow into a process unit must be equal to the flow out at steady state. The need for energy balances is to be determine the energy requirement of a process that is heating, cooling and power requirements.

Using the first law of thermodynamics,
Eout=Ein+Egenerated-Econsumed-EaccumulatedWhere E- energy
5.5.1 Energy balance around jaw crusher
Comminution of solid materials consumes energy which is being used to break up the solid into smaller pieces. Bond’s law is regarded as the most reliable and useful relation for expressing the relationship between particle size and crushing energy.

Bond’s law;
E=10×W1P80-1F80WhereW=bond work index
E=energy required
F80=Diameter which 80% of feed passes
P80=Diameter which 80% of product passes
The bond work index of chromite ore is 9.6
Average feed size is 250mm and the product size is a quarter of this value.

E=10×9.6kWhr/mm10.0625m-10.250mE=192kWhrThis is the energy required to reduce the ore size from 250mm to 62.5mm
5.5.2 Energy balance around ball mill 1
Around ball mill 1, the feed particle size is the product stream particle size of the jaw crusher. Since the feed rate is also the same the value of bond work index at 7.968kWhr holds true. From experimental data the product particle size going out of the mill is 90% passing 75µm.

Using Bond’s law;
E=10×W1P80-1F80E=10×9.6kWhr/mm10.000075m-10.0625mE=10 701kWhrThis is the energy required to reduce size of ore from 62.5mm to 75µm.

5.5.3 Energy balance around ball mill 2
Assuming the average bond work index of hard ores as the work index for chromite ore and slag, given as 24 Kw hr/tonne. For a feed rate of 0.111 t/hr the use of simple proportion is employed to get the following results;
W=0.111t1t×24kWhr=2.664kWhr/mmUsing Bond’s law;
E=10×W1P80-1F80With a feed particle size of 0.1-0.2mm, the average feed particle size is therefore regarded as 0.15mm. The product particle size is 90% passing 75µm.

E=10×2.664kWhr/mm10.075mm-10.15mmE=28.49kWhrThis is the energy required to mill ore from 0.15mm to 75µm.

5.5.4 Energy balance around reactor
Assuming that;
Internal energy is not a function of molar volume
Mixture is an ideal mixture
Heat capacity is constant
Reactor volume is constant
Reaction is well insulated.

Eaccumulated=Einternal+Eadded to systemEnergy due to oxidation reaction
2FeO.Cr2O3+2OH-+12O2?Fe2O3+Cr2O72-+H2O?HRXN=?HFPRODUCTS-?HFREACTANTS?HRXN=((?HFFe2O3+?HFH2O+?HFK2Cr2O7)-(2?HFFeO.Cr2O3+2?HFKOH+0.5?HF(O2))?HRXN=((-824.2)+(-286)+-2033)-(2-1730.13+2-424.72+0.5?HF(0))?HRXN=317kJ/molThe reaction is exothermic, it generates heat on its own.

Energy from electric current being passed
Using the formula;
P=I×VWhere P=power in watts
I=amperage in amps
V=voltage in volts
P=750A×2.5V=1875 Watts=1.875kWEnergy due to heat added to reactor
The heating media is water, with a heat capacity 4.18kJ/mol and the reaction temperature is 150?. Water is thus heated from 25? to 150?.

Q=mCP?TQ=100kg×4.18KJ/kg?×(150-25)?Q=31 350kJTotal energy
Total energy=31350kJ+1.875kJ+317kJ=31 668.9kJ5.5.5 Energy balance around forced circulation evaporator
We can start by applying the energy balance equation to this process:
?H +?Ek +?Ep = Q – W
For the process taking place in this problem, the following assumptions can be made:
?E = 0
Since there is no velocity difference of the inlet and outlet streams they will be neglected
?Ep = 0
There is no considerable difference in height between the inlet and outlet streams
Ws = 0
There are no mechanical parts moving
Thus, the energy balance equation will be described by:
Q = ?H
The change in the enthalpy ?H can be calculated as follows:
?H = n(Hˆ out ? Hˆ in)
The enthalpy of the liquid water entering the system can be obtained by interpolation using the data from Table 2-305 from Perry’s Chemical Engineers’ Handbook, 8th Edition, summarized in the following table.

Table 5.2: Parameters and Data for Steam
Parameter Value
Heat of vaporization of water 2260J/kg
Enthalpy of saturated steam at 110oC 2506kJ/kg
Heat capacity of water 4.2 kJ/kgK
Now we can determine the heat required to boil the feed water, by calculating the enthalpy change:
Q =?H = n(Hˆ out ? Hˆ in )
Table 5.3 Operating parameters.

Parameter Values
Datum temperature 00C
Inlet stream 250C
Temperature within evaporator 1100C
Outlet temperature 1100C
Inlet stream;
Qin=mCP?TQ=766.97kg×CP(mixture of POTASSIUM DICHROMATE AND WATER)×?THeat capacity of mixture of water and potassium dichromate going into the evaporator;
CP(Mixture)=ypotassium dichromateCP(potassium dichromate)+ywaterCP(water)CP(Mixture)=0.8045(0.209kJ/kg?)+0.1955(4.18kJ/kg?)CP(Mixture)=0.983kJ/kg?Q=1275.05kg×0.983kJ/kg?×(25?-0?)Qin=31 334.28kJOutlet stream;
Qout=mCP?TQout=300.8275kg×O.983kJ/kg?×(110-25)?Qout=25 135.64kJQvapour=m?Qvapour=974.405kg×2260J/gQvapour=2202155.3kJOverally;
Qin+Q=Qout+QvapourQ=Qout+Qvapour-QinQ=25 135.64kJ+2202155.3kJ-31 334.28kJQ=2191956.66kJUsing steam tables at 1atm, the evaporation enthalpy of steam is 2506.5kJ/kg;
Amount of heating steam=Qsteam evaporation enthalpy=2191956.66KJ2506.5KJ/kg=874.5kg
5.5.6 Energy balance around dryerInlet stream heat
Heat in potassium dichromate = mCp?T
=300.66kg×0.983kJ/kg?× 52-0 =15 368.54kJ = (15 368.54÷ (24 × 3600)) KJ/s
=0.178kW
Inlet streams enthalpy = Q + 0.178kW
Outlet stream heat
Heat out = m CP?T
=263.33kg×0.983kJ/kg?× 110-52 =15 013.49kJ
= (15 013.49 ÷ (24 × 3600))
= 0.174kW
Heat in the vapor = mCp?T
=37.395kg×4.18kJ/kg? 110-52= 9066.04kJ
= (9066.04kJ ÷ (24 × 3600))
=0.105kW
Outlet streams enthalpy = (0.105kW + 0.174kW)
=0.279kW
Overall balance
(mCp?T)in+Q=(mCp?T)out+(mCp?T)v.

?Q =0.279kW+0.105kW-0.178kW
Pipe Colour codes
Colour Meaning
1752601028700 Service air
20383510985500 Non-potable and hot water
2038359779000 Chemicals
25209510477500 Steam
Fig 5.2 Piping and instrumentation diagram with colour codes
CHAPTER SIX: EQUIPMENT DESIGN6.0 Introduction
This chapter covers the chemical, thermal and mechanical design of equipment which is used in the potassium dichromate manufacturing process. The equipment which is going to be designed includes a ball mill and forced circulation heat exchanger. Also sizing of some ancillary vessels and pumps is going to be carried out.

6.1 Forced circulation evaporator
6.1.1 Introduction
The heat exchanger used for the purpose of concentrating potassium dichromate was designed. A forced circulation evaporator is the selected evaporator of choice since it works well with substances of a corrosive nature.

6.1.2 Selection criteria for heat exchangerUpper –limit pressure requirement
Thermal performance
Temperature range
Mixture of different products
Capacity of flowing fluid
Pressure drop across heat exchanger
Ease of cleaning and maintenance
Material for construction
Cost
6.1.3 Types of heat exchangers considered1. Shell and Tube*
2. Plate heat exchanger
3. Adiabatic wheel exchanger
4. Plate Fin exchanger
5. Fluid heat exchanger
6. Waste heat recovery units
7. Phase change heat exchanger
8. Dynamic scraped surface heat exchanger
* Shell and Tube heat exchanger was designed in this project after considering the selection criteria. The shell and tube was chosen for the following reasons:
Large ratio of heat transfer area to volume and weight.

Pressures and pressure drops can be varied over a wide range
Flexibility in terms of materials of construction.

Ease of construction.

Surface can withstand normal fabrication and field erection stresses.

Ease of modification.

Ease of cleaning and maintenance.

Good design methods exist.

Expertise and shop facilities for successful design exist.

Extended heat transfer surfaces (fins) can be used.

6.1.4 Flow arrangementThree flow arrangements are available, namely:
*Counter flow HEX.

Parallel flow HEX.

Cross flow HEX.

The flow arrangements have their advantages and disadvantages for different applications in the process industry.

* For this project, the counter flow arrangement was applied.

6.1.5 AssumptionsThere is equal amount of heating surface in each pass.
The overall heat transfer coefficient is constant.
The rate of flow of each fluid is constant.
The specific heat of each fluid is constant.
There are no phase changes involving evaporation or condensation in any part of heat exchanger.
Heat losses are negligible.
The fluid properties such as temperature and velocity at inlet and outlet remain the same.
The kinetic and potential energy changes are negligible because there is is little or no change in the velocity of fluid.
Material of construction for shell is transparent.
Material of construction for tube is stainless steel.
Axial heat conduction along the tube is usually insignificant and can be considered negligible.
18BWG (Birmingham wire gauge) with mm tube pitch are used.
According to TEMA standards Baffle spacing should not be less than 1/5 of(shell diameter) .

6.1.6 Calculation of Heat exchanger designThe heat exchanger design is made up of three sections:
Chemical engineering design.

Thermal design.

Mechanical design.

6.1.7 Chemical engineering design
The chemical engineering design focuses on:
Process fluid assignments to shell side or tube side.

The fluids where allocated after considering the following factors:
Corrosion
Fouling
Fluid temperatures
Operating pressures
Pressure drop
Viscosity
Stream flow rates
Selection of stream temperature specifications.

The chemical engineering design results are summarized below:
Table 6.1 Experimental data for design
Parameter Tube side Shell side(steam)
Inlet temperature 25°C 110°C
Outlet temperature 54°C 54°C
Mass flow rate 0.09kg/s 0.06kg/s
Specific heat capacity 0.983KJ/kg. °C 4.2KJ/kg C
Thermal conductivity W/ m°C 0.69W/ m°C
Density 2.676g/cm3 1000kg/m3
Viscosity – 8.9 x10-6m2/s
Internal pipe diameter *19mm –
Tube thickness *3.01mm Outside pipe diameter *25mm –
*chosen from TEMA standards
6.1.8 Thermal design
Thermal design focuses mainly on:
1. Setting shell side and tube side pressure drop design limits.

2. Setting shell side and tube side velocity limits.

3. Selection of heat transfer models and fouling coefficients for shell side and tube side.

Several thermal design features were considered in the design as listed below:
Tube diameter: available space, cost and fouling nature of the fluids were considered
Tube thickness: thickness was determined to ensure there is enough room for corrosion, axial strength, hoop strength and buckling strength.

Tube length: HEX are cheaper when they have small diameter and long tube length. This aspect was taken into account.

Tube pitch: for an economic HEX, the tube pitch should not be less than 1.25 times the tubes outside diameter.

Tube corrugation: These types of tubes is used in inner tubes to give great turbulence.

Tube layout: four types are available, namely triangular, rotated triangular, square, rotated square. Triangular are applied where greater heat transfer is required and square where high fouling is experienced. These layouts where taken into account.

Baffle design: They are used in SHE to direct fluid across tube bundles. Various baffle types were taken into account.

Tube side design calculations
1. Calculation of heat transfer rate
Q=mCp?TcoldQ=218.625kgh-1×4.2KJ/kg?×(110-54)?Q=51 332.4kW
2. Calculation of LMTD, Tlm, cf
16313155270500
341884018034000164782518034000 hot side T1 = 110°C
cold side
T2 = 54°C163131516510000 t2 = 54°C
t1 = 25°Con hot fluid side ?T1 =110°C-54°C = 56°C
on cold fluid side ?T2 =54°C-25°C =29°C
Tm=56-29ln5629 =41.03?use one shell pass and two tube passes
R=T1-T2t2-t1=110-5454-25=1.93S=T2-t1T1-t1=54-25100-25=0.39 Ft = 0.78 {Kern, fig. 18(7)}

Therefore= ( 41.03×0.78)?=32?
3. Design overall heat transfer coefficent
Assume Uass = 3500W/m2.K
Value taken from kern.

4. Calculation of heat transfer area(AO)

=458.3m2
Assuming a 25mm outer diameter, 19mm inner diameter, 6.0m long tubes (3/4in x 16ft) made from stainless steel.

Allowing for tube sheet thickness, assume L= 6.0m
Area of one tube,?d0L= ?×0.025×6=0.0.472m2
5. Calculation of number of tubes
The objective is to find the number of tubes with diameter do, and shell diameter ds to accommodate the number of tubes, with the given tube length. The area is related to tube diameter do and number of tubes Nt ,as below

Nt = 458.3?×0.025×6 =970tubes6. Calculation of bundle diameter
The shell-side fluid is relatively clean, a 1.25 triangular pitch is chosen
Nt is the number of tubes
db is the bundle diameter (mm),
n is a constant(RK Sinnot, page 662)
Db=259000.24912.207=1023.54mmA split ring floating head type is chosen, the bundle clearance = 63mm
Shell diameter , Ds=1023.54+63=1086.54mm
7. Tube side heat transfer coefficient
ht=Nut ktdiWhere, Nut =nusselt number of fluid inside tubeskt=thermal conductivity of fluid inside tubesdi=inside diameter of tubes8. Calculation of Reynolds number
Re=Di ?v? At=?D24 = ?×0.0194 = 2.835 x 10-4 m2
tubes per pass=9702=485total flow area=485× 2.835 x 10-4 m2 = 0.1375m2
slurry mass velocity=218.625kg/s0.1375=1589.9kgs.m2Velocity in each tube(v) =1589.9kgs.m21000kg/m3 =1.589m/s
Using steam tables for steam at 110oC, the assumption is ?=0.00896kg/m.s
Hence,Re=div??= 0.019×1000×1.589 0.00896 =3394.1 Which is greater than 2000.Therefore, flow is turbulent.

9. Calculation of prandtl and nusselt number
Pr=?CPK=0.0089642000.69 =54.54Nu=0.023×Re0.8 ×Pr0.4 = 66.1Therefore, ht=Nut ktdi= 66.1×0.69 0.019 =2400.5Wm2The correlated heat transfer coefficient at the tube side =2400.5W/m2
We select stainless steel as tube material with roughness (?) =0.002mm
10. Calculation of pressure drop (?P)
?P=fLD V22?
The friction factor is determined from the Colebrook equation,
Relative roughness?D=0.0020.019=0.1052
1f =-2.0log?D3.7+2.51Ref1f =-2.0log0.0020.033.7+2.5140654f101-2f=1.892+4.2fUsing an equation solver software (ref www.algebrahelp.com) for the friction factor.

f=0.0368?P=fLD V22?=0.036860.019x 1.5892 21000?P=14 671.1Pa
11. Calculation of pump work done
Wp= ?Pm* = 14 671.1 x0.06=880.266W=2hp12. Calculation of Head
h=?P?g =151201000×9.81 = 1.54mShell side design calculations
Kern’s method was used to calculate the shell side heat transfer coefficient. The Kern Method predicts results in the range 2×103 ? Re =GsDe/ ?106
Calculation of mean diameter( De)
Baffle spacing = Ds5 =1086.545= 217.3mmtube pitch, PT=1.25do=1.25×25=31.25mmCalculation of cross flow area(As)

As=0.03125-0.025×1.08654×0.21730.03125=0.047m2Calculation of shell side mass flux

Gs=0.090.047=1.9kgm2sTherefore ,equivalent diameter ,De :

=431.252×0.87×31.25-12?1924/(0.5?×19=37.9mmMean shell side temperature=0.5110+54=82?Viscosity of steam, ?=?×?=2.676kgm3 ×0.09m/s=0.24kgm2.s
Calculation of Reynolds and Prandtl number
Res=deGs??s =37.9×1.9×2.6760.24 =802.9 Pr=?sCpsks=0.240.983×1000 0.03 =7864Calculation of shell side heat transfer coefficient
The coefficient hs is given by
js=0.0035
hs=jhkfdeRePr13??w0.14 ?w= 0.37 mNs/m2
hs=0.0035×0.19×802.9×78641337.9×10-30.24)0.00370.14 =363.98W/m2?Calculation of shell side pressure drop
The shell side pressure drop is given by:

where jf is the friction factor
linear velocity,us=Gs?=1.92.676=0.71m/s1
at Reynolds number 802.9,from RK SINNOT,page691 figure 12.30, jf = 0.04
?Ps =8×0.041086.5437.96×103217.32.676×0.71220.24)0.0037-0.14 =95kPa
Calculation of overall heat transfer coefficient
Fouling coefficient of water=3500W/m2?Fouling coefficient for steam =400W/m2?1Uo=1363.98+1400+25×10-3ln25192×45+2519×13000+2519×12400U0=1584.6 W/m2?This is below the initial estimate of 3500 W/m2? acceptable
6.1.9 Mechanical design
Data available
Shell side:
Material carbon steel
One shell two tube pass exchanger.

Fluid water vapor
Working pressure 1 atmosphere
Design pressure 0.1114 N/m2
Temperature 110?Diameter 1086.54 mm
Length 6m
Permissible stress for carbon steel is 95 N/mm2
Tube side:
Number of tubes 970
Number of passes 2
Inside diameter 19 mm
Outside diameter 25 mm
Length 6 m
Triangular pitch 1″
Working and operating pressures are same as that of shell side.

Fluid on the tube side is organic slurry:
Inlet temperature 25?Outlet temperature 54?Shell thickness
Let f=85%ts=PD2fj+P+Cc=0.1114×1086.542×0.85×95+0.1114+2mm=2.74?3mmFor practical purposes, the thickness shall be taken as 8mm.

Volume of material used for shell,Vs is given by:
VS=?×do2-di24× LVs=?×1.086542-1.0785424× 5 = 0.068m3Material of construction is carbon steel ,density
Mass of material used ,=?cs×Vs=7485.42×0.068=509kg2.Tube thickness
Using a shallow dished and tropisherical design. The minimum thickness including corrosion allowance is 3mm.
Vt=900?×do2-di24× L Vt=970?×0.0252-0.01924× 6=1.21m3 Density of stainless steel = 7485.00 kg/m3
Mass of material used =?s×Vs=7845.00kgm3×1.21m3= 9492.45kg
3.Baffles
Baffle spacing=Ds5=1086.545=217.3mmThickness of baffles(tb) is chosen as 6mm
4.Tie rods and spacers
These are provided to retain all cross baffles and tube support plates in position.

From IS:4503-1967
For shell diameter 500-1000 mm
Diameter of rod is 12 mm and number of rods = 8
5.Flange design
Flange is ring type with plain face.

Design pressure = P = 0.1114 N/mm2 (external)
Flange material: IS 2004-1962 Class 2 Carbon Steel
Bolting steel: 5% Chromium, Molybdenum Steel
Gasket Material: Asbestos composition
Shell OD,B= 1.08654 m
Shell Thickness, ts = 0.008m
Shell ID=1.07854m
Allowable stress for flange material = 100 MN/m2
Allowable stress of bolting material = 138 MN/m2
6.Determination of gasket width
do di=(y-Pm)0.5y-P(m+10.5Assume a gasket thickness of 0.6mm
y = minimum design yield seating stress = 44.85 MN/m2
m = gasket factor = 3.5
do di=(44.85-0.1114×3.5)0.544.85-0.1114(3.5+10.5d0di=1.001do=1.0011.07854=1.0796mMinimum gasket width, N =do-di2=1.0796-1.078542=0.00106mTaking minimum width as 10 mm
Then do = 1.0896m
Basic gasket seating width, b = 6 mm
Diameter at location of gasket load reaction G = di + N = 1.0796m
7.Estimation of bolt loads
Load due to design pressure
H=?PG24=?×0.1114×1.924=0.316MNLoad to keep joint tight under operation:
Hp = ? G(2b)mp= ? x 1.9 x 0.006 x 2 x 3.5x 0.1114
Hp = 0.0279 MN
Total Operating Load Wo = H+ Hp = 0.3439 MNLoad to seat the gasket under bolting condition:
Wg = ?× G× b× y =?×1.9×0.006×44.85 = 1.606MN
Wg > Wo Hence, the controlling load is Wg = 1.606 MN
8.Calculation of Minimum bolting area
Am = Ag = WgSoSo = allowable stress for bolting material
Am = Ag = 1.606138= 0.0116m9.Calculation of optimum bolting size
g1 = g/0.707 = 1.415g
Choose M18 ×2 BoltsMinimum number of bolts = 44
Radial clearance from bolt circle to point of connection of hub or nozzle and back of
Flange, R = 0.027 m
Bolt spacing , Bs = 0.045m
C = nBs ?=44×0.045? = 0.63C =ID + 2(1.415g + R)
= 1.07854 +2(1.415)(0.008)+0.027
= 1.16 m
Choose C = 1.14m
Bolt circle diameter = 1.14m
10.Flange outside diameter
Flange outside diameter, A = C +bolt diameter + 0.02= 0.764m11.Check for gasket width
AbSg? ×G×N<2ywhere Sg is the Allowable stress for the gasket material=138
Ab = actual bolt area=44×1.54×10-4=0.006776 m2
AbSg? ×G×N=0.42MPa<2y is satisfied
12.Calculation of flange thickness
t2= M CF Y B SFSF is the allowable stress for the flange material= 100MN/m2
K =AB = 1.078541.08654= 0.696For K = 0.696, Y = 4.4
Assuming CF =1
t2= 0.0123
t = 0.11m
Actual bolt spacing Bs= Cn = (3.14)(1.14)44 = 0.08mBolt Pitch Correction Factor
CF = Bs2d+t0.5= 0.596
•CF =0.772
t(act) = tוCF = 0.085m
Select 85mm thick flange. Both flanges have the same thickness
13.Heat exchanger support design
Material: Carbon Steel
Shell diameter = 1086.54mm
R =Ds2=1086.542=543.27mmTori spherical Head:
Crown radius = D,knuckle radius = 0.06×Ds=65.19mmTotal Head Depth, H = D+R2= 65.19+65.194=81.45mmShell Thickness = Head Thickness = 8mm
Maximum allowable stress = 95 MN/m2
Weight of the shell and its contents, W=509kg
Distance of saddle centre line from shell end = A = D/4=271.6mm
Flash vessel
Assuming the diameter of flash vessel is equal to that of the bundle diameter and the height is half that of the shell and tube heat exchanger.

6.2 Ball mill
The use of a ball mill will be employed during for the comminution of ore. There are mainly two types of ball mills;
Grate discharge
Overflow discharge*
An overflow discharge was designed after taking into consideration the selection criterion.
Use of the Nodberg Process Machinery Reference Manual was employed during the design of the ball mill.

6.2.1 Parameters design
Known data from experiments about chromite ore;
Specific gravity =4.06
Bond work index=9.60
Abrasion index=0.12
P80=0.075mm
F80=62.5mm
Feed rate=0.708t/hr
Using bond’s law;
E=10×W1P80-1F80E=10×9.60kWhr/mm10.075mm-162.5mmE=338.4kWhrConverting to Hp;
P=338.4kWhr×1.34P=453.5HpFor a feed rate of 0.708t/hr;
P=453.5×0.708P=320HpA motor with around 320Hp is required. The issue then becomes selecting a ball mill that will draw this power.

Using the tables provided in Appendix 3
Using factors,
Hp=A×B×C×LWhere, A=factor for diameter inside shell
B=factor for percentage loading and mill type
C=factor for speed of mill
L=length in feet of grinding chamber
Step 1
Hp=A×B×C×LStep 2
Effective diameter of a mill is the diameter inside the lining , that is the net diameter. Assuming liners will be 3 inches thick, the net diameter will thus be 6 inches less than the diameter of the shell.

Selecting a range of values from tables;
For net diameter of;
8′-6” (2590mm)A=37.3
For net diameter of;
9′-6” (2895mm)A=49.6
For a net diameter of;
10′-6” (3200mm)A=63.5
For a net diameter of;
11′-6” (3505mm)A=79.3
Step 3
Most overflow discharge mills operate with 35-45% charge. Average value is assumed to be 40%.

Assuming a percentage loading of 40%, the value of factor B is 5.02, for net overflow.

Step 4
Speed is not specified. However ball mills are operated at a percentage slower the critical speed, because grinding efficiency is poor at this point. Most mills are operated 60-90% of critical speed.

Assuming an average speed of operation of 75%. Using tables, for a value of 75% of critical speed, C=0.1838
Step 5
Solving for the length;
L=HpA×B×CAt different net diameter selected in step 2;
At 8′-6”
L=35037.3×5.02×0.1838L=10.2ftAt 9′-6”
L=35049.6×5.02×0.1838L=7.6ftAt 10′-6”
L=35063.5×5.02×0.1838L=6ftAt 11′-6”
L=35079.3×5.02×0.1838L=4.8ftData analysis
Table 6.2 Ball mill design parameters
Shell diameter Net diameter Length Length to diameter ratio
9′ (2743mm) 8’6” (2590mm) 10.2ft 1.1/1
10′ (3048mm) 9’6” (2895mm) 7.6ft 0.76/1
11′ (3353mm) 10’6” (3200mm) 6ft 0.5/1
12′ (3658mm) 11’6” (3505mm) 4.8ft 0.4/1
Selection of mill parameters will be made using the above data. Considering that for most ball mills the length to diameter ratio ranges from 1/1 to 5/1, it is thus safe to assume the parameters for the ball mill with a length of 10.2ft
Therefore, the selected design parameters are,
Table 6.3 Selected design parameters
Net diameter 2.6m
Length 3.1m
Shell diameter 2.7m
Shell thickness 0.1m
Motor power 320Hp
% loading 40
Mill speed 75% of critical speed
6.2.2 Mechanical design
Critical speed
Centrifugal force given by,
Fc=mV2R- mgcosaWhere m is the mass of the ball (kg), V is the linear velocity of the rod (ms-1), and g is the acceleration due to gravity (m s-2).

Since V =2?RN/ 60
cos?=4?²N²R60²g=0.0011N²Rcos?=0.0011N²(D)2The critical speed of the mill occurs when ?=0, i.e. the medium abandons its circular path at the highest vertical point. At this point, cos?=1.

Therefore Nc=42.3(D)
Where Nc is the critical speed.

Nc=42.32.7mNC=25.7rpm=3.63326ms-1 at 2.7m diameterSpeed of the mill= 75% of NC = 2.72ms-1
Steel ball size
Grinding balls diameter can be calculated using theory or the equation below. It is however important to note that ball size is usually chosen depending on the required particle size of product. The finer the product desired, the smaller the balls should be. However for feed with a fairly larger PSD, larger balls can come in handy.

Db=?DXEiKnrWhere, Db=ball diameter
D= mill diameter
Ei= work index of feed
nr= speed , ie percent of critical speed
?=feed specific gravity
K=a constant, 143 for balls
X=average size of feed passing 80%
Db=4.062.70.0625×9.6143×2.72Db=0.0617mTherefore, the design size of the ball mills to be used is 62mm and the material of construction is hardened high chrome.

Mill lifter bars
Single wave lifters were chosen over double wave lifters since they are generally used for mills that require steel balls that have a diameter greater than 60mm.
number of liners=6.6 Dnumber of liners=6.6×2.7=17.82?18 liftersMaterial of construction is manganese steel. Assuming a thickness of 110mm for the mill (Nordberg mineral processing manual).

Mill liners
Thickness of mill liners is usually assumed to be the same as that of the mill lifters used. The same material of construction is also employed.

Liner thickness=110mmMaterial of construction is manganesse steelShaft design
The line shaft used for moving the mill is rotating at 25.7rpm transmitting a power of 454 KW. Assuming that the material of construction is stainless steel. Stainless steel has a maximum allowable shear stress of 45MPa.

To calculate the diameter of the shaft used,
T=P×602?NWhere T= torque
P= power
N= number of revolutions
T=454 000×602?×25.7T=168 692Nm=168 692×103NmmT=??D316Therefore,
D3=168 692×103×1645×?D=267.3mm=0.3mTherefore, the diameter of the line shaft is 0.3m.

CHAPTER SEVEN: PROCESS CONTROL AND HAZOP ANALYSIS7.0 Introduction
One of the main factors considered during the processing of raw material in Chemical and process systems engineering is safety. This is safety for the environment, the chemical plant, equipment and above all human life both those working in the plant and civilians. During other plant’s operation loss of control can lead to accidents and loss of life and also property. It should therefore be mandatory that methods to help anticipate and prevent these troubles be employed. This should be done while designing, fabricating and during the operational stage of the process in the pant. Safety management generally involves;
Hazard identification techniques
Checklists
Safety reviews
Safety audits
Proper application of technical knowledge
For this project a process control and hazop analysis was done on the ball mill and the heat exchanger.

7.1 Alarms, safety, trips and interlocks
Alarms are to be incorporated to alert operators of serious, and potentially hazardous, deviations in process conditions. Key instruments are fitted with switches and relays to operate audible and visual alarms on the control panels.
Safety trip will be incorporated in all control loops involved in the process. Where it is necessary to follow the fixed sequence of operations for example, during a plant start-up and shut down, or in batch operations as in this case, interlocks are to be included to prevent operators deviating from the required sequence.

7.2 Process control around heat exchanger
Control strategy has been developed for;
Flow rate
Temperature
7.2.1 Flow rate control
There can be two types of disturbances in this process, one is the flow variation of input fluid and the second is the temperature variation of input fluid. But in practice the flow variation of input fluid is a more prominent disturbance than the temperature variation in input fluid. So, in feed forward control loop, the input fluid flow is measured and the disturbance in the flow is controlled using a feed forward controller. The output of the feedback and the feed forward controller is added and the resultant output is given to the control valve. With the addition of feed forward controller the control performance is further optimized.

The idea is to ensure that there is always fluid circulating in the tubes.

7.2.2 Temperature control
The schematic diagram of temperature control of a shell and tub heat exchanger is shown in the diagram below. Filtrate from pressure filter is supplied from the overheat tank to the shell side of the heat exchanger. Steam is supplied to the tube side of the heat exchanger. A 2-wire RTD is used to measure the output temperature of the heat exchanger and is connected to the transmitter. The 2-wire RTD transmitter produces a standard 4-20 mA output which is proportional to the temperature. The transmitter helps to reduce the noise in measurement. A separate power source is supplied to the transmitter unit. The data from the transmitter is updated in the PC based controller using a data acquisition (DAQ) device. The PC based controller processes the error signal and computes the appropriate pressure converter via another DAQ device. The current to pressure converter converts the current output of PC based controller to appropriate pressure signal so that the steam valve can be actuated in a proper manner.

Fig 7.1 Process control around heat exchanger
7.3 Process control around ball mill
Control strategy has been developed for;
Temperature on mill bearing
Level sensor in mill
7.3.1 Level control
Helps in keeping check of the quantity of substance at any given time. If the quantity of ore is too high this can lead to poor grinding and overflow of material. Very little ore also means overgrinding of ore. This will mean the particles will be too small for effective usage during electrolysis.

7.3.2 Temperature control
If the temperature of the bearings is too high, then the ball mill speed should be reduced and as well as the feed rate. Excessive temperature hikes can lead to the ball mill tripping, thus the need for effective temperature control on the mill bearings.

Fig 7.2 Process control around ball mill
7.4 Hazop
Hazard and operability study is a systematic critical examination by a team of engineering and operating experts done in order to assess the hazard potential of mal-function or mal-operation of individual items of equipment and the consequential effects on the facility, environment and lives. It should however not be seen as the ultimate review. It is only another tool to be used in the plight of safety in the processing plant. Hazop is based on a theory that assumes risk events are caused by deviations from design or operating intentions.

There are guide words in Hazop analysis. These guide words are applied to flow, temperature, pressure, liquid level and composition and based on the following parameters:
Deviation of these variables from normal operation is considered
The consequences of these deviations on the process are then assessed.

The measures needed to correct these consequences are then established
Process control strategy for temperature and flow should be done and process instrumentation should be recommended and come up with HAZOP analysis of the chosen equipment.

Table 7.1 Hazop key
NONE No forward flow when there should be
MORE More of any parameter than there should be, e.g. more flow, more temperature,
LESS As above, but “less of” in each instance
MORE OF e.g., MORE FLOW caused by reduced delivery head; surging; suction pressurised; valve stuck open leak; incorrect instrument reading.

LESS OF e.g., LESS FLOW caused by pump failure; leak; scale in delivery; partial blockage ; sediments ; poor suction head; process turndown.
LESS e.g., low temperature, pressure caused by Heat loss; vaporisation; ambient conditions; rain; imbalance of input and output; sealing; blocked vent.
MORE THAN Impurities or extra phase Ingress of contaminants such as air, water, lube oils; corrosion products; presence of other process materials due to internal leakage; failure of isolation; start-up features.

7.4.1 Hazop analysis of heat exchanger
Table 7.2 Hazop analysis of heat exchanger
Guide Word Deviation Causes Consequences Action
Less Less flow of filtrate Pipe blockage Temperature of process fluid remains constant High Temperature Alarm
More More cooling flow Failure of cooling water valve Temperature of process fluid decrease Low Temperature Alarm
More of More pressure on tube side Failure of process fluid valve Bursting of tube Install high pressure alarm
Contamination Contamination of process fluid line Leakage of tube and filtrate goes into shell Contamination of process fluid Proper maintenance and operator alert
Corrosion Corrosion of tube Potassium dichromate is highly corrosive Less heating and crack of tube Proper maintenance
None No steam in the shell side Failure of inlet valve to open Filtrate is not heated thus no concentration occurs Install flow indicator before and after the procedure
Less Flow of filtrate Failure of inlet valve to open Crystallization of potassium dichromate might not occur efficiently Install flow indicator before and after process fluid line
7.4.2 Hazop analysis of ball mill
Table 7.3 Hazop analysis of ball mill
Guide Word Deviation Causes Consequences Action
Less Less smooth rotation of mill No adequate lubrication on mill parts Noise leading to a variety of hearing related problems Install interlocks that allow mill to trip in case of lubrication system fault
Use of ear plugs(PPE)
Less Less feed ore fed into mill Damage to level and feed rate sensors Overgrinding of ore and leakages since ore is now dilute Install and maintain level and feed rate sensors
More More ore fed into mill Damage to level sensors and conveyor belt scales Overflow of material at discharge end and low power draw Install good level sensors
Poor Poor grinds Coarse ore, overfeeding , low ball load The overall recovery is low Regular PSD analysis and the use of the correct ball load
CHAPTER EIGHT: SITE SELECTION AND PLANT LAYOUT8.0 Introduction
Factors affecting site selection, decision matrix for site selection and equipment spacing assumptions are considered. Selection of the best site was done using a decision matrix, with the plant layout being made using stated assumptions.

8.1 Site selection
Location of the plant is vital on the final profitability of the project and scope for future expansion. A number of factors can be considered when selecting a suitable plant size.

8.1.1 Factors affecting site selection
Energy availability
Meteorological data
Market study
Transportation facilities
Water supply
Waste disposal
Labour supply
Taxation and legal restrictions
Site characteristics
Safety and Environmental measures
Community factors
Political and Strategic considerations
8.1.2 Suitable sites
ZIMASCO KWEKWE PLANT
This plant is the source of electric arc furnace slag. It is where chromite ore from the respective mines is processed.

REDCLIFF KWEKWE
Area is relatively out of town with easy access to the Kwekwe river, a reliable source of water.

SHURUGWI
Most chromite ore mines are located in Shurugwi. This is usually transported to Kwekwe for processing. This site would offer easy access to low grade chromite ore.

Decision matrix;
Site A – Zimasco Kwekwe plant
Site B – Redcliff Kwekwe
Site C – Shurugwi
Scoring factor;
Importance of each performance is rated on a scale from 1 to5. With 1 representing, not important and 5, vital.

8.1.3 Decision matrix
Table 8.1 Decision matrix for site selection
Criterion Rank Site performance Site total
Site A Site B Site C Site A Site B Site C
Distance to marketing area 5 5 3 3 25 15 15
Proximity to raw materials 5 4 3 4 20 15 20
Transport requirements 4 5 4 2 20 16 8
Labour availability 4 5 3 2 20 12 8
Power availability 4 5 4 3 20 16 12
Land availability 3 3 4 5 9 12 15
Waste disposal 3 4 5 4 12 15 12
Site characteristics 3 4 4 3 12 12 9
Political considerations 2 4 4 3 8 8 6
Disaster preparedness 4 4 3 2 16 12 8
TOTAL 162 133 113
From the above analysis it is clear that the most appropriate area to locate the plant is site A, the Zimasco Kwekwe plant. This because it is already functional and already has efficient transport systems in place.

8.2 Plant layout
Plant layout plays an important role in determining construction and manufacturing costs. The plant layout designed here was carefully planned with attention being given to future expansions and problems that may arise. The layout includes arrangement of processing areas, storage area, and handling areas in efficient coordination. The process units and ancillary buildings should be laid out to give the most economical flow of materials and personnel around the site. Hazardous processes must be located at a safe distance from other buildings.

The ancillary buildings and services required on a site, in addition to the main processing units (buildings), include:
Storages for raw materials
Maintenance workshops.
Stores, for maintenance and operating supplies.
Laboratories for process control.
Fire stations and other emergency services.
Utilities: Service water, power generation, refrigeration, transformer stations.
Effluent disposal plant.
Offices for general administration.
Canteens and other amenity buildings, such as medical centres.
Car parks
8.2.1 Maneuvering space
In carrying out the preliminary site layout, the process units were sited first and arranged to give a smooth flow of materials through the various processing steps, from raw material to final product storage. Process units are spaced 2m apart to allow maneuvering space. The principle ancillary building will be arranged so as to minimize the time spent by personnel in travelling between buildings. The Training centre and laboratories will be located (30m) adjacent to the processing units. Access roads will be provided to each building for construction, and for operation and maintenance. The main storage area will be placed between the loading and unloading facilities and the process units they serve.

Some of the factors considered in the site layout design includes;
Turning circle for articulated vehicles to be 62 feet (26 meters) diameter minimum. For draw-bar vehicles this can be reduced to 69 feet (21 meters).
Turning head for rigid trucks only needs to be 115 feet (35 meters) long.
Off-loading bays at 90° to road should be 102 feet (31m) deep including the road width. Bay should be 12 feet or 3.5 meters wide.
Strong site management is required to ensure maneuvering space is kept clear of storage/goods/debris at all times.
Headroom clearance should be a minimum of 15.25 feet (4.65 m) with careful consideration to ensure all pipe work, etc. is above that level. Approach gradients to flat areas will reduce the effective height.
8.2.2 Plant layout

Fig 8.1 Plant layout
CHAPTER NINE: ENVIRONMENTAL IMPACT ASSESMENT
9.0 IntroductionThis chapter provides information with regards to the impacts of activities related to the potassium dichromate manufacturing plant. The researcher carried preliminary Environmental Impact Analysis (EIA) as part of the project documentation. Under the Environmental Management Act (chapter 20:27) of 2002, it is mandatory that an EIA be carried out before any permits can be granted for construction and operation of a new mineral processing plant or mine in any area within Zimbabwe.

9.1 Environmental legislation
A variety of laws are necessary and relevant for the construction and operation of a potassium dichromate manufacturing plant. Some of them include;
Standards for Waste Management (Section 69-76)
This piece of legislation forms the basis of the EIA of the project reported herein and this project seeks to address the plight of the Standards for Waste Management section which says and l quote,
“Every person whose activities generate waste is required to employ measures that minimize wastes. The specified measures are reclamation, treatment and recycling.

Any person who contravenes this law will be guilty of an offence and liable to imprisonment of not more than five years or to a fine.”
The Environmental Management Act (Chapter 20:27) No. 13 of 2002
It forms a broad legal statement on environmental management throughout Zimbabwe and it is the ultimate piece of legislation in Zimbabwe in relation to the environment. The act encompasses all environmental quality standards for example, water pollution prohibition (section 57), Air quality standard (section 63-68), the Standards for Waste Management (Section 69-76), Standards for Noise (Section 80 – 81) etc.

Water Act (Chapter 20:24)
The act was enacted to provide for the development and utilization of the water resources of Zimbabwe, to provide the protection of the environment and prevention and control of water pollution and to provide for the approval of combined water schemes. The proposed project is in line with this act as its main aim is to try and minimize metal pollution.

9.2 Environmental impacts and mitigation
9.2.1 Site clearance and preparation
Loss of natural habitat and biodiversity
Loss of associated ecological habitats and their fauna
Intrusive activities will most probably scare away animals
Mitigation techniques include;
Activities should be restricted to within the footprint of the development
There should be no side tipping of excavated material
Animals that can be relocated to safer areas like conservation parks, must be removed from the area.

Dust
Air borne particulate matter (dust) will be generated. This most likely worsens during the dry season.

Dust can pose a hazard risk to residents in the vicinity
Mitigation techniques include;
Access roads and ground should be wetted regularly
Workers on site should be issued with dust masks when working
Noise
Heavy machinery used during site clearance and construction generates a lot of noise
Mitigation techniques include;
Workers operating equipment that generates noise should be equipped with noise protection gear.

Workers operating equipment generating noise levels greater than 80 dB continuously for 8 hours or more should use earmuffs. Workers experiencing prolonged noise levels of 70 – 80 dB should wear earplugs.

If necessary, local residents should be given notice of intended noisy activities so as to reduce degree of annoyances
9.2.2 Construction impacts
Loss of land for other uses
The construction of a potassium dichromate plant will involve building large embankment structures. This will result in a loss of the options for alternative land use and thus represents an irreversible commitment of land resources.

Mitigation techniques include;
No viable mitigation but to make the most of every square inch of land used.

Materials transportation
The various materials required for pond and building construction e.g. steel, blocks and lumber, marl etc, will be obtained from sources elsewhere and transported to the site. Transportation of these materials, typically in over-laden and sometimes uncovered trucks, usually results in undue road wear-and-tear. In the case of fine earth materials, dusting and spillages occur on major roadways between source and site. Dusting degrades local air quality and material spillages worsen driving conditions and increase the risk of road accidents. These occurrences represent indirect, short-term, reversible, negative impacts on public health and safety.

Mitigation techniques include;
All fine earth materials must be enclosed during transportation to the site to prevent spillage and dusting. Trucks used for that purpose should be fitted with tailgates that close properly and with tarpaulins to cover the materials.
The clean-up of spilled earth and construction material on the main roads should be the responsibility of the Contractor and should be done in a timely manner (say within 2 hours) so as not to inconvenience or endanger other road users.
The transportation of lubricants and fuel to the construction site should only be done in the appropriate vehicles and containers, i.e. fuel tankers and sealed drums
Appropriate traffic warning signs, informing road users of a construction site entrance ahead and instructing them to reduce speed, should be placed along the main road in the vicinity of the entrance to the Longlands plant.

Flagmen should be employed to control traffic and assist construction vehicles as they attempt to enter and exit the project site. Also As far as possible, transport of construction materials should be scheduled for off-peak traffic hours. This will reduce the risk of traffic congestion and of road accidents on the access roads to the site.

Building material sourcing
Earth materials needed for construction (e.g. marl, sand) are normally obtained from quarry and mining operations. Conscious or unwitting purchase of these materials from unlicensed operations indirectly supports, encourages and promotes environmental degradation at the illegal quarry sites and causes medium to long-term negative impacts at source.

Mitigation techniques include;
Earth materials must be obtained from officially licensed and approved quarries and copies of the relevant licenses made available for inspection at the site by the Contractor.

9.2.3 Operation impacts
Water supply
Workers at the facility will demand water for drinking, washing, and flushing toilets. This demand will be insignificant in terms of resource depletion and impact on the local water supply network.

Use of electricity
The Zimbabwe Electricity Supply Authority (ZESA) will supply power for the development site from the existing mains. The incremental demand will be within the capacity of the system and this will be confirmed in writing by the utility. The expansion should therefore not cause any supply shortages to the rest of the system. However, this increased demand will commensurately increase the use of fossil fuel to generate that electricity, and thus the project will indirectly incur negative impacts associated with greenhouse emissions.

Mitigation techniques include;
Mitigation measures relate to improving energy management and conservation practices.
Sub-meters and real-time energy monitoring equipment, timers, photoelectric cells, thermostats, etc. should be installed in the villas.
Install translucent shades and fluorescent lighting.
Pipe insulation, tank lagging (not asbestos!) and heat recovery systems should be installed wherever it is practical to do so.
Waste material (iron rich residue)
The tails at the end of the process, i.e. the iron rich residue is deposited in dumpsites around the plant, hopefully to be used I the future, for the recovery of iron. However they pose a hazard to the environment since they take up space that can be used for other uses in the society. Iron in large quantities can also affect the ecosystem surrounding the dumpsite if leached into the soil.

Mitigation techniques;
Design of a waste management plant
Use of landfills
Use the residue for construction of roads
Potassium dichromate
Potassium dichromate is a harmful substance and is toxic to humans, animals and plants if not handled properly.

Potassium dichromate is one of the most common causes of chromium dermatitis, chromium is highly likely to induce sensitization leading to dermatitis, especially of the hand and fore-arms, which is chronic and difficult to treat. Toxicological studies have further illustrated its highly toxic nature. With rabbits and rodents, concentrations as low as 14 mg/kg have shown a 50% fatality rate amongst test groups ( Sigma et al.,2011)Aquatic organisms are especially vulnerable if exposed.

As with other Cr (VI) compounds, potassium dichromate is carcinogenic and should be handled with gloves and appropriate health and safety protection. The compound is also corrosive and exposure may produce severe eye damage or blindness. Human exposure further encompasses impaired fertility, heritable genetic damage and harm to unborn children.

Mitigation techniques;
Proper handling techniques should be practised, whilst ensuring that employees in the workplace never come in direct contact with the salt.

Social impacts
During the plant operations many social activities takes place. Some of the activities are positive and some are negative. Some of these activities are illegal, for example drugs and prostitution. Prostitution leads to spread of Aids.

Mitigation techniques;
Aids awareness campaigns to the plant workers is required
Provision of social and entertainment facilities
9.2.4 Decommissioning impacts
Ghost sites creation
Upon the decommissioning of the plant the facilities will be left idle and it can pose a negative impact to the environment. The facilities can be used for different illegal dealing, or it can collapse from human or animal activities destroying the habitat
Mitigation techniques include;
Prior to plant decommissioning the facilities should be used for other purposeful projects
The responsible authority should consider the up keep of decommissioned plants
Unemployment increases
Prior to the decommissioning of the plant many workers will be left without any source of income to cater for their needs. The unemployment will result in many negative implications e.g. increase of crime rate.

Mitigation techniques include;
Workers should be trained for entrepreneur skills so as to generate income in cases of losing their jobs
Inception of other projects which foster employment
CHAPTER TEN: ECONOMIC ANALYSIS
10.0 Introduction
The purpose of this chapter is to demonstrate the economic and financial viability of the project using profitability indicators such as return on investment (ROI), payback period (PP), break-even (BE) point analysis, net present value (NPV) and internal rate of return (IRR). The chapter will cover assumptions on financing estimation of project cost (fixed cost and working capital), estimation of manufacturing unit cost, and calculation of annual net cash flows.

10.1 Estimation of capital investment
10.1.1 Direct costs
These represent material and labour involved in actual installation of complete facility.

Table 10.1 Bill of quantities for a ball mill
Material Description Quantity Unit cost Total cost
Mill shell High manganese steel 10tonnes $1500 $1500
Liners High manganese steel(110mm) 15 $20 $1800
Lifter bars High manganese steel(110mm) 18 $20 $2160
Steel balls Hardened high chrome(62mm) 208t/year $250 $52000
Sliding bearing Stainless steel 2 $1000 $2000
Mounting bearing Stainless steel 2 $1200 $2400
Cement – 80bags $10 $800
Bolts M33 120bolts $4.5 $540
Motor 550Hp 1 $8000 $8000
Diaphragm – 2 $500 $1000
GRAND TOTAL $95 700
Considering that there are two operational ball mills;
Total =$95 700 +$95 700
=$191 400
Table 10.2 Bill of quantities of a Forced circulation evaporator
Material Description Quantity Unit cost Total cost
Shell and tube heat exchanger Stainless steel 1 $2500 $2500
Flash vessel separator Stainless steel 1 $1500 $1500
Forced circulation pump Comes with motor 1 $4000 $4000
Thickener Stainless steel 1 $2200 $2200
Centrifuge – 1 $15000 $15000
Steam compressor – 1 $2500 $2500
Condenser – 1 $2050 $2050
Motor 1.5Hp 1 $1150 $1150
Cement – 40bags $10 $400
Support Carbon steel 10m2 $110 $1100
GRAND TOTAL $32 400
Table 10.3 Purchased equipment cost
Equipment Quantity Unit cost ($) Total cost ($)
Jaw crusher 2Hp 1 7500 7500
Conveyor belts 500m 12 6000
Cyclones 2 2500 5000
Storage tanks 5 1200 6000
Mixing tanks 3 1300 3900
Motors 3 1150 3450
Rectifiers 2 130 260
Reactors 2 4200 8400
Condenser 1 2050 2050
Spray dryer 1 5600 5600
Boiler 1 2250 2250
Recuparator 1 220 220
Leach tank 1 7840 7840
Sumps 2 700 1400
Vibro-feeders 4 360 1440
Valves 20 450 9000
Centrifugal pumps 7 1500 10500
Pressure filter 1 8000 8000
GRAND TOTAL 88 810
Total cost of equipment = $88 810 + $32 400 + $191 400
= $312 610
Cost in 2018=Cost in 2014 × cost index in 2014cost index in 2018Cost in 2018=$312 610 ×10241125=$284 545Purchased cost of equipment PCE=$284 545Total physical plant cost
The plant is handling fluids and solids. The following table shows the typical factors for estimation of project fixed capital cost. (Coulson and Richardson volume 6, 4th edition)
Table 10.4 Estimation of physical plant cost
Factor Use Value
f 1 Equipment erection 0.45
f 2 Piping 0.45
f 3 Instrumentation 0.15
f 4 Electrical 0.10
f 5 Buildings 0.10
f 6 Utilities 0.45
f 7 Site development 0.05
Total PPC = PCE (1 +f1 + f2 + f3 + f4 + f5 + f6 + f7 + f8 + f9)
Total PPC = PCE (1+0.45+0.45+0.15+0.10+0.10+0.45+0.05)
PPC = ($284 545 ×2.75)
=$782 499
Total fixed capital cost
Table 10.5 Estimation of fixed capital cost
Factor Use Value
f 10 Design and engineering 0.25
f 11 Contractor ‘s fee 0.05
f 12 Contingencies 0.1
Total fixed capital =PPC (1 + f10 + f11+ f12)
Total fixed capital = PPC (1+0.25+0.05+0.1)
=$1 095 499
Direct costs
Material and labour involved in actual installation of complete facility (70-85%) of fixed capital investment. The total direct costs is shown in the table below:
Table 10.6 Direct costs
Component Range % Chosen % Cost US$
Purchased equipment cost (PCE) 284 545
Installation of purchased equipment (20-50) of PCE 30 85 363.5
Instrumentation and controls(installed) (15-30) of PCE 25 71 136.25
Electrical (installed) (10-50) of PCE 30 85 363.5
Buildings (including services) (10-40) of PCE 35 99 590.75
Yard improvements (10-15) of PCE 13 36 990.85
Piping (installed) (10-80) of PCE 45 128 045.25
Land (4-8) of PCE 5 14 227.25
Total direct costs 805 262.35
10.1.2 Indirect cost
These include construction costs, rental of construction equipment and machinery. The total indirect costs are shown below;
Table 10.7 Indirect costs
Component Range % Chosen Cost US$
Engineering and supervision (5-30) DC 20 104 684.1055
Construction expense (5-20) DC 15 48 657.945
Contingency (5-20) DC 15 80 526.235
Contractors’ fee (2-15) DC 15 56 368.3645
Total 290 236.65
Fixed capital investment = direct costs + indirect costs
= ($805 262.35+ $290 236.65)
=$1 095 499
10.1.3 Working capital
Working capital allows 15 % of Total fixed capital.

Working capital =0.15×$1 095 499
=$164 324.85
10.1.4 Total capital investment
Total investment required = Total fixed capital + Working capital
= $1 095 499 + $164 324.85
=$1 259 823.85
10.2 Estimation of total production cost
Total production cost = Variable cost + Fixed charges + Plant overhead cost.

10.2.1 Fixed charges
These are normally between (15-35%) of fixed capital investment.

Depreciation
The annual depreciation rate for machinery and equipment ordinarily is about 10 percent of the tied-capital investment, while buildings are usually depreciated at an annual rate of about 3 percent of the initial cost.
Depreciation = (0.10 × $705 671.6) + (0.03×$99 590.75)
= $73 555
Insurance
Considering the production operation and the available protection facilities insurance of rate of 1% of the fixed-capital investment is used to calculate insurance required for the plant
Insurance = 1% of fixed capital investment
Insurance = 0.1 × $1 095 499
=$10 954.99
Maintenance
Given as 5% of fixed capital;
Maintenance = 0.05 × $1 095 499
=$54 774.95
Operating labour
Given as 10% of direct cost;
Operating labour = 0.1 × $805 262.35
=$80 526.235
Total fixed charges = $80 526.235 + $54 774.95 + $10 954.99 + $73 555
=$219 812
10.2.2 Variable costs
Raw materials
Raw materials to be purchased include potassium hydroxide, potassium carbonate.
Potassium hydroxide = $800 per tonne
Potassium carbonate = $700 per tonne
Amount of raw materials to be used per day
Potassium hydroxide = 1.1637t
Potassium carbonate = 0.06t
Number of operating days in a year = 365
Potassium hydroxide = 424.75t
Potassium carbonate = 21.9t
Cost of raw materials = (424.75t × $800) + (21.9t x $700)
= $355 100
Miscellaneous activities
Given as 10 % of maintenance;
Miscellaneous activities = 0.1 × $54 774.95
= $5477.495
Utilities
Given as 5% of direct costs;
Utilities = 0.05 × $805 262.35
=$40 263.1175
Total variable costs = $40 263.1175 + $5477.495 + $355 100
= $400 840.6
10.2.3 Plant overheads
Overhead expenses such as rent, interest and insurance which are related to the production capacity of a firm and not to its actual level of output.

Table 10.8 Plant overheads
Plant overheads Cost $
Storage facilities (3% of FCI) 32 864.97
Medical services (2.5% of FCI) 27 387.475
Safety and protection (1% of FC1) 10 954.99
Total 71 207.435
Total production cost = $71 2017.435 + $400 840.6 +$219 812
=$691 860.035
10.3 Production cost analysis
Assuming that 365 working days.

Annual production = 1.62t/day x 365days
= 591.3t/year
Total cost of production = $691 860.035
Production cost/tonne = $691 860.035591.3tonnes = $1170 /tonne
10.4 Profitability analysis
10.4.1 Profit margin
Average selling price for industrial grade potassium dichromate is $2000/tonne.

The profit margin = 20001170 ×100 = 170.9%
This percentage represents the profitability of the plant.

Total Income Generated = $2000/t x 591.3t
= $ 1 182 600
Gross Profit = Total Income Generated – Total Production Cost
= $1 182 600 – $691 860.035
= $490 739.965
Tax Paid = 15% of the Gross Profit
= 0.15 x $490 739.965
= $73 610.99475
Net Profit = Gross Profit – Tax Paid
= $490 739.965 – $73 610.99475
= $417 128.9703
10.4.2 Payback period
This is the time required for the cumulative net cash flow taken from start-up of the plant to equal the fixed capital investment made. It is given by:
PBP = Total capital investmentNet profit x 100
= $ 1 259 823.85$ 417 128.9703 x 100
= 3.02 years
10.4.3 Rate of return
The annual income from an investment expressed as a proportion (usually a percentage), of the original investment, is what is meant by rate of return.

Rate of Return = Net profitTotal capital investment x 100= $417 128.9703$1 259 823.85 x 100 = 33 %
10.4.4 Break-even point
It represents the sales amount in either unit or revenue terms that is required to cover costs. The main purpose is to determine the minimum output that must be exceeded in order to make profit.

Unit variable cost=Variable Cost K2Cr2O7 annual production=$400 840.6591.3t=$677.9/t Break even in units= Total Fixed CostsSelling price – Unit Variable Cost=$219 812$2000-$677.9=166.2597?166Break even in dollars=Total Fixed costsUnit Price-Unit Variable CostUnit Price=$219 812$2000-$677.9$2000=$332 519 10.4.5 Net present value
Cash flow statement attached in the appendix section.

Discount rate is estimated at 18% and an initial total capital investment of $1 259 823.85 was used.

Table 10.9 Net present value cash flow
PV of Cash Flows ($)
Year Cashflow Calculation PV Factor Cash Flow * PV Factor
2019 $691 860.04 (1/1+0.18) ¹ 0.8475 $586 351.38
2020 $703 297.3 (1/1+0.18) ² 0.7182 $505 108.12
2021 $931 280.95 (1/1+0.18) ³ 0.6086 $566 777.59
2022 $1 221 240.05 (1/1+0.18) ? 0.5158 $629 915.62
2023 $1 508 353.7 (1/1+0.18) ? 0.4371 $659 301.40
Total $2 947 454.11
Net present value NPV=Total PV Cash Inflows-Capital Investment NPV=$2 947 454.11-$1 259 823.85=$1 687 630.26CHAPTER ELEVEN: CONCLUSION AND RECOMMENDATIONS11.0 Conclusion
The manufacture of potassium dichromate by electrochemical oxidative decomposition of chromite slag and low-grade chromite in a concentrated alkaline solution has thus been confirmed to be feasible.

From the results obtained from laboratory experiments, the optimum operating conditions for the electrochemical oxidation of low grade chromite ore and slag is at 60 wt. % alkali, current density of 750 A/m2, ore to slag ratio of 3:1, airflow of 1L/min, stirring speed of 700rpm, a particle size distribution of less than 75µm, reaction temperature of 160? and a residence time of 4 hrs. At these conditions the expected recovery is 89.67%. This process proves to be an efficient way of using EAF slag.

An economic analysis carried on of the project showed that with an initial capital investment of $1 259 823.85, a payback period of 3.02 years and a rate of return of 33% is realised.

11.1 RecommendationsThe following areas of further study, research and considerations are being recommended for a complete and successful implementation of process;
The effects of a binary sub- molten system, with a combination of potassium hydroxide, potassium nitrate and water instead of potassium hydroxide alone should be looked into.

The role of potassium carbonate as a catalyst should be looked into further and other alternatives researched on.

Research should be done into the most effective feed rate of air during the reaction since an assumption was used.

The effect of carbothermic reduction of chromium oxide on ore and the resulting recovery if such a type of ore is used.

There is need for simulation or prototyping of this process to check for over sighted problems that might arise during the operation of this process.

A maintenance contract should be in place with a licensed onsite professional to assure the proper operation and maintenance of plant equipment and instruments
Mining is a hazardous process to the environment as it is. Ways of utilising the iron rich waste produced should be looked into, for example using the waste as tar for roads. If not there is the risk of people and civilians who might inhale the iron rich dust, could develop retinitis, benign pneumoconiosis and even lung cancer in extreme cases.

APPENDIXAppendix 1 Physiochemical properties of oxygen
Alkali concentration (wt.%) ?+/- *a 105COXYGEN *b
Mol/kg 105DOXYGEN *c
5 1.53 83.08 1.5937
13.5 3.06 42.90 1.0115
23 8.15 16.81 0.6847
31.61 20.1 7.141 0.5226
40.70 70.1 2.176 0.4029
50.65 230 0.722 0.3102
*a =Activity coefficients of dissolved oxygen in KOH solutions, date taken from
(Shoor et al. 1969).

*b= Dissolved oxygen content in the KOH solutions, data calculated from (Shoor et al.

1969).

*c =Oxygen diffusion coefficient in the KOH solutions, data taken from (Tham et al.

1970).

Appendix 2 Bond work indices
Soft ores 6-10KWhr/tonne
Medium ores 10-18KWhr/tonne
Hard ores 18-36KWhr/tonne
Data from Metskill by R Sands et al 2010
Appendix 3 Ball mill design factors for calculation

Appendix 4 Cash flow statement

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