Beyond the Blast Furnace [1 ed.] 9780849366765, 9781315138220, 9781351464239, 9781351464222, 9781351464246

Table of contents :

Introduction. Kinetics of Iron Oxide Reduction. Coal Based DR Processes Using Rotary Kilns. Coal-Based DR Processes Using Shaft Furnaces. Gas-Based DR Processes. Current Status and Future Outlook of DR. Smelting Reduction Processes. Important Features of SR Processes. Alternative Avenues of Ironmaking. Use of DRI/HBI and SR/MBF Hot Metal in Iron- and Steelmaking. The Future Beyond the Blast Furnace. References. Index.

Citation preview

Beyond the

BLAST FURNACE

Beyond the

BLAST FURNACE Amit Chatterjee Senior Technical Advisor The Tata Iron ond Steel Co., Ltd. Jamshedpur, India

Boca Raton

CRC Press Ann A rbor London

Tokyo

Library of Congress Cataloging-in-Publication Data Chatterjee. Amit. Beyond the blast furnace / Amil Chatterjee, p. cm. Includes bibliographical references and index. ISBN 0-8493-6676-3 1. Iron— Metallurgy. 2. Direct reduction (Metallurgy). TN706.5.C47 1993 669'.141—dc20

3. Smelting.

I. Title. 93-13331

CIP This book contains information obtained from authentic and highly regarded sources. Reprinted material is quoted with permission, and sources are indicated. A wide variety o f references are listed. Reasonable efforts have been made to publish reliable data and information, but the author and the publisher cannot assume responsibility for the validity of all materials or for the consequences of their use. Neither this book nor any part may be reproduced or transmitted in any form or by any means, electronic or mechanical, including photocopying, microfilming, and recording, or by any information storage or retrieval system, without prior permission in writing from the publisher. CRC Press, Inc.'s consent does not extend to copying for general distribution, for promotion, for creating new works, or for resale. Specific permission must be obtained in writing from CRC Press for such copying. Direct all inquiries to CRC Press, Inc., 2(X)0 Corporate Blvd., N.W., Boca Raton, Florida 33431. © 1994 by CRC Press, Inc. No claim to original U.S. Government works International Standard Book Number 0-8493-6676-3 Library o f Congress Card Number 93-13331 2 3 4 5 6 7 8 9 0

DEDICATION This book is dedicated to four persons who have molded my life, each in his or her own way, who are all very dear to me. They are my father. Dr. A. B. Chatterjea, himself a metallurgist who worked in the area of alternative processes and who became my the inspiration and ideal role model, my mother, Mrs. Usha Chatterjee, whose tender care and affection made me what I am today, my wife, Jenny, who has showered me with her love for over twenty-five years, through the good times and the bad, and, my daughter, Nimilita (Mia), whose omnipresent ‘Tapa don't preach (to me!), until...(you write the book)", drove me on and on. I salute them all and can only say thank you so much. I am delighted to have this opportunity to record my appreciation for your contributions to this book.

Amit Chatterjee, D.Sc., Ph.D. Jamshedpur, India October 1992

FOREWORDS Before addressing the subject of this book, it seems necessary to say a few words about the Chatterjees, both Dr. Amit Chatterjee and his father, Dr. A. B. Chatterjea, since metallurgy is a tradition in this family, as well as for India, with its long history in the art of ironmaking. At any rate, it is a fact that the Chatterjee family has a long tradition in modem iron- and steelmaking technology. This certainly makes it appropriate that a book such as this has been written by Dr. Amit Chatterjee. Regarding my relations with this family of Indian metallurgists, and especially with Dr. Amit Chatterjee, 1 am reminded of my first trip to India, at the end of the 1950s and the beginning of the 1960s, when Amit was a young boy. Tisco was, at that time, the “big steel plant," practically the only large one, with several smaller plants around the country. Now the panorama has changed, and is still changing, but Tisco remains a good example of an efficient plant. Practically all the processes described by Dr. Amit Chatterjee in this book are in use in the world today — all the variations of the blast furnace, including the mini blast furnace, direct reduction, and the potential of the smelting reduction processes. The book itself, then, is timely: i.e., it appears at a time when the classical ironmaking process, the blast furnace with accompanying agglomeration of iron ores and coal coking plants, is developing at the same time that new ironmaking processes are appearing in various forms. In other words, we are witnessing a time of both evolution and revolution. The evolution is illustrated by the fact that classical ironmaking, using the blast furnace process, has been developed to a kind of perfection by, in my opinion, a combination of factors. These include purely technological improvements in materials and processes used to build the blast furnace itself, the stoves as well as corresponding agglomeration, and coke oven facilities, and scientific achievements in knowledge about and understanding of the blast furnace, which now enable us to operate it in the most efficient manner. The revolution is mainly explained in this book by the two “new" technologies, direct reduction and smelting reduction. Direct reduction is now becoming an established technol­ ogy. Although it is quite old, older than blast furnace technology, it had gradually disappeared, and it has only made a “comeback" in the last twenty years; as described in this book, it is now playing an increasing role in modem iron- and steelmaking. It is on this subject that I remember many meetings and discussions with Dr. Amit Chatterjee, during which we discussed direct reduction. We met in India, when he was developing operations in the Ipitata Direct Reduction plant, and in many other places such as Rio de Janeiro for the IIS I annual meeting, in Venezuela for a very interesting seminar, and in Jamshedpur, in India. Smelting reduction, again, is not new; it dates back, like direct reduction, to ancient technology, when iron was produced as solid or more or less liquid by various methods. Then, outside the shaft furnace (which ultimately became the conventional blast furnace), several processes were developed to produce liquid hot metal, especially electric furnaces and rotary kilns. After many unsuccessful attempts, new processes are now appearing. We must be grateful to the author for describing and explaining the new technological processes for producing liquid hot metal and, perhaps, liquid steel directly in the furnace, which is one of the dreams of all metallurgists! I wish this book all success.

Jacques E. Astier, Consultant Paris May 1992

It gives me great pleasure to write a foreword for this book, whose title, Beyond the Blast Furnace, is significant in the sense that it is not After the Blast Furnace, since the blast furnace will maintain its supremacy while the alternative technologies, including direct reduction and smelting reduction, will continue to make parallel impressive growth, particularly in develop­ ing countries. Direct reduction technologies have indeed taken root, mostly in the developing countries that provided the cradle for their remarkable progress over the decades. Smelting reduction technologies, the craze of the 1980s, are slowly finding their feel in the 1990s. These alternative technologies provide a fascinating and challenging spectrum, placing a high premium on technical ingenuity. Coming from the pen of Dr. Amii Chatterjee, with his mastery of the subject, intellectual brilliance, and pragmatic analysis, this publication is a classic. In the midst of myriad claims for wondrous smelting technologies, many of which have fallen by the wayside over the years. Dr. Chatterjee has done a remarkable job of analyzing and highlighting those that hold the most promise, while others, still smoldering, may find few takers. Dr. Chatterjee’s task has not been easy. With a flair for incisive, in-depth review, and a solid theoretical background. Dr. Chatterjee has indeed put in proper perspective the claims of those who state, "Après moi le deluge”. Dr. Chatterjee has also covered the subject of mini blast furnaces, with its many global applications, such as in Brazil and, soon, in India. Additionally, Dr. Chatterjee has drawn upon his immense practical experience in developing, innovative R&D technologies, such as TDR (Tisco Direct Reduction) technology, upon which basis a commercial plant is currently operating very successfully in India. This publication is a masterpiece; il fulfills a long-felt need of discerning readers, metal­ lurgists, and busy executives.

B. R. Nijhawan, Senior UNIDO Consultant Kokomo, USA April 1992

PREFACE This book encompasses the entire gamut of processes which aim to produce iron units without going through the classical blast furnace and the accompanying cokemaking and sintermaking facilities. Thus, the book covers direct reduction, smelting reduction, mini blast furnaces, iron carbide manufacture, and some attempts at direct steelmaking, both with and without the use of direct reduced iron. The book begins by covering the theory and mechanism involved in reduction of iron oxide employing solid and gaseous reductants. It then goes on to describe features of coal and gas based DR processes which are in commercial operation. Subsequent chapters of the book deal with some of the suggested processes of smelting reduction, including the COREX process, which is the only smelting reduction process which has been commercialized so far. The last few chapters of this book discuss other alternative processes of making iron and describe the use of sponge iron in electric furnace steelmaking, both using hot DRl as well as conventional charging, in which DRI is charged cold. The book concludes with a chapter which examines the future of alJ the alternative processes and comes to the conclusion that these aJtemative processes wilJ supplement hot metal production by highly efficient classical blast furnaces in the years to come. Perhaps the most important chapters in this book are Chapters 3 and 5 , which detail various aspects of coal- and gas-based direct reduction, respectively, and Chapter 7, which deals with the smelting reduction processes available at present. 1 hope the readers will find this book of interest. It has been prepared over a period of ten years, chiefly with the assistance of Dr. P. Basu and Mr. B. D. Pomdey, erstwhile colleagues in the R&D Division of Tata Steel.

Amit Chatterjee, D.Sc., Ph.D. Jamshedpur, India September 1992

ACKNOWLEDGMENT A book of this nature, covering different processes that supplement the blast furnace, can never be credited to one individual. This book is a product of synergistic team efforts in which many individuals made contributions. While it is not possible for me to acknowledge the help rendered by all individually, I would like to thank the following people. Dr. Projjal Basu, a colleague of mine in the R&D Division of Tata Steel, helped in the preparation of the basic manuscript and painstakingly carried out instructions given to him. This book owes a lot to him. Similarly, Mr. B. D. Pandey, also of the R&D Division, Tata Steel, extended his assistance in preparing some of the material used in this book. I am indebted to two internationally renowned authorities in the field. Dr. B. R. Nijhawan and Dr. J. E. Astier, who have known me since my childhood, for having agreed to write the two forewords for this book. I am sure that their contributions have added luster to Beyond the Blast Furnace. Perhaps my personal exposure to the alternative processes of ironmaking began with my association in the development of the Tisco Direct Reduction (TDR) process of sponge iron making, which has been acknowledged both in India and overseas as one of the most sustained research efforts undertaken in the iron and steel industry. I would like to thank all those who worked with me in the TDR Pilot Plant and then in Ipitata Sponge Iron Limited for their cooperation and assistance for over two decades. I would also like to acknowledge with gratitude the support extended by the management of Tata Steel, Mr. R. H. Mody and Dr. J. J. Irani, in particular, in this endeavor. Secretarial assistance in the preparation of this book was provided by Mr. Prakash Rao and Mr. K. H. Kartha, who have tirelessly worked in my office and have endured the many changes and corrections introduced at various stages during the preparation of the manuscript. The help rendered by the R&D draftsmen in preparing the figures and by Mr. Ramesh Singh, another colleague in R&D, in overall coordination, is also acknowledged with thanks. I would like to close by paying homage to The Almighty for having given me the strength to complete this task, started almost a decade ago. CRC Press, Inc. deserves kudos for the publication. I hope the readers will derive some benefit from this book.

Amit Chatterjee, D.Sc., Ph.D.

ABOUT THE AUTHOR Amit Chatterjee, D.Sc., Ph.D., Senior Technical Advisor and until recently Deputy General Manager (Research and Development), Tata Steel, Jamshedpur, India, graduated in Metallurgical Engineering from the Bañaras Hindu University (BHU), India, and obtained his Ph.D. and D.l.C. from the Imperial College of Science and Technology, London, in 1970. He joined Tata Steel in 1972, following employment at British Steel and in Thyssen, Germany. After rising to the post of Joint Director (Research and Development), he was assigned to Ipitata Sponge Iron Limited in 1982 to serve five years as its Managing Director. Ipi tata is India’s first coal-based direct reduction plant employing the Tisco Direct Reduction process, which was developed through Dr. Chatterjee’s sustained efforts spanning almost two decades. He returned to Tata Steel in 1987 as Director (Research and Development). Dr. Chatterjee has over 270 publications to his credit, many in international journals and symposium proceedings. He was awarded a higher doctoral degree. Doctor of Science (Engineering), by the University of London for his outstanding work on coal-based direct reduction and oxygen steelmaking. He has received the National Metallurgist Award given by the Government of India, the SAIL Gold Medal, the Visvesvaraya Gold Medal (both from the Institution of Engineers, India), and the Bhoruka Gold Medal of the Indian Institute of Metals; he is a Fellow of the Institute of Metals, London, and is also a Chartered Engineer of the Engineering Council, England. He is acknowledged as one of the foremost technical authori­ ties on process metallurgy of iron- and steelmaking, and has coedited books on blast furnace ironmaking, steelmaking, continuous casting, and coal carbonization.

TABLE OF CONTENTS Chapter I Introduction ............................................................................................................................ I References ................................................................................................................................ 6 Chapter 2 Kinetics of Iron Oxide Reduction ........................................................................................ 7 I.

Reduction of Iron Oxides in Systems Employing Solid and Gaseous Reductants ...... 7

H.

Kinetics of Reduction of Iron Oxide ............................................................................. 9 A. Mass Transfer Through the Stagnant Gas Film ................................................ I 0 B. Mass Transfer Between the Solid Oxide Surface and the Reaction Interface ............................................................................................... I 0 C. Chemical Reaction of Oxygen Removal at the Interface .................................. I 0 D. Nucleation and Growth of Iron .......................................................................... 11

m.

Mechanism of Iron Oxide Reduction in Coal-Based Processes ................................. 11

IV.

Mechanism of Iron Oxide Reduction in Gas-Based Processes .................................. 13 A. Reduction of Dense Iron Oxide ......................................................................... 13 B. Reduction of Porous Iron Oxide ........................................................................ 15

V.

Carbon Deposition and Its Benefits ............................................................................ 17

References .............................................................................................................................. 18 Chapter 3 Coal Based DR Processes Using Rotary Kilns ................................................................. 19 I.

Process Technology in General ................................................................................... 19

U.

Specific Rotary Kiln OR Processes in Commercial Usage ........................................ 22 A. SL/RN ................................................................................................................. 22 B. CODIR ................................................................................................................ 23 C. ACCAR ............................................................................................................... 28 D. DRC .................................................................................................................... 31 E. TOR .................................................................................................................... 32 F. General Comments on Rotary Kiln OR ............................................................ 36 G. Processes Using Plant Wastes in Rotary Kilns ................................................. 37

Ill.

Process Analysis .......................................................................................................... 38

IV.

Raw A. B. C.

Material Characteristics ....................................................................................... 40 Iron Oxide Feed .................................................................................................. 40 Noncoking Coal for Direct Reduction .............................................................. .43 Fluxes .................................................................................................................. 49

V.

Ring A. B. C.

Formation in Rotary Kilns .................................................................................. 49 Softening Behavior of Iron Oxide .................................................................... .49 Kiln Operating Conditions ................................................................................. 51 Effect of Composition ........................................................................................ 51

VI.

Flow A. B. C. D.

of Materials ......................................................................................................... 52 Bed Movement ................................................................................................... 52 Effect of Operational Variables ......................................................................... 55 Effect of Kiln Geometry .................................................................................... 56 Effect of Build-ups Within the Kiln .................................................................. 60

VII. Process Control Parameters ......................................................................................... 61 A. Feed Rate and Proportion of Raw Materials in the Charge Mix ...................... 61 B. Kiln Temperature Measurement and Control .................................................... 62 C. Control of Gaseous Atmosphere ........................................................................ 63 D. Gas Pressure Inside the Kiln .............................................................................. 63 E. Kiln Rotation ...................................................................................................... 63 F. Inclination of the Kiln and Retention Time ...................................................... 63 G. Sponge Iron Discharge Temperature ................................................................. 63 H. Amount of Char in the Kiln Discharge Product ................................................ 63 I. Waste Gas Temperature and Composition ........................................................ 64 References .............................................................................................................................. 64 Appendices ............................................................................................................................. 65 Chapter 4 Coal-Based DR Processes Using Shaft Furnaces ............................................................. 81 I.

KM Process .................................................................................................................. 8 I

H.

MIDREX Electrothennal Direct Reduction Process ................................................... 84

UI.

Shaft Furnaces Using Gasified Coal ........................................................................... 85 A. Coal Gasification Processes ............................................................................... 85 B. Evaluation of Gasification Processes Based on the Requirements of a DR Unit ............................................................................... 87 C. DR Process Flowsheet Based on Coal Gasification .......................................... 89

References .............................................................................................................................. 90 Chapter 5 Gas-Based DR Processes ..................................................................................................... 91 I.

Retort Processes ........................................................................................................... 91 A. HyL ..................................................................................................................... 91 B. HOGANAS ......................................................................................................... 91

n.

Process Using a Fluidized Bed .................................................................................... 94 A. FIOR ................................................................................................................... 94 B. HIB ..................................................................................................................... 94

Ill.

Shaft A. B. C. D.

E. F. G. H.

Furnace Processes .............................................................................................. 95 WIBERG ............................................................................................................ 95 PLASMARED .................................................................................................... 95 ARMC0 .............................................................................................................. 97 PUROFER .......................................................................................................... 97 NSC .................................................................................................................... 99 HyL III .............................................................................................................. 100 MIDREX .......................................................................................................... 103 AREX ............................................................................................................... 110

References ............................................................................................................................ I 12

Chapter 6 Current Status and Future Outlook of DR .................................................................... 115 References ............................................................................................................................ 126

Chapter 7 Smelting Reduction Processes .......................................................................................... 127 l.

Introduction ................................................................................................................ 127

Il.

Historical Background ............................................................................................... 127

Ill.

Process Classification ................................................................................................ 128 A. Shaft Furnace Processes ................................................................................... 128 B. Electrical Processes .......................................................................................... 128 C. Converter Processes ......................................................................................... 128

IV.

Descriptions of Different Processes .......................................................................... 128 A. COREX ............................................................................................................. 128 B. Sumitomo's SC Process ................................................................................... 136 C. Kawasaki Smelting Reduction Process ............................................................ 137 D. INRED .............................................................................................................. 140 E. ELRED ............................................................................................................. 143 F. COMBlSMELT ................................................................................................ 145 G. PLASMASMELT ............................................................................................. 147 H. KRUPP-COIN .................................................................................................. 149 l. PClG for Smelting Reduction .......................................................................... 150 J. Hlsmelt Process ................................................................................................ 154 K. Reactor Steelmaking ......................................................................................... 156 L. HyL-AISI Direct Steelmaking ......................................................................... 159 M. FLPR ................................................................................................................. I59 N. DIOS ................................................................................................................. I60 0. Other Developments ......................................................................................... 161

V.

Current Status and Future Outlook of Smelting Reduction ...................................... l62

References ............................................................................................................................ 164

Chapter 8 Important Features of SR Processes ............................................................................... 167 I.

Process Categorization ............................................................................................... 167

Il.

Kinetic Aspects of Smelting Reduction .................................................................... 169 A. Reduction by Solid Carbon .............................................................................. 170 B. Reduction by Carbon Dissolved in Liquid Iron .............................................. 171 C. Reduction by Carbon Monoxide ...................................................................... 171

UI.

Importance of Slag Foaming in Smelting Reduction ................................................ 174

IV.

Practical Considerations ............................................................................................. 175

References ............................................................................................................................ 179 Chapter 9 Alternative Avenues of lronmaking ................................................................................ 181 I.

INMETCO Process .................................................................................................... 181

U.

MIDREX FASTMET Process ................................................................................... 184 A. Process Description .......................................................................................... 184 B. Raw Materials and Energy ............................................................................... 186 C. Process Applications ........................................................................................ 186 D. Economics ........................................................................................................ 187

Ill.

Iron Carbide Manufacture ·······················o········o·············o··········o········o············o·•·o······ 189

IV.

lronmaking in Mini Blast Furnaces ......... o......... o• .. o.. o•·· ........... 194 A. Special Features of MBF .............. oo········o··········o························o·o····•ooo·······ooo 194 B. Charcoal versus Coke MBF o···········o··············o••o·····o····················o···············o··· 195 C. World Scenario ···········································o··············o···············o··o····•ooo•o········· 196 000···· ••••••••••••••••••••

References .................

0 •••••••••••••••••••••• 0 •••••••••••••••••• 0 •••• 0 •••• 0 ••••••••••••••••••••••••••• 0

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••••

••••••••••••••••••••••••••

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Chapter 10 Use of DRI/HBI and SR/MBF Hot Metal in Iron- and Steelmaking ·················o········ 20 I I.

Use of SR/MBF Hot Metal and DRI!HBI in Steelmaking ............................

Il.

Use of Sponge Iron in Electric Arc Furnaces .................................................. 201 A. Recent Developments in Electric Arc Furnaces with Particular Reference to the Use of Sponge Iron ........................ 203 B. Charging and Melting Practices ....................................................................... 205 C. Furnace Performance with Continuously Charged DRI .................................. 207 D. Features of DRI Usage in EAFs ...................................................................... 213

0 ••••••••••

0

201

••••••••

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UI.

Use of Hot DRI in Steelmaking .............................................

IV.

Replacement of Scrap by Sponge Iron in LD Steelmaking ..................................... 216

0 ••••••••••••••••••••••••••••••••••

214

V.

Advantages of Using HBL/DRJ in Steel Ladles ........................................................ 220

VI.

Use of Sponge Iron in Open Hearth Furnaces .......................................................... 220

VII. Use of DRI/HBI in Foundries ................................................................................... 220

A. B. C.

Sponge Iron Usage in Cupolas ........................................................................ 221 Use of DRIIHBI in the Induction Furnace ...................................................... 223 Use of DRJ for Blast Furnace Lronmaking ...................................................... 223

VIII. Reoxidation of Sponge Iron ....................................................................................... 224

References ............................................................................................................................ 227 Appendix .............................................................................................................................. 228 Chapter 11 The Future: Beyond the Blast Furnace .......................................................................... 231 References ............................................................................................................................ 236 lndex .................................................................................................................................... 237

Beyond the

BLAST FURNACE

Chapter 1

INTRODUCTION The classical blast furnace is still the principal means of hot metal production. In fact, over 95% of the total iron in the world today is produced using this process, as was the case a century ago. Figure 1.1 presents the percentage contribution of hot metal in steelmaking across the globe in 1990.* Such a long and sustained track record of blast furnaces arises out of the continuous developments in design, such as increases in capacity; alterations in profile; improved burden distribution systems like movable throat armor and Paul Wurth (PW) top; high top pressure operation, etc.. It also reflects upgrades of the process technology, such as the use of higher hot blast temperature, oxygen enrichment of the blast, auxiliary fuels including coal injection, prepared burdens, automatic measurement and control systems, and artificial intelligence systems, resulting in a significant increase in productivity, a decrease in coke rate, and an improvement in the quality of hot metal in classical blast furnaces. It is interesting to note that, in the year 1990, 527 million tons of hot metal were produced by 541 blast furnaces, which means an average production per furnace of almost 1 million tons of hot metal,' although many of the new furnaces constructed between 1986 and 1991 were either “very large” or “very small” — the latter comprising the so-called mini blast furnaces (Figure 1.2).' There are, however, certain limitations inherent in the blast furnace process. The process depends on metallurgical coal (required for the manufacture of blast furnace grade coke), it is an economically viable process only when it produces relatively large quantities (around 2 million tons per annum, or mtpa), and it requires elaborate supporting facilities such as raw materials handling and preparation systems, a sinter plant, coke ovens, a gas cleaning system, etc. These factors make the conventional blast furnace highly capital-intensive and often beyond the reach of smaller manufacturing units. Since blast furnace ironmaking is carried out in a single reactor, intrinsically different unit operations, like counter current heat and mass exchange, reduction by gases, reduction by solid carbon, high temperature gas generation, and finally, smelting and liquid drainage, all occur together. This means that the only way of ascertaining the efficiencies of the individual processes occurring within the blast furnace, which is a semicontinuous reactor, are periodic analysis of hot metal and slag and continuous gas analysis using a variety of probes, even within the burden in some cases. Although attempts have been made recently to introduce these probes, particularly in the furnace stack area, when corrective measures have been taken to alter the process parameters based on the data gathered from such probes, the responses to these measures could only be assessed after several hours. Therefore, the blast furnace is a reactor with a certain amount of mystique attached to it, and though this may lend it a romantic air of mystery, it does not make it a controlled reactor with high productivity and “hit rate”. The limitations imposed by concurrent reduction and smelting in a blast furnace have led to renewed interest in alternative forms of ironmaking which can be more readily linked to fast, high-productivity steelmaking processes. These alternatives include direct reduction (DR), smelting reduction (SR), iron carbide manufacture, and of late the so-called mini blast furnaces (MBFs), which really do not constitute an alternative system in the classical sense but form an extension of normal blast furnace ironmaking into the hitherto uncharted territory of small scale operation. It must be mentioned at this stage, to provide perspective, that while the blast furnace process is fully established, many of the alternative systems still exist only on paper, and to this extent it is difficult to make direct comparisons of the broad process systems.

Beyond the Blast Furnace

FIGURE 1.1. Percentage consumption o f hot metal, scrap, and sponge iron in the world in 1990 at a charge o f 1240 kg/t of crude steel.

2-4

4-é

6 -a

8-10

10-12

12-14

14-16

Hearth diameter, m

FIGURE 1.2.

Blast furnaces newly built and shut down during the period 1986 to 1991.

Because all metallurgical reduction processes occur at high temperatures, they consume substantial amounts of energy. In the blast furnace process, the energy is supplied by the hot blast, the combustion of coke, and increasingly, as shown in Figure 1.3, by direct coal injection through the tuyeres. ‘ All the alternative processes aim to eliminate or reduce the energy supply via coke. Thus, for an appreciation of alternative processes, it is necessary to take stock of the available energy resources of the world.^ As shown in Figure 1.4, coal represents the largest reservoir of energy still available to mankind. As a result of the continually dwindling reserves of oil and gas, coal is expected to meet an increasing share of metallurgical energy require­ ments in future. Furthermore, oil and gas reserves are concentrated only in some regions, whereas coal is more evenly distributed world wide. At the current rate of usage, it is estimated that the resource life of natural gas would be around 60 years, of oil only 33 years, and of coal 225 years. Oil and natural gas should, therefore, be used for the production of high value products, as far as possible, and coal should form the basis of all the alternative ironmaking processes that stretch “beyond” the blast furnace.

Introduction

FIGURE 1.3.

Blast furnaces of different size operated with coal injection in the world (as o f November 1991).

tO

*^ ^

FAR EAST WITHOUT P.R. CHINA

/ ^

AUSTRALIA/

S

/ \

ZEALAND

W.1

□

EXPLOITATION LIFE TIME

COAL

n atural

oil

GAS

FIGURE 1.4.

Available energy resources in the world.

Furthermore, it should be understood that energy accounts for about 20 to 25% of total manufacturing costs in steelmaking, and 65% of this cost is for the coal and coke consumed at the ironmaking stage. The contributions from fuel oil, natural gas, and electric power have risen gradually during the last two decades, but of late, the accent is again back on coal primarily because of the comparative world prices and relative availabilities of these forms of energy. Figure 1.5 clearly illustrates how the global market price of coal has changed only

Beyond the Blast Furnace

Year

FIGURE 1.5.

Fluctuations in the prices o f crude oil and coking coal in the world.

slightly since 1978-1979 compared with crude oil, which experienced major price swings even as late as 1991. It can thus be stated with some degree of certainty that in the long run, coal will continue to be less expensive and its price will certainly be more stable than that of the other forms of energy. Use of coal and similar solid reductants has therefore increased in the new technologies being developed and adopted in the iron and steel industry all over the world. In some areas, where natural gas is abundant, it automatically becomes the preferred source of energy, but such cases are infrequent compared with the use of coal. Vagaries in the scrap market and an anticipated shortfall in its supply in the coming years have given impetus to the development of DR processes for the production of sponge iron and of the upcoming technique of using natural gas (or coal gas) for the manufacture of another scrap substitute, iron carbide. Production of steel using Direct Reduction-Electric Arc Fur­ naces (DR-EAF) has now become a firmly established process which is being adopted increasingly all over the world.^ Sponge iron, or direct reduced iron (DRi) as it is also often called, has also been used in many other steelmaking processes, including oxygen steelmak­ ing, which is by far the predominant means of producing steel. The past 10 years have thus witnessed major structural changes in the steel industry — an increase in essentially solid charge-based steelmaking (e.g., the Electric Arc Furnace and the recent Energy Optimization Furnace) and the overwhelming popularity of continuous casting being two key factors. At the same time, paradoxically, a major problem facing the steel industry all over the world, including the developing countries, has been inadequate availability of high quality scrap. This has been accentuated by the ever-increasing adoption of continuous casting. While the technology for making quality steels requires large quantities of prime quality scrap, continu­ ous casting improves energy utilization and reduces the generation of in-plant scrap. To overcome this problem, DR process technology, based either on coal or natural gas, has come of age, which has led to the increasing use of DRI — a term which will also be used in this book to denote sponge iron — in steelmaking. While the DR industry has reached the “adolescent” stage for supplying solid metallics, SR is a “newborn”; as it endeavors to become a complementary source of liquid hot metal and iron carbide (a source of iron units in the solid form), it is still in the “embryonic” stage. The focus of interest in liquid iron production has always been on improving the efficiency and lowering the cost of production. Keeping the inherent limitations of the conventional blast furnace in mind, a unique set of criteria has been formulated,"^ which should be fulfilled by any

Introduction smelting reduction process if it is to be a serious alternative to the blast furnace for producing hot metal in small tonnages. These include: •

Use of fine ore concentrates or powdered iron ore directly, without any prior agglom­ eration, Use of less expensive fuels, like coal, lignite, etc., without coking. No physical or chemical limitations either on the reductant or on the iron oxide feed. Full control of the raw material feed and ease of removal of the end product, as well as of the outgoing gases, A completely stirred system with immediate response to changes in process parameters. Productivity index (t/m^ of reactor volume per unit time) at least five times that of the blast furnace. Low investment cost per unit of iron produced at a smaller scale of operation. Ability to meet the more stringent environmental standards which have been formulated in recent years and are the forerunners of even stricter regulations to follow in future.

Both SR processes and MBFs can fulfill many of these stipulations. In developing countries like India and Brazil, which not only have limited capital availability and limited demand for steel products but which are also deficient in suitable grades of coking coal for use in blast furnaces, MBFs are required to augment conventional ironmaking, including iron produced through SR processes in the future. The blast furnace will no doubt continue to contribute the lion’s share to the steelmaking capacity of large integrated steel plants, even in the foreseeable future, but the use of small units is bound to mushroom, especially in catering to local markets. In India for example, small electric furnaces of 5 to 25 t capacity in the mini-mill sector account for almost 25 % of the crude steel capacity of the country. Based on some recent projections of the future steel demand and firm modemization/ex pans ion programs, both of the existing integrated steel plants and the mini-steel sector, a gap of about 5 million tons of steel is anticipated in India by the turn of the century. This will have to be met by installing new plants, preferably dispersed throughout the country, to meet the needs of local markets.^ The alternative methods of ironmaking would merit serious consideration under such circumstances. The technology to be adopted in such new mini or midi steel plants spread throughout the world would primarily depend on energy considerations, flexibility of operations with different raw mate­ rials, and the unit capacity at which the process would be economically viable. Unlike SR, which produces liquid metal, DR includes a family of processes in which iron ore or pellets are reduced in the solid state and oxygen is removed by either solid or gaseous reducing agents. In these DR processes, reformed natural gas or noncoking coal is mainly used as the reductant and is the primary source of energy. The basic principles involved in DR are shown schematically in Figure 1.6. It must be remembered that the final product is a solid which must be melted during steelmaking in a manner similar to scrap. For gas-based direct reduction, the types of reactors used are fluidized beds, retorts, and shaft furnaces, while for coal-based DR, rotary kilns are used almost exclusively. A large number of processes are available today for DR,^ and can be grouped as follows:• • • • • • •

Coal-based processes using rotary kilns, Coal-based processes using shaft furnaces (which are less popular), Processes using steel plant wastes as the primary oxide feed to rotary kilns, Continuous processes in a shaft furnace using reformed natural gas as the reductant (which are most common), Retort processes using gas or coal as the reductant, Gas-based processes using a fluidized bed.

Beyond the Blast Furnace Shaft- furnace

Reforf

Pellets Gas for Reduction

Gas for Preheating

Lump ore

Cooling

Natural Gas

Lump Ore

Gas

Steam ~

Reformer Natural Gas — «

Gas reformer

| Off Gas

Steam or Recycle Gas or Gasified coal

DRI Reduction Gas

Off Gas

1st Stage 2nd Stage 3rd Stage

FIGURE 1.6.

Process routes for D R I production.

Different aspects of DR-SR technologies, iron carbide manufacture, and MBFs will be described in subsequent chapters in order to provide a thorough understanding of the process fundamentals, operating parameters, constraints, and scope of applicability, particularly in countries which are deficient in coking coal.

REFERENCES 1. Peters, K. H. and Lungen, H. B., Cokemaking Ini.. 14(1), 1992. 2. Janke, W., Proc. Lnt. Conf. Alternative Routes to Iron and Steel under Indian Conditions (ARTIS), Indian Institute o f Metals, Jamshedpur (India), February 8-10, 1988.

3. Chatterjee, Amit, Singh, R., and Chakravarty, P. K., Steel Furnace Mon., 12(3), 1977. 4.

Proc. Shenyang lnt. Symp. on Smelting Reduction, Shenyang, China, September, 1986.

5. Chatterjee, Amit, Ironmaking Steelmaking, 17(3), 1990. 6. Chatterjee, Amit, Singh, R., and Pandey, B. D., Steel India, 16(2), 1983.

Chapter 2

KINETICS OF IRON OXIDE REDUCTION L REDUCTION OF IRON OXIDES IN SYSTEMS EMPLOYING SOLID AND GASEOUS REDUCTANTS It is well established that a solid-solid reaction is always slower than a reaction between a gas and a solid, for two reasons. First, the solid-solid contact area is limited, while when a gas reacts with a solid, the gas envelopes the entire surface of the solid particles; and second, solid state diffusion is much slower compared with mass transfer to/from gases. Therefore, it was recognized quite early that the overall reaction during the reduction of iron oxide takes place in two stages,* reduction of the iron oxide mFe,Oy(s) + pCO(g) = nFe,0^(s) + rC02(g)

( 2. 1)

and gasification of carbon C(s) + CO^Cg) = 2CO(g)

( 2 .2 )

The complete set of reactions involved in SR and DR processes is presented in Table 2.1, which also shows the heats of reaction (AFl°) and free energies (AF°) for each reaction at 1000 and 1400 K (about 700 and 1 100°C) because these are the critical temperatures in these processes. The chemical reactions involved are complex because three different species of iron oxides are involved, the reduction of all of which takes place through reactions of the gassolid type. At any temperature, the concentration of the reducing and the product gas(es) in the gas phase must be such that the partial pressure of the reducing gas always exceeds the equilibrium value as obtained from the calculated CO/CO 2 or H2/H 2O ratios. Solid state DR processes have to operate within a narrow temperature range, 900 to 1100°C, to prevent the formation of any semisolid phases. In large reactors which treat substantial volumes of material, it is not always possible to maintain the operating temperatures in the entire reactor within these limits. Nonetheless, it must be appreciated that these temperature limits exist in ail solid state reduction processes, and in fact, even in gas-based DR similar temperature stipulations are valid, albeit for different reasons. The total heat requirements of DR processes are made up of heat required for heating the charge to the reaction temperature and for carrying out the endothermic reduction reactions; this could amount to a major contributor, particularly when solid fuels (e.g., coal) are used and the CO required is obtained mainly through the solution loss reaction (Reaction 2.2 above). This is a basic problem in all rotary kiln DR processes, where because of the nature of the reactor, the solid and the gaseous phases are almost totally separated from each other. Heating of the charge can be carried out more efficiently in countercurrent reactors (e.g., shaft furnaces) in which the sensible heat in the upward flowing reducing gas has a chance to be transferred to the solids moving in the opposite direction. Therefore, gas-based DR processes can approach the ideal heat transfer conditions, whereas rotary kilns are the worst sufferers because of their very inefficient gas-solid heat exchange. Since DR processes, in which neither melting nor even partial fusion can be allowed, always operate at temperatures below those encountered in blast furnaces and SR processes, the rate of the reduction reaction, which is controlled essentially by the reducibility of the iron

Beyond the Blast Furnace TABLE 2J Iron Oxide Reduction Reactions Equation for reaction Heat of reaction, AH° kJ/mo!e C or H 2

SI No.

1. C + 02 = C02 2. C+C02 = 2C0 3. 4. 5. 6. 7. 8. 9.

C + 1/2 O2 = CO C + H2O = CO + 3Fc203 + CO = 2Fc304 + CO2 Fc304 + CO = 3FeO + CO2 FeO + CO = Fe + CO2 Fc203 + 3CO = 2Fe + 3CO2 FeO + C = Fe + CO

10. 3Fe203 + H 2 = 2Fe.04 + H 2 O 11. Fe3 0 4 + H 2 = 3FeO + H 2 O

12. FeO + H2 = Fe + H2O 13. Fe20. + 3H2 = Fe + 3H2O

FIGURE 2.1.

1000 K -394.63 + 170.62 -1 12.01 + 135.73 ^ 7 .0 7 + 15.69 -19.87 -14.99 + 150.75 -12.18 +50.58 + 15.02 + 19.92

1400 K -395.43 + 166.31 -114.56 + 135.59 ^ 8 .6 6 + 17.87 -1 7 .7 0 -13.22 + 148.62 -1 7 .7 4 +48.79 + 13.22 + 17.69

Free energy of reaction, AF° k j/m ole C or H 2

1000 K

1400 K

C O /CO 2 or Hj/HjO ratio

1000 K

-395.85 -5.31 -200.58 -7.99 -95.27 -1 .7 6 +3.68

___

___

-75 .0 2 -235.60 -65 .5 6 -113.72 -9.54 + 12.64

2.610

1.63 -92.59 +0.92 +6.36

-62.38 -123.18 + 19.00 +3.18

—

—

1400 K

_ —

—

—

—

—

1.05 X 10-^ 0.809 1.557

5.71 X ICr^ 0.440 2.961

—

—

—

—

1.46 X ICr^ 1.117 2.149 —

2.53 X ICr^ 0.195 1.314 —

Cross section of partially reduced iron oxide particles.

oxide particles, becomes critical. The porosity of the iron oxide particles is one of the most important factors affecting its reducibility. With hard, dense ores, the partially reduced particles show a topo-chemicaJ type of magnetite at the core^ preceded immediately by wustite and an outer shell of metallic iron (Figure 2.1 ). For very porous iron ores, a more diffused type of reduction is observed with no distinct interfaces but a gradual transformation from iron at the periphery to hematite at the core.

Kinetics of Iron Oxide Reduction

9

In the case of hard and dense iron oxides, oxygen is removed from the iron-wustite interface only according to reactions 7 to 13, given in Table 2.1. The other oxides are reduced to the next lower oxide by diffusion according to the following reactions: Fc304 + Fe^+ + 2e~ = 4FeO

( 2 .3 )

4 Fc203 + Fe^^ + 2e~ = 3 Fc304

( 2 .4 )

This mechanism requires that the reducing gas, carbon monoxide, and/or hydrogen, should readily diffuse inward through the iron layer to the iron-wustite interface, and the product gas, carbon dioxide, or water vapor should diffuse outward equally easily so that there is no accumulation of any of the species involved. A totally different mechanism is applicable in the case of porous oxides, through which the reducing gases can penetrate faster than they can react at any one interface. In this case, the reduction takes place according to reactions 5,6, and 7 of Table 2.1. Both mechanisms require that the reducing and product gases diffuse freely through the iron layer, in opposite directions. In addition, the concentration of the reducing gas must exceed that of the product gas to the extent that their ratio is greater than the equilibrium ratio shown in Table 2.1. For this to occur, the flow rate of the gases passing through a bed of solid particles, whether fixed or moving, must be large enough to ensure that a stagnant layer of gas does not build up around the particles. Fluidized bed reactors, where each individual particle of fairly small size is sup­ ported by an upward thrust of gas, approach these ideal conditions; a somewhat similar situation also exists in shaft furnaces with a larger size of oxides. Such a situation, however, cannot prevail in a rotary kiln, and thus rotary kilns are at a disadvantage in reaction kinetics. Once melting occurs, which is the case in all SR processes operating at higher temperatures, more of the heat energy is transferred automatically to the outgoing gases. Utilization of this heat energy in large volumes of the exit gas subsequently becomes the predominant factor affecting the process efficiency and economics. Even from a fundamental standpoint, there­ fore, it is obvious that proper utilization of large quantities of a rich outgoing gas is of vital importance in SR processes. To this extent, the classical blast furnace is a wonderfully conceived SR reactor, ensuring the optimal utilization of the outgoing gases within the system itself.

II. KINETICS OF REDUCTION OF IRON OXIDE Solid state reduction of iron oxide is a heterogeneous reaction involving solid and gas phases separated by an interface. The rate at which the iron oxide can be reduced to metallic iron is the primary determinant of the rate of production in any DR process. The production rate, in turn, governs the economic viability of the process and its competitiveness with other systems. It is therefore necessary to understand the mechanism of iron oxide reduction along with the kinetics, particularly since the mechanism is quite complex and involves a large number of elemental steps, the slowest of which determines the overall reaction rate (i.e., the rate-controlling step). The following are the elemental steps involved in iron oxide reduction;^ Mass transfer between the bulk gas and the solid specimen surface. Diffusion of the reducing agents from the bulk gas to the solid specimen surface or of the gaseous reduction products from the surface to the bulk gas passing through a stagnant gas film. Mass transfer between the solid oxide surface and the internal reaction interface through the layer of solid reduction products, which is influenced by interparticle diffusion of

Beyond the Blast Furnace

10

reducing agents or gaseous reduction products through a porous layer of solid reduction products, and solid state diffusion through a dense layer of lower oxides and/or through a dense layer of reduced iron. Chemical reaction involving oxygen removal at the reaction interface. Nucléation and growth of the metallic iron phase. Heat transfer to the reaction interface.

• • •

A. MASS TRANSFER THROUGH THE STAGNANT GAS FILM The supply of the reducing agent to the surface of iron oxide is controlled by diffusion through a stagnant gas film around the oxide particle, which is determined by the gas flow rate. The reaction rate increases as gas velocity increases, which decreases the thickness of the gas film. The amount of the gas transferred through the gas film is then less than that required for the reaction. There exists a critical gas velocity beyond which the reduction rate is independent of the gas flow rate. This critical velocity varies according to the overall reaction rate, since the contributions of the resistance of each elemental step to the overall reaction are inter­ related.

E. MASS TRANSFER BETWEEN THE SOLID OXIDE SURFACE AND THE REACTION INTERFACE Mass transfer in this case largely depends on the structure of the product formed on the oxide and is influenced decisively whenever the metallic layer is dense or less porous. The steps involved are: 1.

2.

3.

Solid diffusion through the layer of reduced oxides: reduction processes generally involve the transfer of oxygen and iron ions between the reaction interface and the oxide surface when the reducing agent has a chemical potential insufficient for reduction to metallic iron but sufficient for reduction to a lower oxide. In this case, oxygen is removed at the surface of the oxide, and the iron ions diffuse by a lattice vacancy mechanism which controls the overall reduction rate. Solid diffusion through a dense layer of reduced iron: in this case, the removal rate of oxygen is determined by the diffusion of oxygen through the dense metallic iron layer and the growth rate of the iron layer. Intraparticle diffusion process: when a porous spongy structure is formed on the surface of the oxide (which is why DRJ is also referred to as sponge iron), the reducing agent required for dissociation of oxygen or the removal of gaseous products from the oxide surface is provided by intraparticle diffusion through the pores formed in the layer of sponge iron. Moreover, in the reduction of porous ores, the diffusion process of carbon dioxide or water, which is formed, must be considered since the porous ore has a large reaction interface. In such cases, intraparticle diffusion within the gas phase could be a rate-determining step. It is, therefore, obvious that not only the porosity but also some structural factors, such as the extent of open pores present, influence the reduction rate. Moreover, pore distribution plays a role in the diffusion mechanism of the gas phase as well as the pressure dependence of the reduction reaction.

C. CHEMICAL REACTION OF OXYGEN REMOVAL AT THE INTERFACE In the reduction of any dense iron oxide, the volume of the oxide/iron interface is quite small compared with the total volume of porous oxide. As a result, the ratio of the rate of interface reaction to that of pore diffusion of the reducing gas or the gaseous product through the metal layer which is formed, is small. If the difference in the reducing gas concentration suddenly changes at the interface, the dissociation reaction of oxygen at the interface becomes

Kinetics of iron Oxide Reduction

11

the rate-determining step. The quantity of oxygen removed at the interface is directly propor­ tional to the reaction surface and the partial pressure of reducing gas, i.e., dN dl

-

(2.5)

= k -A , P

D« NUCLEATION AND GROWTH OF IRON In the reduction of FeO to Fe, supersaturation of iron occurs at the beginning of reduction, and the iron nucleates on FeO. This can be associated with a slow reaction rate, often observed at low temperatures, and with a rather rapid reduction rate after the nucléation of iron. The period at the beginning of reduction is called the induction period, and during the reduction of Fc203 to Fe with hydrogen at 350 to 600°C, the induction period decreases with increasing temperature. This is particularly so with decreasing reduction potentials. An induction period has also been observed during reduction with CO-CO 2 at 800°C to 1050°C, and the reduction rate increases following the nucléation of the metallic phase. In the reduction of Fc203 to Fc304 and further to Fe with solid carbon, a minimum reaction rate has been observed before the nucléation of iron. The time required for the nucléation of metallic iron from FeO is largely dependent on the initial oxygen present in the oxide. The quantity of oxygen removed in the nucléation period increases with increasing initial oxygen content. Iron nucleated on FeO grows Linearly with time for reduction with both hydrogen and carbon monoxide. However, the growth rate of iron on FeO using hydrogen is approximately 40 times higher than that at the same partial pressure of carbon monoxide at 800°C. This is because iron forms a thin and dense layer during hydrogen reduction, while in carbon monoxide reduction, the layer is thick and porous. The retarding effect on nucléation occurs only in the FeO-Fe step and not in the formation stage of Fe304 and FeO from Fe203. This is because Fe304 and FeO both have a cubic structure and are crystallographically similar, so that the transformation in this stage occurs without an induction period. But in the reduction of Fc203 to Fc304, the lattice transforms from rhombohedral to cubic, so that the reduction behavior between Fe203 and Fe304 is different. The nucléation and growth processes of iron are very important because they are concerned with the structure of the reduced phase, which in turn affects the subsequent reduction rate.

III. MECHANISM OF IRON OXIDE REDUCTION IN COAL^BASED PROCESSES In the case of iron oxide reduction with coal, the reactions between carbon and iron oxide particles begin only at the points of contact but are disrupted once metallic iron is formed in the intermediate phase. Thereafter, the reduction can proceed only as a result of the diffusion of carbon atoms through the metallic iron layer to the residual oxide. Thus, in DR processes based on either solid or liquid reductants, it is essential to convert the reductant to a reducing gas. In any rotary kiln process based on coal, the reduction of iron oxide is essentially achieved through the following reactions: 3 FC203 + CO = 2FC304 -}- CO 2

AH° = -52.87 kJ/mol

(2.6)

Fc304 -r CO = 3 FeO -r CO 2

AH° = 36.25 kJ/mol

(2.7)

FeO + CO = Fe -r CO 2

A W =17.31 kJ/mol

(2.8)

Beyond the Blast Furnace

12

The carbon monoxide required for the above reactions is generated by the carbon in the coal feed reacting with the carbon dioxide produced by the reduction reactions. This reaction, commonly known as the Boudouard reaction, may be represented as CO, + C = 2CO

AH° = 172.46 kJ/mol

(2.9)

This reaction is highly endothermic and therefore is favored only at high temperatures. Thermodynamically, carbon dioxide cannot exist at temperatures above 1000°C; however, reduction takes place owing to the two reactions, 2.8 and 2.9, taking place separately. The overall reaction may be represented as: FeO + C = CO + Fe

AH° =155.16 kJ/mol

(2. 10)

The reduction of Fc304 or FeO to iron at temperatures below 600°C is thermodynamically impossible according to the Boudouard equilibrium curve. Nonetheless, reduction does take place at these low temperatures because the rate of reaction of the gases with carbon is relatively slow compared with the rate of reaction of the gases with the iron oxides. Therefore, in coal-based DR processes, the gaseous reduction of iron oxide to metallic iron, as well as the Boudouard gasification reaction, take place simultaneously and interdependently. A characteristic feature of this type of DR is that the reducing gases react with the oxide from all sides of the bed of materials, but the flow velocity of the reducing gases is comparatively lower than that encountered in blast furnaces. The required separation of the gas phases, outside and inside the charge, is ensured by their partial pressures, thus preventing the oxidizing flue gases above the charge from reacting with the reduced oxides. This condition prevails only when the oxygen removal rate from the oxide is quite high. The major portion of the gas mixture in contact with the charge is composed of carbon monoxide and carbon dioxide. The iron oxide reduction rate becomes much slower beyond an 80% degree of metalization (this is the ratio of metallic iron to the total iron present), since the removal of the last traces of oxygen is a time consuming process. This results in an increase in the nitrogen content of the gas mixture in a rotary kiln toward the discharge end. In a given kiln cross section, it is, therefore, essential that the two gas phases are kept separate, which is achieved in practice by increasing the filling degree of the charge in a rotary kiln. When the bed depth is increased, the ratio of the exposed charge surface to the total charge volume decreases, thereby restricting the interaction between the two gas phases. In this connection, the design of the kiln outlet end is of special significance. The discharge end of any kiln is, therefore, narrowed, or a dam is provided (in the form of a refractory ring) in order to increase the filling degree. In addition to the improvement in the reducing conditions, this measure provides a longer retention time in the high temperature zone, leading to carburization of the reduced iron. However, the extent of increase in filling degree is restricted, since the heat transfer between the gases and the charge deteriorates with an increase in the degree of filling. It is, therefore, necessary to operate any rotary kiln with an optimum amount of charge (i.e., filling degree), particularly at the discharge end. Both the gas reduction reaction and the coal gasification reaction which take place simultaneously can be considered to be first order reactions. The rate of the gas reduction is proportional to the oxygen concentration in the oxide and the extent of deviation of the existing partial pressure of carbon monoxide from the equilibrium value. Similarly, the rate of the gasification reaction is proportional to the carbon concentration in the charge and the deviation of the carbon monoxide partial pressure from the Boudouard equilibrium value.

Kinetics of Iron Oxide Reduction

FIGURE 2.2.

13

Reduction curves of dense briquettes reduced with carbon monoxide.

lY. MECHANISM OF IRON OXIDE REDUCTION IN GAS^BASED PROCESSES The reason why gas-based DR processes have become so well accepted is due as much to the abundant availability of natural gas in many parts of the world as to these processes’ economic and technical advantages in reducing iron oxides. It is, therefore, essential to understand the mechanism of hematite reduction with hydrogen and carbon monoxide mix­ tures, which are the reducing species present in reformed natural gas. A. REDUCTION OF DENSE IRON OXIDE When pure carbon monoxide is used to reduce dense Fe203, the reduction stops at around 40% at 700°C and at around 85% at 800°C but is completed at higher temperatures. Figure 2.2 represents the reduction curve of dense Fe203 briquettes with carbon monoxide."^ Incom­ plete reduction at low temperature is a result of the high amount of carbon deposition on the surface of the oxide, which seals up the pores and prevents further diffusion of gases. The addition of hydrogen to carbon monoxide, however, improves the rate of reduction, as is indicated in Figure 2.3. Here again, the reduction is incomplete at 700°C and 800°C with a 5 OCO-5 OH2 mixture, but as is shown in Figure 2.4, the degree of reduction is higher compared to that with pure carbon monoxide. The reduction rate is faster with hydrogen than with carbon monoxide, and the rate is intermediate with a mixture of hydrogen and carbon monoxide. Thermodynamically, hydrogen reduces iron oxide more easily than carbon monoxide, as is revealed by the free energy changes which occur at 1000°C for both cases: F e 203 -b 3 H . = 2 F e -b

3 H 2O

Fe^O , -b 3 C O = 2 F e -b 3 C O 2

AF° = -150.62 y/m ol

( 2 . 11)

AF° = -129.70 kJ/mol

( 2 . 12)

For reduction with a gas mixture of hydrogen and carbon monoxide, the relative contribu­ tion of each gas to the overall reaction rate depends on the reactivity of the molecules and their selective absorption on the iron oxide surface. The overall reaction rate is not only equivalent to the specific rate of chemical reaction but also depends on the active surface area of the oxide and the diffusion coefficient of the reducing gas. The Arrhenius plots for the reduction of

Beyond the Blast Furnace

14

3

C

6

X o E

m

25

50

75

100 % CO

Reducing gas composition FIGURE 2.3.

Effect of gas composition on the rate of reduction of dense briquettes at 1000°C.

FIGURE 2.4.

Reduction curves of dense briquettes reduced with a 50CO-50H2 mixture.

Kinetics of Iron Oxide Reduction

FIGURE 2.5.

15

Arrhenius plots for the reduction o f dense briquettes reduced with CO/H 2 mixtures.

dense iron oxides as a function of different CO/H 2 ratios is given in Figure 2.5, and the calculated activation energies are as follows: Activation energy Gas composition

(kj/mol)

100% Carbon monoxide

31.59 37.40 40.12 45.31 53.56

75% CO + 25% Hp 50% CO + 50% H2

25% CO + 75% H 2 100% Hydrogen

In the reduction of iron oxide with carbon monoxide or CO-H 2 mixtures, the iron phase always consists of two zones: the outer is more porous and is mainly iron carbide, while the inner is iron. In the case of reduction with hydrogen, the iron carbide layer is totally absent, and only the iron zone exists. During reduction with hydrogen, the initial portion of the reduction curves is almost linear, indicating a constant rate of reduction, and the activation energy value for this case indicates a chemically controlled process. On the other hand, reduction with carbon monoxide (Figure 2.2) shows a gradual decrease in the reaction rate with the progress of reduction. This is a result of the diffusion of gases through the increasing thickness of the porous iron layer. The activation energy value for carbon monoxide reduction signifies that the process is controlled both by chemical reaction and by gaseous diffusion. In reduction with carbon monoxide, the reduced iron reacts with carbon monoxide or the deposited active carbon to form iron carbide, the extent of which can be gauged from the carbon content of sponge iron. Since during reduction with a gas mixture of carbon monoxide and hydrogen the rates are intermediate between those for reduction with pure carbon monoxide and with pure hydrogen, the rate-controlling process is a combination of those prevailing in the reduction with the pure gases. With increasing CO/H 2 ratios, the amount of iron carbide formed would increase, resulting in increased resistance to diffusion of gases. B, REDUCTION OF POROUS IRON OXIDE It is to be understood that the reduction rates of porous iron oxides are faster owing to their larger surface area, as shown in Figures 2.6 and 2.7. Though both these figures show that some

Beyond the Blast Furnace

16

FIGURE 2.6.

Reduction curves of porous briquettes reduced with carbon monoxide.

FIGURE 2.7.

Reduction curves of porous briquettes with a 50CO-50H2 mixture.

reduction curves overlap at certain temperatures, Figures 2.8 and 2.9 clearly indicate a minimum rate at 975°C for carbon monoxide reduction and another minimum at 1025°C for reduction with a 50CO-50H2 mixture. The activation energy values are much lower than those obtained for dense oxides, but again vary with gas composition as is shown below: Reducing gas

Activation energy (kj/mol)

100% Carbon monoxide 75% CO + 25% H2 50% CO + 50% H2 25% CO + 75% H2 100% Hydrogen

12.05 12.55 15.31 17.20 21.51

From these activation energy values, it is evident that the rate of reduction is controlled by the gaseous diffusion for all the gas mixtures. Arising out of the fast chemical reaction rates in the case of porous oxide material, the rate of gaseous diffusion through the material becomes the slowest step. At the rate minima temperatures, which are rather high, sintering of iron (or

Kinetics of Iron Oxide Reduction

17

FIGURE 2.8.

Effect of temperature on the rate of reduction of porous briquettes with carbon monoxide.

FIGURE 2.9.

Effect of temperature on the rate of reduction of porous briquettes with a 50CO-50H2 mixture.

iron carbide) around the wustite grains isolates the grains from direct contact with the reducing gas and, hence, hinders reduction. The characteristics of reduction with CO-H 2 mixtures does not vary significantly from carbon monoxide reduction except for a 50°C shift in the rate minimum temperature. This probably occurs because the amount of iron carbide formed during reduction by carbon monoxide is greater than that formed in the case of CO-H 2 mixtures, which allows the same degree of sintering to occur at a lower temperature in the former case.

V. CARBON DEPOSITION AND ITS BENEFITS Carbon deposition on reduced iron takes place for the following reasons: 1.

At temperatures below 900°C, the fine carbon formed by the reaction 2C 0 = C02 + C reacts with freshly formed iron to give iron carbide.

(2.13)

Beyond the Blast Furnace

18

2.

At temperatures above 900°C, carburization of reduced iron occurs by the reaction 3Fe + 2CO = FeX + CO,

(2.14)

In DR processes where the degree of metalization does not usually exceed a maximum of 95 to 96% and is often restricted to around 90 to 92%, the carbon deposited on the iron oxide acts as a reducing agent for the remaining oxides at the high temperatures encountered during steelmaking and also lowers the melting point of DRl. In steelmaking, it is also advantageous if the materials charged already contain sufficient carbon (as iron carbide) to produce specific steel grades. Carburization of DRl is thus beneficial and can be controlled by adjusting both the reduction temperature and the gas composition, in order to control the carbon content of the metalized material within certain limits. Thus, better carbon control and higher carbon levels are more easily accomplished in gas-based DR processes rather than in rotary kilns where the temperature and composition of the gas generated from coal cannot be varied appreciably.

REFERENCES 1. Chatterjee, Amit, Proc. XV Symp. of SIDOR, Puerto Ordaz, Venezuela, November 1988. 2. Ross, H. U., McAdams, D., and Marshall, T., Direct Reduced Iron — Technology and Economics of Production and Use. Iron and Steel Society of AIME, USA, 1980. 3. Tokuda, M., Yoshikoshi, H., and Ohtani, M., Trans. Iron Steel Inst. Jpn.. 13(5), 1973. 4. El-Geassy, A. A., Shehata, K. A., and Ezz, S. Y., Trans. Iron Steel Inst. Jpn.. 17(11), 1977. 5. Von Bogdandy, L. and Engell, H. J., The Reduction of Iron Ores, Springer-Verlag, Berlin, 1971.

Chapter 3

COAL^BASED DR PROCESSES USING ROTARY KILNS L PROCESS TECHNOLOGY IN GENERAL In all coal-based DR processes, sized iron ore (pellets can of course be used if desired) and a coarse fraction of noncoking coal are fed into the rotary kiln from the inlet end in the required proportions. The coal not only acts as a reducing agent, but it also supplies the heat required for maintaining the temperature profile of the charge within the kiln. A finer fraction of coal is also often introduced from the discharge end of the kiln to help complete the reduction process. The temperature of the charge bed inside the kiln is confined to a maximum of around 950 to 1050°C so that the entire reduction occurs in the solid state. Noncoking coal contains sulfur, which can be partially absorbed in sponge iron during the course of reduction — to control this, a flux such as limestone or dolomite must be added to the coal charged into the kiln. The product discharged from the kiln is indirectly cooled in a rotary cooler to room temperature, so that no water normally comes in direct contact with the reduced product, which, being a nonequilibrium material, is prone to reoxidation. Figure 3.1 shows a simplified flow sheet of direct reduction in rotary kilns. The final product with a high degree of metalization exhibits a "honeycomb” structure under a microscope; it is therefore called "sponge” iron. Since sponge iron is magnetic in nature, this feature is taken advantage of in separating the sponge iron from the non magnetic portion consisting mainly of coal ash or char. The most critical factor in the reduction process is the controlled combustion of coal and its conversion to carbon monoxide. This is achieved by controlled introduction of air through • •

•

The discharge end of the kiln, Kiln shell mounted blowers, which supply air to secondary air pipes so that air enters axially along the kiln length in the space above the charge bed. These secondary air pipes are sometimes equipped with swirlers at the tip to aid combustion and face either toward the inlet or the discharge end of the kiln, Radial/submerged injection into the charge bed through nozzles in the preheating zone, covering about 30% of the kiln length from the charge end.

Figure 3.2 shows a schematic view of the cross section of a rotary kiln and the process of reduction.' The coal bums in the presence of insufficient air to carbon monoxide: C -r 1/2 O, = CO

(3.1)

The carbon monoxide reduces the iron oxide gradually to metallic iron as the charge travels down the length of the kiln. To this extent, the "direct” part of DR is not reduction with carbon, as in the case in blast furnace ironmaking, but DR refers to the direct conversion of iron oxide to metallic iron in the solid state. Carbon dioxide is obtained as the gaseous product of reduction reaction, and it again reacts with the carbon present in the coal to produce carbon monoxide: CO -r C 2 = 2CO This cycle continues, maintaining the reducing conditions prevailing in the kiln.

19

(3.2)

Beyond the Blast Furnace

20

COARSE

SPCmOE ^ON

FIGURE 3.1.

FINE SPONGE IRON

CHAR

Basic principles of sponge iron making in rotary kilns.

SECONDARY AIR

TEMPERATURE

950“-1050®C f 4 CO, ^ 2C0 F6|0 j ♦ C 0 bs-F6|0^ + C0|

FejO^ 4 CO _ ^ F e O 4 COj FeO 4 C O _ ^ F e O 4 COj

SUBMERGED AIR INJECTION

FIGURE 3.2.

Schematic diagram showing the cross-sectional view o f a rotary kiln and the process o f reduction.

As shown in Figure 3.3, all rotary kiln-based DR processes operate on the countercurrent principle, in which gases move in a direction opposite that of the flow of solids.^ Various unit operations occur both in parallel and in series, e.g., transport, mixing, grain separation, heating, gas generation, and reduction. Segregation of charge materials (ore, coal, flux) because of size and density differences, as well as because of the slope and rotation of the kiln, must be prevented, by adopting an appropriate design and practicing correct operating mea­ sures.^ Another area of critical importance is the prevention of localized areas of high temperature, which can result in the formation of accretions on the kiln lining and ball-shaped clusters within the bed.'^ The difference between the gas phase and the charge bed temperature (normally about 100 to 150°C) plays an important role in this context. Most of the established rotary kiln-based DR processes operate on the principle that the reduction is carried out exclusively by carbon monoxide obtained by the gasification of the fixed carbon content in coal. In case the carbon monoxide generation falls short, fine coal with a grain size of -3 or -6 mm can be combusted to make up for the lack of carbon monoxide.

Coal-Based DR Processes Using Rotary Kilns

FIGURE 3.3.

21

Temperature and degree of metalization at various locations in the kiln.

Volatiles in coal ( water & ash f r e e ) , wt %

FIGURE 3.4.

Proportion o f heat content o f volatiles as part of totaJ heating value for different Indian coals.

In this situation, about 25 to 50% of the total coal requirement is injected into a zone where the fresh coal is immediately subjected to temperatures above 800°C, at which reduction reactions take place. The balance coal (-15 and +6 mm or -15 and +3 mm) is introduced at the charging end of the kiln together with the ore, desulfurizing agent, and sometimes an amount of recycled char. The volatile content of this cocurrent coal is released at temperatures up to approximately 600°C, i.e., before the reduction reactions start. Some of these volatiles are utilized to preheat the kiln charge; however, the majority leave the kiln with the waste gases without contributing to the actual process of reduction. In order to lower the overall energy consumption, it is therefore essential to optimize the utilization of the hydrocarbons in the volatiles. The energy content of coals is determined by their fixed carbon contents and volatiles,^ the latter mainly consisting of hydrocarbons. Figure 3.4 indicates the extent to which the hydrocar­ bons contribute to the overall heating value of coals. It can be seen that this contribution varies between 28 and 42%, depending on the type of coal. In general, this figure increases with increasing volatiles content, except for some coals with a relatively high oxygen concentration

22

Beyond the Blast Furnace

Volat-iUs in coal (wat«r & ash fre e ), wt V*

F IG U R E 3.5.

Waste heat recovery for electric power generation — dependence of power generated on coal quality.

in the volatiles. Unless the volatiles present in the coal are effectively utilized, the specific coal consumption in relation to energy consumption, particularly the heat content of the kiln off gas, will be high. Therefore, some modem rotary kiln plants are equipped with waste heat recovery systems which make use of this otherwise wasted energy. Figure 3.5 illustrates the amount of electric power which can be generated per ton of DRl, depending on the coal quality as well as the amount of fixed carbon consumption per ton of DRl. However, such a waste heal recovery system and generation of electric power may not always be economically viable except for plants having a production capacity of more than 200,000 tpa of DRJ using at least two kilns.

II. SPECIFIC ROTARY KILN DR PROCESSES IN COMMERCIAL USAGE The characteristic features of the important coal-based DR processes in commercial usage at present are presented in Table 3.1.^ The important processes in this group include SL/RN, CODIR, ACCAR, DRC, and TDR; there are many common features in all these systems, and only the unique features of each are highlighted below.

A. SL/RN This process, based on the direct use of noncoking coal, is the most widely applied DR production route.^ it was developed jointly by the Steel Company of Canada (Stelco), Lurgi Chemie, Republic Steel Company and National Lead Corporation in 1964. Figure 3.6 shows the basic flowsheet of the process. The reactor is a refractory lined cylinder rotating on an inclined axis, which is continuously charged at a controlled rate with appropriate amounts of size-graded iron oxide (pellets, lump ores, or ore fines), limestone, coal and return char. The materials charged into the kiln gravitate toward the discharge end and are thereby progres­ sively heated to the reduction temperature of about 1000 to 1100°C as the materials roll down the gradient. The product discharged from the kiln is cooled in an externally cooled rotary cooler to around 100°C and passed through magnetic separators and screens to separate

Coal-Based DR Processes Using Rotary Kilns

23

sponge iron, coal ash, and char. Waste gases leaving the kiln at the inlet end pass through a dust chamber and postcombustion chamber and are cooled and cleaned in electrostatic precipitators, scrubbers, or bag filters before being released to the atmosphere. Alternatively, clean kiln gases can be used in waste heat boilers to recover the sensible heat, and the steam generated can be utilized for heating purposes or conversion to electric power. The rotary kiln plants based on SL/RN technology which are presently in operation are shown in Table 3.2. In the year 1991, SL/RN plants alone produced 0.83 million tons of DRI, accounting for about 56% of the world’s total production of coal-based DR processes, and the dominance of this process continues. During the past 20 years, SL/RN plants have demon­ strated the flexibility of the process with regard to the use of a large number of various ironbearing materials and coals, as is summarized in Tables 3.3 and 3.4. The salient features of the technology include: • • • • • • •

Flexibility with regard to the type of iron-bearing materials which can be used, such as lump ore, pellets, ilmenite, iron sands, and steel plant wastes, Use of a wide variety of solid fuels ranging from anthracite to lignite, including charcoal, Improved heating system for the charge by submerged air injection in the preheating zone of the kiln in plants installed recently, Optimized coal injection facilities, Waste gas conditioning by controlled post combustion, Cold briquetting of DRJ fines whenever required, Waste heat recovery through steam generation, in some cases.

As summarized in Table 3.5, regular operation of some of the new plants has shown high performance levels with respect to product quality, plant availability, and energy consump­ tion. It may be observed that the SL/RN process has satisfactorily treated raw materials of diverse characteristics, and the process is well established.

B. CODIR The C O D IR (Coal-Ore-Dlrect Reduction) process was developed through experience gained from other reduction processes, namely the K R U P P -R E N N process and the WELZ process for reducing zinciferrous raw m a t e r i a l s . I t is macroscopically very similar to the SL/RN process and is capable of reducing high grade lump ores as well as fine ore concen­ trates into metalized products. Figure 3.7 shows the flow sheet of the process. Once again, the iron oxide is fed into an inclined rotary kiln, together with the reductant and the desulfurizing agent; it is heated and then reduced to metalized sponge iron by the hot kiln reducing gases traveling in the countercurrent direction. The material leaving the kiln is a mixture of DRI, surplus fuel, and ash, all subjected to cooling in a rotary cooler and subsequently to screening, magnetic separation, and gravity separation. Some CODIR coolers use a mist of water injected directly into the cooler rather than indirect water sprays on the cooler shell, but the water is used in controlled amounts to prevent oxidation of hot sponge iron. Commercial operation of the first Krupp (now CODIR) plant was started in 1974. This plant, with an annual capacity of 120,000 to 150,000 t (depending upon the reducibility of iron ore charged), was constructed at Dunswart Iron and Steel Works in South Africa with a rotary kiln of 75 m length and an internal diameter of 4.6 m. Since its start-up in 1974 and until September 1990, the plant produced more than 1.5 million tons of DRJ with a metalization averaging 92%. It is claimed that by enrichment of combustion air with oxygen (23% O 2), the energy consumption in the process has been reduced from 20.0 GJ/t to 14.7 GJ/t DRI. Figure 3.8 shows typical figures of thermal energy consumption at Dunswart Steel’s CODIR plant.

64 min.

4 max. preferred

—

—

As low as possible

Wide size range acceptable

Noncoking coals and fossil fuels (solid, - 1 0 mm)

4 max. preferred

—

—

As low as possible

Wide size range acceptable

Noncoking coals and fossil fuels (solid, - 1 0 mm)

Total gangue. (Si 0 2 + AI2 O3 ) Si 0 2 A1203 S P

Size range of iron bearing materials, mm

Reductant used

Ore fines, lumps, and pellets

64 min.

Ore fines, lumps, and pellets

KRUPP

Fe total

Desired analysis of oxide feed, %

Raw material

SL/RN

Noncoking coals (solid, -6 /5 0 mm). oil and gas

10-15 (lump ore). 10-16 (pellets)

—

5-25

As low as possible

Noncoking coal/ Noncoking coal. noncoking coal anthracite, etc. and oil/noncoking cocurrent coal coal and coal-oil (-3 2 mm), blown coal ( - 1 0 mm) slurry

5 -2 0 (lump ore). 6 - 2 2 (pellets)

As low as possible but 0.08% max.

—

—

—

.5-3.0 . - 2 .0 0.05 1 0

1

4 max. preferred

4 max. preferred

4.5 max. preferred

Lump ore and pellets

DRC

64 min.

Lump ore and pellets

TDR

64 min.

65 min.

Lump ore and pellets

Allis Chalmers

2.89 17.39 47.21 4.28 5.69 0.348 0.91 10.75

9.0 1.0

50.0

Mostly below 325 mesh which is pelletized

C Zn

Fe (total)

Anthracite (ash 9.8%, Coke breeze (6 -2 0 mm) SO.77%, lignite, VM 11.6% and oil and C (fix) 78.6%) mostly below 1 0 mm

Mostly below 325 mesh which is pelletized

Fe (Met) FeO Fe2 0 3 Si 0 2 CaO S Zn C

49.43

Typical raw materials composition, %

Typical raw materials composition, %

Fe (total)

In plant dust in the form of filter cake and dry dust caught in the dust collectors

Kawasaki

BF sludge, BOF sludge, mill sludge, and yard sludge

Sumitomo

TABLE 3A Features of Some Direct Reduction Processes Using Noncoking Coal

s 3

Cö

s Si-

Cö

M

Rotary kiln

Type of reduction reactor Rotary kiln

Water (make-up). m3

Limestone. kg Electric power. k w h

Natural gas/fuel oil. coal. GJ

Approximate specific consumption per ton of sponge iron

Low sulfur

Low sulfur

Desirable properties of the reductant

Ported rotary kiln

Low sulfur in oillgas (0.75%). low ash coal

60 (dolomite) 1 20

Rotary kiln

Low sulfur (preferably below I % )

40 65-1 10

Rotary kiln

Low sulfur

Rotary kiln

Coke consumption = 220 kg. Oil = 70 1

Rotary kiln

Beyond the Blast Furnace

26

FIGURE 3.6.

SL/RN process of direct reduction.

TABLE 3 2 SL/RN Plants in Operation

Company Western Titanium'* Highveld (prereduction)^ New Zealand Steel Acos Finos Piratini Siderperu S IIL ISCO R New Zealand Steel Western Titanium Sands^ Bihar Sponge Iron

Country

Start-up date

K iln

units

K ile

size,

m

Australia

1969

1

2.4

X

30

South Africa

1968/84

11

4.0

X

60

New Zealand Brazil

1969

1

4.0

X

75

1973

1

3.6

X

50

Peru

1980

3

2.9

X

India South Africa New Zealand Australia

1980 1984 1985

2

3.0 4.6 4.6

X

4 4

1986

1

India

1988

1

Iron ore

Coal

Capacity, t

emende cone. Lump ore

Subbituminous Bituminous

150,000

Lignite

175,000

62

Beach sand cone. Pellets/ lump ore Pellets

X

40 80 65

Lump ore Lump ore Beach sand

4.6

X

65

4.6

X

80

Ilmenite cone. Lump ore

X

11

Bituminous

15,000

60,000

Coke/ Anthracite Bituminous Bituminous Lignite cone.

60,000 720,000 900,000

Subbituminous

150,000

Subbituminous

150,000

1 2 0 ,0 0 0

Pre-reduction of ilmenite

In the middle of 1989, a CODER plant with a design capacity of 150,000 tpa was commis­ sioned for Sunflag Iron and Steel Company Ltd. at Bhandara, in western India, as a part of an integrated mini mill. The rotary kiln of Sunflag has a length of 80 m with an internal diameter of 5 m. Between August and December 1989, this plant completed a continuous production campaign of 138 d at a production rate of approximately 70% of the rated capacity. The

Coal-Based DR Processes Using Rotary Kilns

27

TABLE 3 J Typical Composition of Coals Used in SL/RN Plants AFP

NZS

SIDERPERU

iSCOR

BSIL

Plant Location Coal type

Brazil Bituminous Lignite

New Zealand Lignite

Peru Coke breeze

South Africa Subbituminous

India Subbituminous

49 42 9 0.3 18

81 3 18 0.7

58 28 14 0.7

50 25 25

Proximate Analysis, % by wt Fixed carbon Volatile matter Ash Sulfur Moisture

45 51 4 0.3 17

40 25 35 0.4 9

1

0 .6

6

10

TABLE 3A Typical Chemical Composition and Size of Iron Bearing Feed in SL/RN Plants Plant Location Ore type

Lump

AFP

NZS

WSL

ISCOR

BSIL

Brazil

New Zealand

Australia

South Africa

India

Beach sand

Ilmenite

Lump

Lump

57.8 3.7 4.2

27.0

Pellet

Chemical composition, % by wt. Fe total SiO, A 1 ,0 , CaO M gO TiO, S P Grain size, mm

69.4 0.58

6 8 .2 1.2

0.5 0.03

0 .8

1.1

0.3

3.1 7.8 0.04 0.06 -0.5

0.005 0 .0 2

5-15

5-20

1.0

0.9

6 6 .0

2.5 1.3 0 .2

0.15 58.5

-0.5

0.009 0.04 5-15

67.4 1.5 1.0

0.05 0.05 0.009 0.03 5-15

composition of coals used in the in d u strial CODIR k iln s mentioned above is shown in Table 3.6. One of the unique features of CODIR technology is the countercurrent injection of coarse coal with sizes of 5 to 25 mm, going up to 35 mm, for highly reactive coals. In this system developed in 1981, the majority of coal is injected from the discharge end of the kiln and distributed in such a manner that the bed temperature does not fall below 950 to 1000°C owing to localized coal concentration. The coarse coal particles that settle in the high temperature zone of the kiln are mixed with the burden while releasing the volatiles so that these can also participate in the reduction of iron oxide, thus substituting fixed carbon. If the coal quality allows effective char recovery, even the total amount of fresh coal can be injected and only the recycled char added together with the feed. This coal injection system, depicted in Figure 3.9, makes use of the coal volatiles for reduction, thereby reducing the requirement of fresh coal to a minimum. After the incorporation of this coal injection technique, it has been possible to reduce the energy consumption in the CODIR process to 14.65 GJ/t of DRI. Figure 3.10. shows a comparison of the process heat balance with and without coarse coal injection. It is clear that the energy consumption decreased from 20 GJ/t DRI (with finer coal injection) to 15 GJ/t (in the case of coarse coal injection). With the use of 80% bituminous coal and 20% semianthracite, the specific coal consumption decreased from 690 to 530 kg/t DRI.

Beyond the Blast Furnace

28

TABLE 3.5 Performance of Some New SL/RN Plants Typical chemical analysis. % by wt. Fe total Fe metallic FeO

SiOj

A IP 3 CaO M gO S P Metalization

FIG URE 3.7.

Product quality IS C O R

B S IL

(South Africa) (India) 90.4 84.0 8.1

5.8

1.8

0.3 —

93.7 8 6 .2

9.64

Specific consumption (per ton of D R I) PelletsAump ore Coal Electrical energy Water

1.43-1.45 i 380-400 kg fixed carbon 70-80 kWh 2.5 m

2.4 Manpower

0.35-0.60 man-hours

Plant availability

90-92%

1.8

0.05 0.05

0 .0 1

0 .0 2

0.05 93.0

0.04 92.0

General flow sheer o f the CODIR process.

C. ACCAR The ACCAR (Allis-Chalmers Controlled Atmospheric Reduction) process was developed in 1972 at Niagara Metals ( C a n a d a ) . T h e heart of the process is a ported rotary kiln, as is shown in Figure 3.11, in which natural gas/fuel oil can be used either singly or together, without any external reforming to supplement coal which is used as the primary reductant in the process. Except for this fundamental difference (i.e., the use of gas + coal/oil or coal/oil or gas + coal), keeping the coal percentage at a maximum of 80% in all cases, the process is similar to the SL/RN and CODIR processes. In the ACCAR system, coal is charged along with the oxide feed, whereas oil/gas is injected directly into the ore bed, along the length of the reactor through the radial ports.

Coal-Based DR Processes Using Rotary Kilns

FIGURE 3.8.

29

Thermal energy consumption per ton of sponge iron in Dunswart Steel’s C O D IR plant.

TABLE 3,6 Analysis of Coals Used In Industrial CODIR Kilns Coal type

Cfi., %

Volatiles, %

S, %

Ash, %

Caloric value, GJ/t

Bituminous coal' Bituminous coaP Semi-anthracite Anthracite Coke breeze

41.3 58.0 68.0 78.0 84.5

28.2 28.0 13.0 11.0 4.0

0.5 1.0 1.3 0.8 0.5

30.0 13.0 17.7 10.2 11.0

21.7 28.8 27.7 30.6 29.5

‘ India. South Africa.

Gas f e m p t r ^ of of food coal

femprature .

Area of distribution of injected coarse ceat (60- 80% of, total cpaO

70 75

•-Preheating zone

FIGURE 3.9.

Kiln temperamre profiles and injected coal distribution area.

Beyond the Blast Furnace

30 WITHOUT COARSE COAL INJECTION 20 GJ/T DRI RESP. 0.4éT C-FIX/T Fe

LOSSES

WITH 70% COARSE COAL INJECTION ENERGY CONSUMPTION ^5 RESP. 0.3S T C-FIX/T Ft

-IRON ORE WITH 41% Ft -COAL WITH C-FIX 41.S % VOL. H. 27.3 % ASM 11.1 % SULPHUR 0.74% -METALLIZATION : 92%

FIGURE 3.10.

Energy balance for rotary kiln operation with and without coarse coal injection from the discharge

end.

FIGURE 3.11.

Flow sheet of the ACCAR process.

symmetrically arranged in equally spaced rows and located up to 80% of the length of the kiln from the discharge end. The arrangement of the ports and valves is such that oil/gas and air are alternately injected through the same port; when a particular port is under the charge bed, oil is injected, and when the same port approaches the “above bed” position, air is injected through it. This radial oil-with-air distribution system apparently permits closer control of the temperature profile along the entire length of the reactor. It is claimed that the ACCAR process using coal + oil has certain distinct advantages, such as higher carbon content in the product, high degree of metalization, lower energy consumption, lower operating temperatures, etc., over other rotary kiln DR processes using coal exclusively as the reductant. The lower kiln operating temperature is indeed a significant advantage, particularly if the coal used for reduction has a low ash fusion temperature. However, all these advantages due to the usage of oil must be judged from the point of view of overall economics of the process, particularly in relation to the relative price of oil and coal.

Coal-Based DR Processes Using Rotary Kilns

FIGURE 3.12.

31

Flow sheet of the D R C process.

The process is capable of treating both lump ore and iron oxide pellets and reportedly produces a nonpyrophoric sponge with 90 to 92% metalization and 1.0 to 2.0% carbon because of the deposition of carbon in the sponge iron following the cracking of oil, especially close to the exit end of the kiln. The extra carbon in ACCAR sponge iron is advantageous for electric furnace steelmaking. The energy consumption in the ACCAR process ranges between 13.4 and 14.6 GJ/t of sponge iron. The ACCAR process has a rather limited track record. After the 100 tpd demonstration plant was operated in Canada for a number of years, the process was commercialized in 1983 in a 500 tpd unit in India. The unit in India has since switched to the use of 100% coal (abandoning the use of oil for economic reasons) with a 20 to 25% reduction in the kiln capacity. D. DRC The DRC (Direct Reduction Corporation) process was developed from the Hoc kins process for synthetic titanium dioxide production originally operated by Western Titanium Ltd., Australia.*^ Raw materials consisting of lump ore/pellets, coal, and limestone are continuously fed to a refractory-lined kiln rotating at less than 1 rpm, as is shown in Figure 3.12. Combustion air is supplied by shell mounted fans via internal air tubes. Coal is blown from the discharge end to supplement the heat generated by the combustion of carbon monoxide escaping from the bed. Thermocouples installed along the kiln length accurately measure temperatures and allow close control of temperature using the air tubes. The product is discharged via a sealed transfer chute again to an indirectly cooled rotary cooler and then sent for screening and magnetic separation. This yields coarse and fine DRl, nonmagnetic fine waste, and char, which is recycled in much the same way as in all other rotary kiln processes. A 50,000 tpa demonstration plant based on this process went into operation in 1978 at Rockwood, Tennessee, USA, some typical operational details from which are presented in Table 3.7. The distinguishing features of the DRC process are as follows: 1.

The direction of the outlets of the secondary air tubes in the preheating zone of the kiln is such that it ensures a cocurrent flow (with respect to the kiln gases), but in the reduction zone, the tubes are positioned to give a countercurrent flow. It is claimed that

32

Beyond the Blast Furnace TABLE 3 J

Typical Operational Results of the DRC Plant at Rockwood Characteristics of iron ores F * p''total? %

Swelling index

Gangue,

65 67

15

- 1 2

7 4

-25

6 6

—

Size, mm

Type Lime bonded indurated iron oxide pellets Acid indurated iron oxide pellets Hematite ore Pelletized mill scale Manganese lump ore

-15

- 2 0

-30

10

63 5

%

6

5 —

10

—

Characteristics of coals F.C., % Ash (wet), % V.M. (wet), % Moisture, % Ash fusion temperature (l.D.) in reducing condition, °C Free swelling index (FSI) Relative reactivity Sulfur, % Ash: Silia ratio Char strength index

38 to 84 5 to 2 0 6 to 33 2 to 25 1066 to 1510 0 to 6 40 to 500 0.4 to 2 . 0 45 to 8 8 26 to 6 8

Temperature profile inside the kiln with different ore/coal combinations, °C Sub-bituminous coal/pelletized ore Semi-anthracite coal/lump ore Bituminous coal/lump ore Bituminous coal/pellets

940

1010 1050 1060

Typical analysis of sponge iron produced, % Total iron Metallic iron Metalization Carbon Sulfur

2. 3.

89 82 92 0.05 0.008

to 94 to 91 to 97 to 0.25 to 0.05

this system of air injection inside the kiln ensures optimum heat evolution and heat transfer in various zones. In the DRC process, a part of the coal and char is injected pneumatically into the kiln from the discharge end. The trajectory of the countercurrent reductant injection can be altered by changing the position/inclination of the delivery pipe.

So far, only one 300 to 400 tpd commercial plant based on the DRC process is in operation in South Africa, and the operations have been very successful; yet this technology is not in use in other units. K TDR The TISCO Direct Reduction (TDR) process of sponge iron manufacture in rotary kilns was developed in India in a 10 to 12 tpd pilot p l a n t . T h e technology developed by Tala

Coal-Based DR Processes Using Rotary Kilns

FIG URE 3.13.

33

Flow sheet of rotary kiln direct reduction by the TDR process.

Steel has now been commercialized in a 300 tpd plant which has since been upgraded, with the association of Lurgi, to approximately 400 tpd.^"^ Figure 3.13 shows the simplified flow diagram of the TDR process. The salient features of the process are:^"^^^ • • • •

•

•

The process uses noncoking coal exclusively as the reductant, and fuel oil is used only for preheating the kiln. Coal is introduced in specified size ranges and proportions from both ends of the kiln along with a flux. The process prescribes the use of dolomite as the flux for scavenging sulfur, in a specified size range and proportion. The normal carbon content of the product is 0.2%. However, this can be increased to some extent by spraying a small quantity of oil or a coal-oil slurry at the discharge end of the kiln through a specially designed lance. There is a provision for radial and axial injection of air up to 30% of the kiln length from the inlet end, and in the remainder of the kiln, air is introduced axially through secondary air tubes protruding up to the center line of the kiln. The trajectory and amount of countercurrent coal injected from the kiln discharge end helps in maintaining a flat temperature profile and proper reduction conditions all along the kiln length.

The rotary kiln of the Ipitata plant (the only commercial venture using this process) has a length of 72 m, an internal diameter 3.75 m (after lining), and a lining of 225-mm-thick, highalumina, refractory bricks, castables. It is inclined at 1.432° (2.5% slope) and rotated by thyristor-controlled DC shunt motors to give any speed between 0.1 and 0.8 rpm. It has a fixed inlet hood with provision for collecting the back spilled materials and dust separately. Suitable arrangements are made for sealing the inlet end of the kiln and feed pipe through which raw materials enter into the kiln. The kiln outlet hood accommodates the countercurrent coal

Beyond the Blast Furnace

34

TABLE 3«8 Specific Consumption and Requirement of Raw Materials in the TDR Process ÌW materials Iron ore CoaJ Dolomite

Incoming size, mm

Specific consumption, kg/t

Loss as fines,

- 2 0 and +5 -1 5 0 - 4 and +1

1550-1600 750-850 70

80 50-100

kg/t

throwing pipe, a retractable oil lance, and the start-up oiJ burner used for initial heating of the kiln. The kiln has ten secondary air blowers with inlet vane controls mounted at equal intervals on the shell along its length in order to introduce controlled amounts of air into the kiln for controlling the temperature. The temperatures inside the kiln are measured by thermocouples fixed at regular intervals along the kiln length. There is a provision for taking samples of the in-process raw materials at different locations along the length of the kiln. Waste gases from the kiln enter into the postcombustion chamber to complete the combus­ tion of unbumed carbon monoxide. The gas is then cooled and cleaned in a venturi scrubber, and finally, a waste gas fan exhausts the cleaned gas through the chimney. The rotary cooler is 36.5 m long, having a shell diameter of 3.5 m. It is inclined at 1.432° and can be rotated at an adjustable speed of 0.14 to 1.4 rpm through a thyristor controlled DC drive. The discharge end of the cooler is sealed to prevent any ingress of air. Hot sponge at about 950 to 1000°C discharged from the kiln is cooled to below 100°C as it passes through the cooler, by indirect cooling using water sprays located along the outer surface of the inclined cooler. Materials discharged from the cooler consist of sponge iron and char, which are subsequently separated by screening and magnetic separation. The specific consumption of raw materials and their annual requirement for the 300 tpd Ipitata plant are given in Table 3.8. Table 3.9 shows the physicochemical characteristics of the ores used at Ipitata, while the characteristics of noncoking coals used are given in Table 3.10. Out of the total sponge produced, the proportion of -1-5 mm, -5 and +2 mm, and -2 mm fractions generally ranges between 70 to 80%, 15 to 20%, and 5 to 10%, respectively, depending on the decrepitation behavior of the ore used. The yield of +2 mm sponge iron produced is on the order of 60 to 62% of the ore charged, and the recovery of iron into sponge iron is 90 to 92%. The extent of nonmetallics in the finished product is restricted to less than 2%. Table 3.11 shows the size distribution of the magnetic and the nonmagnetic fractions of the cooler discharge product. As is evident from the screen analysis, the ore and coal do not always undergo extensive degradation inside the kiln. Under stable kiln operating conditions, the extent of ash and char accompanying sponge iron has been found to range between 20 and 30%. The presence of 50 to 52% carbon in the coal char compared with 44 to 46% in the feed coal (Table 3.12) signifies proper operating conditions inside the kiln. The waste gas leaving the kiln contains 2 to 3% carbon monoxide, which is subsequently burned in the postcombustion chamber so that no unbumed carbon monoxide is released to the atmosphere. The dust loading in the waste gas is 40 to 60 g/Nm^, and after cleaning in the waste gas cleaning system, the dust content is decreased to less than 300 mg/Nm^. The typical quality of sponge iron produced by the TDR process is shown in Table 3.13. Metalization of the product is controlled to meet customers’ requirements and can be main­ tained at any level between 90 and 95%. The total iron content in sponge iron is governed by the metalization and by the level of impurities in the iron oxide feed. Since only high grade iron ore is processed, the normal range of total iron in sponge iron is 90 to 93%, the gangue content is 5 to 7%, and sulfur is controllable in the range of 0.020 to 0.030%. The phosphorus content of DRJ is entirely dependent on the phosphorus present in the ore and can range between 0.040 and 0.100% in the case of most Indian ores.

Coal-Based DR Processes Using Rotary Kilns

35

TABLE 3,9 Physico-chemical Characteristics of Iron Ores Used at Ipitata Source

I

2

3

66-67 . - 1 .8 2.0-3.4 0.04-0.05

6 6 -6 8

0.7-1.3 1.3-2.1 0.03-0.05

65-67 . - 2 .0 2.3-2.9 0.04-0.06

3 -9 7 9-90 3-11 2 -4

4 -1 0 84-91 2-5 1.5-3.0

2 -9 80-90

1.0

0.8

1.2

1.5

1.9

2.8

+6.3 mm -5 0 0 um

95 3

93 4

94 5

Bulk density, t/m^

2.5

2.3

2.4

Chemical analysis, % Fe SÌO2 A JP3 P

1 2

1 6

reen analysis, mm + 2 0 -2 0 - 8

and and

+ 8

+5

-5

2 -1 2

3-5

Reducibility, % per min

-dt1/4o 4 Nm-^ 100 kWh I4 t 550 NmVmin 1.5 kg/cm^ 900°C 10 g/Nm-^

Outputs Productivity Fe-Cr (50% Cr, 34% Fe. 5% C) Slag^ Top gas volume Caloric value Top gas pressure Top gas temperature Fe-Cr temperature Heat for decomposition Sensible heat of Fe-Cr Sensible heat of Slag Sensible heat o f top gas Heat loss Chrome ore analysis, %: CaO MgO

67 i/mVd 100 t/d 1630 kg 11000 Nm^ 5.02 MJ/Nm-^ 1.1 kg/cm^ 700°C 1600°C 1.67 GJ 12.55 GJ 3.35 GJ 55.65 GJ 4.18 GJ

SiOp

AI2 O,

S

P

Cr/Fe

5.0

13.0

0.020

0.020

2.2

45.0 0.1 Coke analysis.. % VM Ash

10.0 P

FC

23.2 0.9 ^ Slag analysis. % CaO SÌO 2

0.15

7 5 .5

MgO

AI2 0 ,

32

10

27

31

Beyond the Blast Furnace

142

Coal

Concentrate

Oxygen

^ ^ a s t e gas

\ Slag

Flash

smelting Prereduction

1900

V

1450

Hot metal F IG U R E 7.12.

/

r El-smelting Final treatment

i

El , Connection to grid

[NRED process — two stage reduction in one single reactor.

power for oxygen production and internal consumption.^' This is illustrated below for a suggested 50 tph INRED plant. The preconditions are Iron concentrate, hematite Coa] moisture Coa] ash, dried sample Net heating value, dried coal sample Steam pressure Steam superheated to Feed water temperature Consumption Coal, dry weight Oxygen, 95% O 2 . 5 bar gauge Electrical energy for the furnace Electrical energy for auxiliary equipment Electrical energy for oxygen plant Electrical energy produced in the condensing turbine generator

65% 7% 15% 28 MJ/kg 45 bar gauge 450°C 1 10°C 620 kg/t 680 mVt 300 kWh/l 100 kWh/t 280 kWh/t 680 kWh/t

The plant can be connected to the grid, but electrical energy must be purchased

fo r

start -

up.

3. Operation The data for a 50 tph unit were projected from the figures obtained in a 5 tph plant which was operated during 1978. It has been reported that• •

• • • • •

Fine grained concentrates of iron ore and different kinds of iron oxides can be smelted without difficulty. Use of materials with grain sizes up to 100 mesh has actually been tested. Coals with high ash and sulfur can be used; coking grade coal is not required. It is possible to flash smelt without forming a large circulating load of dust. The waste heat boiler did not experience any clogging in the prototype unit. The boiler tubes in the flash smelting chamber did not experience much wear. Process control can be achieved by using measuring and analyzing equipment as well as a computer. Once started, the INRED process can be stopped and restarted with minimal loss of time.

Smelting Reduction Processes

143

4. Metallurgical features The most important metalJurgicaJ features of the INRED process include the following: • • •

•

• •

Normal carbon analysis is 3.5%, and the carbon content of hot metal is a function of the size of the coke bed in the electric furnace; Normal silicon analysis is 1%, but this can be altered by varying the size and tempera­ ture of the coke bed; Approximately 70 to 80% of the sulfur is eliminated as sulfur dioxide during flash smelting. The remaining sulfur is distributed between slag and metal with a partition coefficient of 25 when the slag contains less than 1% Fe; Arsenic, lead, zinc, alkali, and to some extent, phosphorus, are eliminated in the waste gas. Build-up of these contaminants in the dust is prevented by the dust bleed in the final phase of gas cleaning; The circulating dust load in the system is 5% of the amount of concentrates, lime, and ash entering the system. Loss of iron in the normal dust bleed is 1.5 to 2.0%; Cyanides are not traceable in the waste gas, and NO, is below the level allowed for coalfired plants in the United States.

5. Process features The INRED process is characterized by the following: • • • • • ® •

Rapid transfer of large amounts of thermal energy to the charge with high thermal efficiency Ability to use high ash coals and lean metallic raw materials Heavy purge of sulfur, allowing use of high sulfur coal and ore High tolerance level for arsenic, lead, zinc, and alkali Low volume of gas generation, which facilitates gas cleaning and sulfur dioxide re­ moval Simplicity of reactor design using proven components and few moving parts Use of a conventional submerged arc electric furnace with predictable low wear rate of lining, ensuring low maintenance cost

The INRED process has the potential to be utilized at lower investment and production costs than those of some of the conventional processes. The estimated capital cost for a 50 tph INRED plant was reported (in 1984-1987) to be $60 million dollars (US) including 20 million dollars (US) for the oxygen plant. At that stage, the production cost of hot metal in an INRED unit was estimated to be 136 dollars (US) per ton. Despite the various attractive claims in its favor, the process has not yet been commercial­ ized, and its future is open to question because precise control of a rapid, flash smelting process is always likely to be difficult.

E. ELRED This process, advocated by ASEA and Stora Kopparbergs Barglags in Sweden, is designed to produce pig iron by two stage reduction of iron ore concentrates and/or waste oxide with coal.^^-^^ In the first stage of the process, the concentrates or waste oxides are prereduced with coal powder to a partly reduced metallic product, which also contains carbon, and in the second stage this product is reduced and smelted to pig iron in a DC electric arc furnace. Both the concentrates and the coal can thus be used directly, without pretreatment in the form of sintering plants or cokemaking. Figure 7.13 shows a schematic flow sheet of the ELRED process. The flue gases from both stages shown can be used to generate electrical energy. Owing to the thermodynamic equilib-

Beyond the Blast Furnace

144

Air Water

F I G U R E 7.13.

Schematic flowsheet of the ELRED process.

rium conditions and the fact that equilibrium cannot be fully achieved, more electrical energy is obtained from the process flue gases than is consumed in the final reduction stage. A small surplus is therefore always available, which can be introduced to the grid. The net energy consumption projected for the ELRED process is 15.5 GJ/thm at an efficiency of 33%. The process allows the use of coal of different grades, ranging from anthracite to lignite, but whichever coal is used, it must be ground to 0.2-0.3 mm before it can be injected into the bed. The ore concentrates also require a fine-grained size, less than 0.1 mm, with an iron content of over 65%. This process also has the capability to use a suitable mixture of various flue dusts from iron and steel works, and it is claimed that by using such waste oxides as the feed material, even small scale ELRED plants of 120,000 tpa can prove economical. In this process, prereduction takes place under pressure in a circulating fluidized bed, which differs from the traditional fluidized bed primarily in terms of the use of substantially higher gas velocities. The excess of carbon in the bed and the high gas velocity prevent sticking. The vigorous material circulation in the bed results in uniform temperature distribution. Coal powder and air are injected directly into the bed, generate the required reduction temperature of 950 to 1000°C, and supply the (CO + H2) mixture as well as excess carbon to the process. Therefore the process does not require a separate plant to generate the reducing gas. The reducing and fluidizing gases, formed as a result of the partial combustion of coal, result in intensive mixing within the bed. The flue gas from the reactor carries with it a large amount of bed material to a cyclone system, where the solid material is separated and then returned to the lower part of the reactor. This system also includes an arrangement for continuous charging and preheating of the concentrates. After dust, water, and carbon dioxide have been removed, 30 to 50% of the gas can be utilized for fluidizing in a part of the reactor. A prereduced, partly metalized product, having a controlled carbon content, is continuously discharged at the bottom of the reactor. The degree of metalization of the product can be controlled at a level of 60 to 70% by adjusting the residence time and temperature in the reactor. Final reduction takes place in a DC arc furnace, where a hollow carbon electrode located at the center of the furnace roof is used to carry the prereduced feed along with suitable fluxes. The materials quickly pass through the hot plasma melting, carburizing, and reducing zones with a good yield, starting with the feed charged at 700°C. A liquid heel of 30% is maintained for restarting. The liquid iron contains 3 to 4% carbon and approximately 0.05% silicon and

Smelting Reduction Processes

145

Iron ore

Coal

Dolo mite

Li__: -- ----- iu

Direct reduction plant

;l / r n

-Off !

char

Wesfe and

Ash ¡Elecfric power generation

ORI

1Ch Energy generation plant

Electric energy

X iff

gas

I/' SAP

-Slag

^

Slag

Smelting plant

jHotl

metal.

d ff

gas

Metallurgical treatment

Steel

F I G U R E 7.14.

Steelmaking plant Scrap

Principles of the CO MB IS MELT process.

manganese. Roughly half of the sulfur and most of the phosphorus in the feed report to the hot metal. Low silicon and manganese in the liquid iron expected to be produced in the ELRED process make it suitable for “slagless steelmaking”. The energy content in the flue gases from the prereduction and final reduction stages can be extracted in a combined power plant, comprising a gas and a steam turbine. The surplus power generation, over and above the requirement of the ELRED system, is expected to be 300 to 400 kWh/t of iron. An economic evaluation of the ELRED-BOF (LD) route has been made on paper consid­ ering BF-BOF and DR-EAF routes, for a capacity of 0.5 mtpa of crude steel. For the ELRED route, it was assumed that only 70% scrap would be consumed due to low silicon and manganese in the pig iron, but 18% scrap recycling was assumed in the other cases. It showed that the ELRED-BOF could be 30% less expensive than the DR-EAF route. While the pilot plant work confirmed the raw material requirements, the local conditions, especially energy and raw materials costs, became important factors, and the overall economy of the process varied markedly from country to country. The future of the ELRED process is most uncertain, since further process development work has now been suspended. F. COM BISMELT CO MB ISMELT is a new iron and steelmaking technology jointly developed by Lurgi and Mannesmann Demag of G e rm a n y .F o r the production of hot metal, COMBISMELT com­ bines the SL/RN coal-based DR process with the submerged arc furnace (SAF) for open slag smelting. Hot DRJ produced in the SL/RN kiln is smelted in the SAF, utilizing the electrical power generated from the heat of the kiln waste gases and the residual char. Heat recovery from the waste gas is, in most cases, sufficient to generate about 70% of the electrical energy required by the SAF. The constant electrical load required by the SAF makes this link technically feasible. This concept allows the production of hot metal of adjustable quality, containing 0.1 to 2.0% C, which can be used for steelmaking. The basic flow sheet illustrating the process principles is shown in Figure 7.14. Depending on the type of coal used in the DR kiln, the heat recovery resources are different, as is shown in Figure 7.15. About 25 to 50% of the input energy is lost in the off-gas, and up to 25% may be lost with the kiln discharge material, thus giving a total heal recovery potential of approximately 50%. This indicates that the heat recovery system must be selected, keeping in view the type of coal used. In the case of lignite, the heal recovery system will be mainly

146

Beyond the Blast Furnace Input 4.é

Input 475

24 %

Anthracite coal

F I G U R E 7.15.

Input 415

12 % Bituminous coal

1% Lignite coal

Heat recovery possibilities in the COMB IS MELT process for different coal grades.

based on the utilization of waste gases, while if anthracite coal is used for DR, the heat recovery system should cater to the burning of the kiln discharge material as well as to the utilization of the kiln off-gas. The sensible heal in the kiln discharge material can be recovered by feeding it directly into an electric furnace. To be able to do this, the application of a highly reactive, low ash coal is a prerequisite. The sensible heat of DRl would correspond to approximately 200 kWh/l at a charging temperature of about 600°C. With low reactivity coals containing high ash, the kiln discharge should be cooled down for magnetic separation of DRl from the nonmagnetics. In such a case, only the chemical heal in the kiln discharge may be utilized by burning the excess carbon to produce steam. It is reported that the circulating fluid bed (CFB) system gives the best thermal efficiency, and therefore this is the most efficient practice for producing steam. With the CFB system, energy up to 1 100 kWh/t iron can be recovered. With typical Indian Singareni coal, power generation up to 865 kWh/t can be achieved; additional power may be generated by adding fresh coal to the CFB. In general, the capacity of the in-plant energy recovery system is determined by the number of rotary kilns installed. If two rotary kilns are coupled with one electrical power generating unit, the power level that can be supplied depends on the operational conditions of the kilns. If both the kilns are in operation, the upper constant power level will be attained. If one of the kilns is down for maintenance or other reasons, the lower power level will be attained. It should be mentioned that the electrical energy generated by the in-plant energy recovery system is supplied at a relatively constant power level. Furthermore, the resulting short circuit capacity of this type of power plant is relatively low, if no external mains are available for connection. This emphasizes the need to search for a tailor-made smelting concept indepen­ dent of additional external power facilities. In order to utilize the electrical energy generated in the combined SL/RN-CFB system to the optimum extent, an electrical type furnace should be applied, featuring constant power consumption units in combination with a low short circuit capacity of the mains. This ideal smelting unit is the submerged arc furnace (SAF). In a SAF with an open slag bath and increased slag height, the conversion of electrical energy into thermal energy mainly takes place in the horizontal direction by resistance heating of the slag, and more flexible adjustment of the active power load of the furnace becomes possible. Furthermore, an important advantage of this open slag smelting technique is the lower current-voltage ratio, resulting in a power factor as high as 0.95. The SAF process can also accept DRl with lower metalization, and hence, variations in the degree of metalization do not give rise to operational difficulties. The slag volume with increased slag height represents an important heat accumulator. The raw material charged into the slag starts melting immediately, and the chemical reaction rates are increased. The charging of the material is preferably done batchwise into the open slag bath, so that the batch charged is immediately enveloped by liquid slag before it starts melting. This also enables the

Smelting Reduction Processes

147

Iron ore Slag formers

F I G U R E 7.16.

Schematic flowsheet of the PLASMAS MELT process.

furnace to be operated with a high proportion of fines in the charge. The high slag volume, together with excellent mixing of the components, permits metallurgical reactions to approach equilibrium closely. The overall economics of the process would depend primarily on the type of coal used, its ash content, and its charring characteristics. A typical kiln discharge product in India (kilns operating with high ash coals) usually contains about 78% sponge iron, 2% charry materials, and 20% ash. Assuming that the sponge iron portion of the kiln discharge contains 90% total iron with a metalization of 90 to 92% and that the char contains no volatile matter and around 50% each of fixed carbon and ash, the kiln discharge would have a total iron content of 64 to 65%, with almost 30% gangue and 1,0 to 1.2% carbon. This type of material, containing such a high amount of gangue, cannot be fed directly into a SAP for the production of suitable quality hot metal. G. PLASMASMELT Plasma, sometimes referred to as the fourth state of matter, is an ionized gas comprising molecules, atoms, ions (in their ground or in various excited states), electrons, and protons. A plasma is electrically quasineutral, and plasmas may be generated by passing an electric current through a gas. The history of plasma technology is essentially as old as the history of electricity, but the current plasma technology was developed toward the end of the 1950s. However, the application of plasma technology in iron and steel metallurgy is of more recent origin. The major problem earlier was that the plasma generators were not developed to the point where they could be used in continuous operation with satisfactory efficiency and electrode life in an industrial environment like a steel works. SKF Steel’s (Sweden) PLASMASMELT process, developed around 1975, is designed to optimize energy utilization in hot metal making. Here, iron ore is prereduced in two fluidized beds by reducing gas generated at the final smelting reduction stage. Prereduced ore and coaJ fines are injected into a coke-filled shaft furnace where smelting reduction takes place in a raceway in the coke column. Figure 7.16 shows a schematic flowsheet of the PLASMAS MELT process. Iron ore fines are mixed with fluxing agents and preheated in a fluidized bed preheater, using gas from the smelter or the fluidized bed reactor where ore is prereduced by contact with off-gas from the shaft smelter. The prereduced ore is fed by pneumatic injection into the shaft smelter along with the pulverized coal reductant. Fluxing agents can also be introduced here instead of along with the ore in the preheater. The shaft furnace is charged at the top with coke

Beyond the Blast Furnace

148

Eltctricity, kWh

Degree of prereduction, 7 ®

FIGURE 7.17. prereduction.

Energy consumption in the PLAS MAS MELT process as a function of oxygen removal during

that forms a temperature resistant reaction chamber in which smelting takes place. Plasma generators mounted in tuyeres near the bottom of the shaft smelter provide the required thermal energy. The plasma gas for the plasma generators is produced from clean and recompressed process gas from the smelter. The gas temperature in the plasma generator is between 3000 and 5000°C, but as a result of the strongly endothermic reactions, it drops rapidly outside the generator to 1700 to 2000°C in the actual reduction zone. Liquid iron and slag are collected at the bottom of the shaft and are tapped in a manner similar to that in a blast furnace. The gas leaving the smelter consists primarily of carbon monoxide and hydrogen, a part of which is fed to the prereduction stage, after cleaning in a hot cyclone and cooling to about 850°C. A small portion of this gas is also cooled, cleaned, and compressed for use as plasma gas and for injection of prereduced ore and coal. The main reducing agent is coal. The functions of the coke in the shaft or final reduction furnace are to form a reduction chamber which is permeable to gas and liquid and capable of withstanding high temperatures and to prevent furnace wall refractory wear. Coke is also required to ensure reducing conditions along the refractory walls, to compensate for minor fluctuations in the feed rate of the reducing agent and to ensure a uniform carbon content in the hot metal. Ore fines are reduced to 50 to 60% at 750 to 800°C in the prereduction furnace. When the gas leaves the prereduction stage, it still contains approximately 10 to 15% carbon monoxide and hydrogen, and can be used for drying and preheating the ore concentrate. Figure 7.17 shows the energy consumption of the PLASMAS MELT process as a function of the extent of oxygen removal during prereduction.^^ When electricity is expensive, it is preferable to replace electric heating, partially or wholly, by the combustion of coal with oxygen in the final reduction stage (PLAS MAS MELT OXYGEN). In practice, it is expected that energy con­ sumption of about 11.7 GJ/thm can be attained with 50% prereduction, which is lower than the typical specific energy consumption of around 14 GJ/thm in a blast furnace (Table 7.2). The hot metal produced has low sulfur and normal silicon and manganese contents, which can be adjusted to the desired values. The slag is well reduced and therefore the iron yield is high. After completing the pilot tests in a 1.5 MW plant in 1981, a half-scale plant of 18 MW ( 3 x 6 MW plasma generators) with an annual capacity of 40,000 to 50,000 thm, based on steel works waste products, was built. This variant of the process is called PLASMADUST. The plant is located at Landskrona in Southern Sweden and began operation in 1984. In 1986, a plant for the production of 80,000 tpa ferro-chromium was started at Malmo, also in southern Sweden. Because the PLAS MAS MELT process uses approximately half the amount of fossil fuel required by a blast furnace, it is possible to produce hot metal with a much lower sulfur

Smelting Reduction Processes

149

TABLE 1 2 Energy Consumption in PLASMAS MELT Compared with a Blast Furnace

Blast furnace PLASMASMELT, coal PLASMASMELT, oil

Coke, kg

Coal, kg

Oil, kg

390 75 75

75 200

—

—

140

—

Electric power, kWh

GJ

105 1120 1080

14.23 11.78 11.50

N o te : The coal consumption is based on a coal containing 9% moisture and 10% ash. Energy for sintering is included for the blast furnace.

content. The process is so designed that it utilizes energy optimally, although the electricity consumption is rather high. Based on certain assumptions regarding coal quality and electrical energy, the cost of producing hot metal by the PLASMAS MELT process is claimed to be 25% lower than that of the conventional blast furnace route. In locations where electricity is expensive, this process may be a disadvantage, but it is very flexible in energy utilization, and there is the possibility of producing fuel gas as a byproduct, as in most other SR processes. The fuel gas produced is a mixture of carbon monoxide and hydrogen with a heat value of 11.7

MJ/NmL The limited amount of coke consumed in the PLAS MAS MELT system can always be purchased. Because it uses a lower gas volume per ton of hot metal, the furnace and most equipment are expected to be proportionally smaller than in the case of a blast furnace. Significant capital investment and operating costs can thus be saved compared with the blast furnace. In countries with cheap hydroelectric power and in industrialized countries with normal price ratios between electricity and coal, the process could have good prospects.

H. KRUPP-COIN A combination of a converter for smelting and a shaft furnace for prereduction is a possible means of producing liquid iron with lower energy consumption, using only coal as the source of energy. The COal INjection (COIN) process developed by Krupp is one such process, in which coal fines with low to medium volatile content are used as the source of energy.^^'^'^ Partial combustion of coal to carbon monoxide and hydrogen is brought about by injecting coal and oxygen into a converter type melting vessel. The resulting gas causes violent turbulence in the bath, giving rise to a high melting rate of sponge iron and scrap which are continuously charged from the top of the vessel. A premelt is produced, which is used for steelmaking. The oxygen necessary for the partial combustion of coal is fed into the bath through nozzles at the bottom of the converter. Fine coal is pneumatically injected through annular nozzles arranged concentrically around the oxygen nozzles, and the coal stream also serves to protect the oxygen nozzles from bum-off. Hot gases (CO + H2) from the vessel are used for fine ore reduction in a circulating fluidized bed; alternatively, these gases can also be used as an industrial fuel, depending on location. The basic flowsheet of the process is shown in Figure 7.18.^^ The converter is equipped with a gas-tight hood, and all equipment is designed to hold one bar over-pressure. Hot gas produced in the reduction vessel possesses sufficient pressure to penetrate it through the prereduction furnace. In the case of highly volatile coals, the fine coal is first injected into the hot off-gas, leaving the vessel at about 1400°C, where it is charred and cools the gas to 1000°C. The char is mostly separated in a hot cyclone before the ore reduction unit and transferred to the coal/oxygen injection system. The reducing gas then passes through the steps of the ore reduction system before it is finally cooled to about 500°C by fresh, cold ore. In this step, the greater part of

150

Beyond the Blast Furnace 400 Nm^O.

UO kg Stmi-Coke

40 kg lime . 1------— — Mtlting Convtrter 8S kg C

0 .181 Slaa

4

400 kg Coal OOV.Volatntj 0.9t Iron

1 t DPI fO%|

[ Rtduction of ort finti [ Gas volume ( Nm^) %CO % H , V.CO, V.HjO Gas utilization 0 1000 80 I 20 “ h s t t p ;F e0-^ F t| 56

1000

15

24

5

29 %

37

8

45%

64

10

74%

2 Step: FcjOj-t-FeO 43

1000

1.4 t Concentrate“

12

3 Step:C-precipitat

840

I I

12

I 13

Offgas with 0.5 Gcal/tDRI

FIGURE 7 .1 8 .

Flowsheet of the COIN process using Fine ore reduction in a circulating fluidized bed.

the residual carbon monoxide decomposes to carbon dioxide and carbon, thus giving an off­ gas that need not be recycled and saving the carbon dioxide scrubber. The carbon formed is mainly on the ore surface and can either serve as a reductant or eventually go into the meltergasifier. The product from the melter-gasifier will be intermediate between steel and hot metal: 1% carbon, 0.1 to 0.2% sulfur with low silicon and manganese. The consumption figures per ton of liquid steel, i.e., after refining, as estimated by Krupp, are as follows: M a teria ls

Fine ore Scrap Lime Coal (30% VM), dry Oxygen Electrical power Refractories for furnace and ladles

Am ounts 1400 kg 160 kg 95 kg 600 kg 400 Nm^ 150 kWh 14 dollars (US)

The off-gas from the reduction unit is of low caloric value; therefore, a gas credit should not be given. Investment costs are indicated as 240 dollars (US)/annual ton of liquid steel, which includes the reduction furnace, the melter-gasifier, the refining vessel, and the coal and gas supply system. The COIN process thus foresees a way of converting ore concentrates with coal to molten iron, which is the ultimate goal of all smelting reduction processes.

I„ PCIG FOR SMELTING REDUCTION A method to gasify coal in a molten iron bath was introduced toward the end of the 1960s by the Applied Technology Corporation (ATC), Pittsburgh, Pennsylvania, and was followed

Smelting Reduction Processes

151

by a number of similar developments in the iron and steel i n d u s t r y . T h e PCIG (Pressurized Coal Iron Gasification) process originated from the development of a new coal-based smelting reduction method of ironmaking. The process was developed jointly by Interproject Service AB (IPS) of Sweden and Nippon Steel Corporation (NSC) of Japan in a pilot plant at the Metallurgical Research Station (MEFOS) in Lulea, Sweden. In the coal gasification process using a molten iron bath, the gasifier holds molten iron at a high temperature (1400 to 1600°C), and pulverized coal is blown into this molten iron with oxygen and steam as the gasification agents. The iron bath is violently agitated by the action of an oxygen Jet blown onto the surface of the bath. This agitation mixes the coal and the molten iron so that the coal is instantaneously cracked by the high temperature of iron to liberate hydrogen and carbon monoxide. The reaction is enhanced by oxygen and steam. The coal ash fuses and floats on the surface of the molten iron, while the sulfur in coal dissolves in the molten iron and is transferred to the molten slag. The process of gasifying coal using a molten iron bath has a number of applications. It can be used in metallurgical processes for heat and power generation and for producing ammonia, methanol, and other industrial chemical products. It is possible to produce liquid iron using the gas from the gasifier for prereduction of iron ore and, using the gasifier itself, for smelting of the prereduced material. Another possibility is to make DRI, where this gas is used only for reduction; in this case, it is necessary to shift the gas to a hydrogen rich gas if DRI with a high degree of metalizad on is to be produced. The smelting furnace is charged with prereduced iron, coal, and oxygen. When oxygen alone is used as the gasification agent,a large amount of surplus heal is generated in the bath. The coal which is added is used for reduction of prereduced iron, for generating additional heat in the melt, and for generating reducing gas for the prereduction. These characteristics make it favorable to apply this process in smelting reduction, since the main reaction occuring during smelting reduction is

Fc203 + 3C

2Fe 4- 3CO,

(7.2)

A W = -443 kJ/mole Fc203

The prereduction furnace consists of a two stage bed which is charged with preheated iron ore fines, reducing gas from the smelting furnace, and in some cases oxygen for maintaining the heat balance. The energy balance in the smelting furnace (SF) improves if a part of the carbon monoxide and hydrogen are burned to carbon dioxide and water in the SF itself, with a high heat transfer efficiency from the gas to the metal bath. The oxidation degree (OD) of the generated gas in the SF also influences the reduction potential of the reducing gas for the prereduction furnace (PF). It may be favorable to remove the water and carbon dioxide in the SF gas in order to minimize the amount of export gas from the process. Considering the above points, three different cases (A, B, and C) were chosen for a comparative evaluation.'^ The basic principles of these cases are explained in the schematic representation in Figure 7.19. In case A, the coal, to a large extent, is used for heating the bath (87 to 93%); a minor part is used for reduction of the prereduced material. The total coal consumption and the coal for reduction show an irregularity at low degrees of prereduction, due to the change in the ratio of coal for reduction and coal for heating. The coal added generates large amounts of slag and dust, which decrease the iron yield in the process and make material handling problematic. The dust, which represents about 80% of the total iron loss, may have a sulfur content that could preclude recirculation. It is, therefore, important to

152

Beyond the Blast Furnace

• OFF-GAS •DUST STEAM

Case B FIGURE 7.19.

Case C

Schematic diagram of the PCIG system.

know whether the gas for prereduction must be cooled down and cleaned to remove the dust (if possible in the hot stage), or whether it may be injected raw into the prereduction furnace without any pretreatment. The SF unit should be operated at an internal pressure of 3 to 5 bar to allow for direct transfer of gas to the prereduction furnace. The amount of surplus gas generated in the iron bath and not used for prereduction varies markedly with the degree of prereduction. Some part of this energy can be used downstream inside the plant (5 to 10 GJ/t), but the total economy of the process is highly influenced by the possibility of obtaining credit for the surplus or byproduct gas sold outside the plant. The net energy consumption is strongly influenced by the degree of prereduction of the iron ore and is minimal at high degrees of prereduction. Thus, the energy balance in the smelting furnace will be influenced by the degree of prereduction, as a result of which the amount of coal for heating must increase with the degree of prereduction. This will increase heat losses in the gas, slag, and dust phases, and the surplus gas volume will increase. Hence, a process based on case A, with an exact gas balance or with a small amount of byproduct gas used downstream in the plant, requires iron prereduced to a very high degree, which imparts high demands on the prereduction furnace, especially if it is a fluidized bed type of furnace. Case B describes two different ways to optimize the process. One alternative is to apply post combustion in the SF, where the heat transfer is influenced by factors such as degree of oxidation, temperature of the melt, agitation of slag and bath, slag type, etc. The net energy consumption increases with an increase in the degree of postcombustion. Another way to decrease the coal consumption further is to recycle the off-gas from the prereduction furnace, in order to utilize the gas more effectively by removing carbon dioxide from it.

Smelting Reduction Processes

153

In case C, which is based on a smelting furnace designed for a very high degree of postcombustion, the reducing gas for prereduction is produced in a separate coal gasifier designed for injection of coal, oxygen, and slag formers into a molten bath at above atmospheric pressure. Gas cleaning, i.e., carbon dioxide removal equipment, is not utilized in this case, and surplus gas is generated in the smelting furnace, if needed, together with additional gas from the gasifier. Case C manifests a high process flexibility compared to the other two cases. The advantages are the following: •

The condition for reduction and gasification can be optimized based on the design of the reactors. The internal pressure in the reactors can be different and controlled separately. Slag/metal composition and amount can be optimized for the special conditions in each unit. The composition of the energy bearing materials can be different; i.e., there is an option in choosing the energy bearing materials.

• • •

The rationale for case C is to utilize post-combustion as a method of increasing the capacity of the smelting furnace. A comparatively small amount of coal is necessary to maintain the heat balance if high postcombustion degrees and high energy transfer efficiencies are to be obtained.The gas which leaves the smelting furnace has a low heating value and can be utilized downstream in the plant or mixed with the gas from the gasifier and utilized outside the plant. The gasifier is pressurized and works mainly as a generator of gas for the prereduction furnace. The reactor pressure is sufficient for direct injection of gas to the prereduction furnace, which improves the energy consumption in the process. The net energy in case C is lower than that in the other two options, even though excess gas must be produced. Another advantage of using two reactors is that the specific productivity in the reduction unit can be increased because of the smaller amount of coal needed for heating in this unit. The different cases mentioned earlier were studied and simulated with a calculation model, and the results are given in Table 7.3. The analysis shows the following: 1.

2.

3.

Case A: The results show that 500 kg of coal will generate more than 1000 Nm^ gas (dry) with a composition of 65% carbon monoxide and 35% hydrogen. Steam is added as a coolant for the melt, and iron ore is added to compensate for the iron loss. The gas can be used as a reduction gas, a feed stock for the chemical industry, or for generation of electric power. Case B: The results show that 1.032 t hot metal is produced with 630 kg of coal and 543 Nm^ of oxygen. The nitrogen content in the off-gas from the prereduction furnace is limited to 2%, and the gas is partly recirculated after purification from carbon dioxide and moisture. Case C: If the smelting reduction and gasification are separated as in case C, the total coal consumption will increase somewhat for the same production as in case B; but the consumption of utilities such as oxygen and steam for carbon dioxide removal will decrease.

The flexibility of the process in the production of hot metal and gas is very important, depending on the location and the local conditions. The principle of the process makes it possible to utilize PCIG as a coal gasifier and/or ironmaking process, even if a particular plant has a limited range of the gas to iron ratio. The process developers claim removal of sulfur with slag and dust under steady state operations. Many other impurities in coal, such as Zn, Pb, Cd, and As, are also expected to be removed from the gas phase as dust.

154

Beyond the Blast Furnace TABLE 1 3 Comparison of Various Approaches in the PCIG Process Case A Material

Case B

PCIG

SF

500 33

632

Case C

PF

SF

PCIG

PF

404

256 17

1597

Input Coal, kg Iron ore, kg Prereduced iron, kg Lime, kg Dolomite, kg Oxygen, Nm^ High pressure steam, kg

—

27 41 232 99

—

1403 62 32 361 —

__ 1568

91

1372 55 13 251

—

—

14 21 1 19 51

1032 209 18 669

62 14 523

— — —

___

—

— —

—

__

___

38

Output Hoi metal, kg Slag, kg Dusi iron, kg Gas volume, Nm^

___

122 28 1022

1032 266 29 1079

__ — —

1974

— —

523

Gas C om position, % CO 65.8 Hj 33.0 CO, H ,0 N, PRD, %

51.4 9.6

0.7 —

41.1 15.1 22.0 16.2 0.8 —

36.3 5.4 31.0 10.8 2.0 37.0

65.8 33.6 36.3 21.2 0.8

—

—

—

16.4 8.2 —

0.7

47.5 27.3 0.7 50.0

J. HISMELT PROCESS The Hlsmelt process, which is a bold step toward the development of an efficient smelting reduction process, utilizes a bath under highly nonequilibrious conditions. It closely resembles the Japanese National Project, Direct Iron Ore Smelting (DIOS) and American Iron and Steel Institute (AISI) processes, on which work is progressing. Figure 7.20 schematically represents the basic outline of the processes, and some of its features are as follows:^^'^^ With a horizontal smelt reduction vessel, it is possible to achieve extremely high bottom injection rates and acceptable containment of the liquids in the reactor. A circulating fluidized bed allows recovery of a portion of the chemical potential and sensible heat in the off-gas through prereduction and preheating of fine iron ore and generation of high pressure steam. An iron bath is utilized as an intensive reaction medium for the rapid reduction of ore and dissolution of coal. The bath acts as a thermal and chemical buffer for the process and plays a pivotal role in heat transfer from the combustion processes occurring in the top space. Bottom coal injection maximizes the recovery of carbon to the bath, releases volatile components in the coal, and generates the required level of bath turbulence. This practice facilitates more efficient heat transfer, faster reduction of the ore, and sufficient mixing with the slag to ensure low FeO levels. Hot air-based operation facilitates simultaneous high postcombustion and appreciable heat transfer efficiency. Substantial quantities of inert nitrogen ensure transfer of sensible heat

Smelting Reduction Processes

155 Process offgas

FIGURE 7.20.

• •

Hlsmelt — simplified process flowsheet.

without reaction and limit the postcombustion flame temperature, which may otherwise cause excessive refractory wear. Pressurized operation eliminates the ingress of air and permits an increase in process intensity. The off-gas is partially utilized by post-combustion to preheat the air in regenerative heats. Sensible and chemical heat in the remaining off-gas can be utilized for materials preparation, steam raising, and other site-related operations.

In the Hlsmelt technology, two-stage processes have been considered, incorporating a coalinjected smelting reactor with the off-gas passing to a prereduction unit where the iron ore can be reduced by the carbon monoxide and hydrogen in the off-gases. Initially, attention was paid to the 20 to 30% postcombustion range, which enabled the use of highly volatile coals and, at the same time, provided a level of prereduction (30 to 50% metalization) compatible with economic coal consumption. It was found that improved coal consumption could also be achieved by incorporating an intermediate stage, in which the smelter off-gas would be upgraded by coal or hydrocarbon injection. This so-called hot gas reforming stage enabled the utilization of high post-combustion in the smelter while still obtaining a good level of metalization (40 to 50%) in the prereduction stage. The additional consumption of fuel/ reductant in the hot gas reforming stage was more than offset by the reduction of coal consumption in the smelting stage. Such a three stage process could offer considerable potential in terms of coal consumption and the ability to use highly volatile coals. Alterna­ tively, a carbon dioxide scrubber could be incorporated to obtain recycled gas enriched in carbon monoxide to be used during prereduction.^^ High degrees of metalization and high preheat temperatures are required for processes such as COREX, which operate with lower degrees of postcombustion because very little excess energy is available for smelting. An increase in postcombustion substantially improves the energy balance and permits smelting of less reduced or colder feed materials. It is claimed by the developers of the Hlsmelt process that an increase in postcombustion level from 0 to 60%, coupled with 80% heat transfer efficiency, can result in a seven-fold increase in the energy available for smelting. CRA of Australia and Klockner Stahlwerke of Germany Joined together to study the process in a small-scale pilot plant at Maxhuette, Germany, from 1981 to 1987. in 1987, CRA assumed all responsibilities for further development and formed a new Joint venture with Midrex Corporation in late 1988. The new venture achieved simultaneous high postcombustion and high heat transfer efficiency by applying the principle of high top and bottom injection

156

Beyond the Blast Furnace

FIGURE 7.21.

Hlsmelt process heat balance.

rates coupled with numerous novel and radical modifications to the configuration of the vessel, the process, and the flowsheet. So far, the Hlsmelt process has achieved post combus­ tion levels of about 60% with 85% heat transfer efficiency. On the basis of the pilot plant results, the heat balance for a commercial plant is expected to be as shown in Figure 1 .2 \P As a next step, Hlsmelt Corporation of Perth, Western Australia, the manager of the joint venture between CRA and Midrex Corporation, has planned to invest $100 million dollars (US) toward the construction of an advanced research and development facility on the site of an old blast furnace in Kwinana, South Perth, and the plant is expected to come into operation in 1993. K. REACTOR STEELMAKING This process, developed by Daido Steel Company, Japan, is based on nonelectric scrap melting technology.^"^ The process compares well with the KS or KMS methods advocated by Klockner, Germany, the EOF method of KORF, and the cupola method, in terms of versatility and energy needs. The process uses a cold iron-bearing charge, along with noncoking coal and oxygen as the three raw material constituents. The Reactor Steelmaking Method (RSM) improves the KS process by providing better heat efficiency in the furnace, as the waste gas combustion heat is utilized to preheat the burden. It also has an advantage over the EOF method in terms of a higher postcombustion ratio, up to 50%. The process has an edge over the cupola method because RSM has fewer restrictions on the size and quality of the feed. The RSM process can be used to make liquid pig iron or steel of any desired carbon level. The molten iron produced in this process can be charged, along with scrap, into an electric furnace. The primary purpose of this process is to utilize low cost primary energy resources effectively, and the secondary purpose is to have a flexible operation in terms of the type of cold charge. To bum carbon fuel as perfectly as possible in the melting furnace, an iron bath type melting reactor is used because its higher combustion rate of carbon fuel. The technical details and essential features of the process are shown in Figures 7.22 and 7.23. An experimental reactor of 1.01capacity with a maximum scrap melting rate of 0.5 t/h and a pilot scale 5 0 1 reactor have been used at Daido Steel; they report the following experiences: 1.

At the maximum scrap melting rate of 0.5 t/h, the process retains 0.3 to 0.5 t of molten iron in every tapping. In the lower part of the furnace, the bath is made shallow in order

Smelting Reduction Processes Basic idea

FIG URE 7.22.

157 Technique for embodiment

Essential factor

Conception of the development of the RSM process.

Scrap

I -

FIGURE 7.23.

2.

Schematic illustration of the new Reactor Steelmaking process.

to increase the interface area of molten iron and high temperature gas. Carbon monoxide is generated continuously by blowing carbon powder and oxygen into molten iron from the bottom of the furnace. Oxygen is also blown into the furnace through the ceiling, for postcombustion. A higher carbon yield is obtained in a shorter time by injecting carbon powder of particle sizes less than 0.4 mm into the molten iron, compared to carbon addition onto the iron bath surface. The postcombustion ratio increases as the top blowing oxygen ratio increases. However, if oxygen is supplied by top blowing only, considerable slag foaming occurs; this leads to unstable operation and lower heat efficiency. The top blown oxygen ratio is controlled within a range of 0.6 to 0.7, and the value of CO 2 KCO 2 + CO), under this condition, is about 0.5.

158

Beyond the Blast Furnace Melting of

\ \

Scrap U

U

Scrap IH I

Preheating Pourlnj

jLf»PPln9j|. Tippins 0 d

^ Scrip FIGURE 7.24.

Improvement o f the thermal efficiency in the new Reactor Steelmaking process.

The rate of heat absorption in molten iron increases with an increase in the oxygen supply rate and time. For an oxygen rate of 700 and 400 Nl/min through the top and the bottom respectively, postcombustion of 0.59, carbon consumption of 0.74 kg/min, furnace heat generation of 17.2 MJ/min, and molten iron heat absorption of 6.3 MJ/min are obtained, giving a heat efficiency of 37%. Figure 7.24 shows the heat balance after sufficient heat accumulation in the furnace is accomplished. The heat efficiency is given as a percentage of the heat generated by complete combustion of carbon, and increases with an increase in the oxygen supply rate and also with scrap preheating, eventually reaching 32%. This heat efficiency compares well with a targe scale electric furnace with 40% efficiency of coal-to-electricity generation and 80% efficiency for melting in an arc furnace, giving a total heat efficiency of approximately 30%. The heat loss due to absorption by the furnace refractory decreases in proportion to the reactor scale-up; higher heat efficiency is expected in a full scale reactor. The furnace design shown in Figure 7.23 has a lower circular cross section to make the bricklaying work easier. It has a sliding gate provided at the bottom of the furnace, enabling tapping without tilting the furnace and to supply oxygen even when tapping, in order to increase productivity. Hot waste gases introduced into the preheating chamber are cooled as they pass through the scrap, and the temperature of waste gas and scrap near the damper is reported to be below 500°C. Figure 7.24 shows the heat flow in the scrap melting process using pure graphite as the heat source. With an average scrap preheating temperature of 550°C, the heat efficiency goes up to 42%. An approximate comparison shows that the energy consumption of the 0.5 t reactor is about 25% less than that of a large scale UHP arc furnace, the reactor requiring 3.2 GJ/t versus 4.3 GJ/t by an EAF. Yield of iron is 94.8%, compared with 96% in an EAF; substantial (up to 4.5%) losses take place due to splashes deposited on the furnace wall and dust. While the major consideration of energy saving and resource utilization have been quan­ tified, some additional problems will come up in a reactor during practical operation. A theoretical calculation for a 50 t reactor (Figure 7.25) shows the possibility of 57% heat efficiency, mainly accruing from fewer losses through refractories and 650°C scrap preheat­ ing. Sensible heat recovery from the waste gases, after scrap preheating, is also an interesting additional possibility.

Smelting Reduction Processes

159 Sensible heat of scrap

F I G U R E 7.25.

Heat flow estimation for a 50 t/h Reactor Steelmaking unit.

F I G U R E 7.26.

Conceptual diagram of the AlSl-H yL Direct Steelmaking process.

L. AlSl-HyL DIRECT STEELMAKING The American Iron and Steel Institute (AISI) and the Iron and Steel Division of the Mexican Company HYLSA are collaborating on the development of an in-bath smelting reduction process designed to eliminate coke ovens and batch processing in the BF-LD route, and at the same time, to produce steel.^^ The process is essentially direct smelting of preheated, prereduced iron ore with coal and oxygen in a liquid iron bath. Figure 7.26 shows a conceptual diagram of the process. Iron ore and coal are added to a bath of liquid iron, while oxygen is blown from the top and bottom of the vessel. The heat given off during the reaction preheats the raw materials and ensures the formation of liquid steel at the required temperature. The coal reduces the iron oxide, producing steel and carbon monoxide, and the oxygen lowers the steel to the specified carbon content and postcombusts carbon monoxide to carbon dioxide. Research has indicated the possibility of a self-sustaining process to produce quality steel at lower production and capital costs than those involved with conventional processes. Present studies are concerned only with the pilot plant level, but the collaborators are aiming to make the process commercially viable within this decade.

M, FERRUM LIQUID PHASE RECOVERY (FLPR) PROCESS^" Recently, the Moscow Institute of Steel and Alloys (CIS) claimed to have developed a single stage smelting reduction process for the production of liquid iron. In this process, -20-mm coal and iron ore are charged in a rectangular smelting vessel operating at near

160

Beyond the Blast Furnace

atmospheric pressure. The reduction of iron oxides takes place through a foamy slag phase. A very high degree of postcombustion (nearly 70%) is achieved by oxygen lancing through side tuyeres, located at two levels along the longer wall of the reaction vessel. The bottom row of tuyeres is placed in such a way that the slag-metal layer is not disturbed and the appropriate oxygen potential across the depth of the bath is also maintained. This process concept has been tested at a 350,000 tpa pilot plant scale at Novoliptskin CIS, but unfortunately, no other details are presently available.

Ne DIRECT IRON ORE SMELTING (DIOS) PROCESS In an effort to develop a flexible route for liquid iron production using non-coking coal, the Japan Iron and Steel Federation (JISF), along with eight Japanese integrated steel producers and the Coal Mining Research Centre of Japan, began in 1988 a US$ 95 million R&D program that is expected to last seven y e a r s . W o r k on most of the fundamental studies of various process steps, individually and in combination, has already been completed; after establishing a basic understanding, the construction of a 500 tpd pilot plant at the Keihin Steel Plant in Japan is presently underway and is expected to be completed in September 1993. Operational tests are expected to begin in October 1993. The main units of the DIOS are a fluidized bed (FB) for prereduction, an iron bath smelting reduction furnace (SRF), and an outlet gas reforming unit, where fine coal is added to the gas generated in the SRF, to enable effective utilization of the gas while optimizing the total DIOS process system in relation to the operational conditions. For achieving the optimum surplus energy supply (4-8.5 GJ/thm) with a relatively low fuel consumption, an in-depth analysis of the system has been carried out. Accordingly, three types of system combinations were selected and the amount of surplus energy and fuel consumption were calculated depending on the operational conditions of each sub-process, i.e., the degree of postcombustion, the degree of prereduction, and the gas reforming efficiency. In the single-combination process, the production of molten iron is completed in the SRF itself, subject to a postcombustion ratio of 55% or more. This system is feasible only when the CO generated in the SRF undergoes a large degree of postcombustion and the heat transfer efficiency is high. In the double-combination process, gases generated in the SRF are utilized for preheating and prereducing the iron ore to achieve improved energy effi­ ciency in the system as a whole. In this mode, the amount of gas generated in the SRF must match the gas requirements of the PRF. Compared to a single stage operation, this configu­ ration can become viable with a lower postcombustion ratio as well as with a high degree of reduction. In the triple-combination process, the sensible heat of the gases generated in the SRF is utilized to reform the gases by adding fine coal, whereby an enhanced reduction potential of the gases is achieved, resulting in improved energy efficiency. Analyzing the above process options of DIOS, it has been found that the double- and triple-combination processes, with postcombustion in the range of 40-60%, would be most suitable in terms of energy optimization.^^ The basic studies on DIOS began with a simulation model, followed by an assessment of the basic performance conditions necessary for the system constitution. Later, small and large scale facilities were utilized to achieve the basic conditions that must be established. All these test results formed the basis of construction for the 500 tpd pilot plant for further investigation. Table 7.4 illustrates the role of the participating agencies in the complete scheme of the work.^^ A 5-ton smelting reduction furnace at the Kashima Steel Plant has been used to study various aspects of the process. It is equipped with side tuyeres, to blow oxygen into the slag layer for more vigorous stirring and for improving the heat transfer efficiency of the gases in the slag layer. The postcombustion ratio improved by 5-10% when employing side blowing of oxygen.^^ Similarly, by bottom blowing nitrogen, the heat transfer efficiency increased but the extent of postcombustion was found to decrease. On the basis of these findings, the stirring

Smelting Reduction Processes

161

TABLE 7.4 Share of the Elemental Studies for the DIOS Process Test Facility and Condition

Testing Place Sakai Steel Plant, Nippon Steel

A 100-ton iron bath smelting reduction furnace (inner vol. 138 m^) Batch operation under normal pressure Oxygen blown at top and bottom Amount o f oxygen: 35,000 Nm^/h

Fukuyama Steel Plant, NKK

A 5-ton iron bath smelting reduction furnace directly connected to a pre­ reduction fluidized bed Operating pressure: max. 1.9 kgf/cm^ Oxygen blown at top and bottom Amount of oxygen: max. 2,500 Nm^/h Tapping system available

Test Target Achieving a high P.C. and a high heat transfer efficiency Extracting problems involved in an up-scaled furnace, and examining the necessary countermeasures Examining the possibility o f using coal with high volatile matter Clarifying the in-fumace behavior Examining a suitable gas reforming technology Achieving a high P.C. and a high heat transfer efficiency Examining the possiblity o f using coal with high volatile matter Achieving prolonged, consecutive operation with the PRF directly connected to the SRF Examining the steady operational conditions of the fluidized bed Assessing the reduction ability o f the fluidized bed

Bubbling fluidized bed: 1 m x 8 m high Kashima Steel Plant, Sumitomo Metal Industries

A 5-ton bath smelting reduction furnace Batch operation under normal pressure Oxygen blown at top, bottom, and side

Achieving a high P.C. and a high heat transfer efficiency Examining the possibility o f using coal with high volatile matter Assessing the effect of injecting oxygen into the slag bed from the side

Amount of oxygen: max. 1,400 Nm^Ai Chiba Steel Plant, Kawasaki Steel

A circulating fluidized bed (0.7 m X 7.3 m high)

Examining the steady operational conditions o f the fluidized bed Assessing the reduction ability o f the fluidized bed

Kobe Steel Plant, Kobe Steel

A gas reforming test facility composed of a carbonizing furnace and a plasma heater

Looking for a gas reforming technology

intensity was optimized to achieve a high postcombustion and a high postcombustion effi­ ciency simultaneously. Coals from various sources with volatile matter contents of up to 40% have been tested and found to be suitable for the DIOS p r o c e s s . T h e results obtained so far have been very encouraging and the developers believe that the pilot plant trials will establish the technology for commercial implementation of the process in the future.

O. OTHER DEVELOPMENTS Tests with the circulating fluidized bed of the prereduction of ore fines are being conducted in Australia. A process proposed by KHD Humboldt Wedag, based on the smelting in an iron

162

Beyond the Blast Furnace

bath reactor, was tested at Lulea, Sweden, and later abandoned. The three stage process included prereduction to FeO in a circulating fluidized bed followed by meltdown in the smelting cyclone and reduction to a carbon-bearing premell in a horizontal/cylindricaJ iron bath reactor. Developers in Japan rely on charging of coal and partially reduced materials into a BOFlike converter. NKK is developing a converter type SR process for ironmaking. Kobe has constructed a flowsheet involving prereduction to FeO in a MIDREX type shaft furnace and SR with high levels of postcombustion. Recently, in Japan, a national seven year program sponsored by MITl to develop an SR process has taken off. Klockner/CRA, Voest Alpine, Krupp, BSC/Hoogovens, NKK, Kawasaki, Kobe, NSC, and Sumitomo are all working to­ gether to develop an SR process using primary energy with an optimal degree of prereduction and postcombustion and low carbon consumption. [MIS, Mexico, is working on a direct steelmaking route to produce steel from a 100% solid burden that is either scrap or iron ore (sponge iron). The process is expected to have tremendous potential for decreasing production costs as it relies only on coal, iron ore, and scrap as raw materials and uses minimal electricity.

V. CURRENT STATUS AND FUTURE OUTLOOK OF SMELTING REDUCTION Out of all the SR processes, COREX is the only one that has reached the stage of commercial exploitation, and a 300,000 tpa plant was commissioned in December 1987.* Except COREX, none of the other processes has gone beyond the concept stage, though many have operating pilot plants of various capacities, which are used to test raw materials and collect data for technical and commercial upscaling of the process in question. Many countries have considered these processes, but no industrial set-up has so far been established anywhere except the COMBISMELT process, which, as an essentially DR-EAF route combination, has recently been tried in New Zealand. Klockner and CRA are presently working on the Hlsmelt process, where the next development stage envisaged is a demonstration plant of 0.2 mtpa capacity.** One of the greatest anomalies for most of the SR processes is that the cost of the auxiliary facilities (oxygen plant, power plant,etc.) is more than the main unit itself. It was found in 1987 that for a 300,000 tpa COREX plant, consisting of the COREX plant proper, oxygen plant, and power plant, the capital cost estimate would be 77, 29, and 93 million dollars (US), respectively, i.e., a total of 199 million dollars (US). Taking the costs involved outside the battery limits and allowing for a contingency fund of 15%, the total cost is 240 million dollars (US), as shown in Table 7.5. The net works cost of hot metal produced by such a COREX plant, if the credit to the gas is given in terms of fuel equivalent (for blast furnace gas and coke oven gas), would work out to 65 dollars (US)/thm versus 103 dollars (US)/thm for hot metal produced from a blast furnace (under typical Indian conditions), as is revealed in Table 7.6. The major impact on the cost of hot metal produced is the credit given for the CO REX gas in terms of electric power. This difference arises since 17.15 GJ of gaseous fuel is available per ton of hot metal produced by the COREX process, versus 6.9 GJ equivalent of gaseous fuel, 1.21 GJ equivalent of tar, and 3.5 GJ equivalent of solid fuel (coke breeze, nut coke, and pearl coke) produced by a conventional blast furnace. From the capital cost figures given, it can be seen that the major cost component of 93 million dollars (US) is for a power plant, and this is based on the use of gas turbines and *

The second commercial COREX plant, with a rated capacity of 600,000-700,000 tpa is being built in Pohang, South Korea, and is expected to be ready in 1994-1995. ** Active research is presently underway concerning the Reactor Steelmaking Process and the AlSI-HyL Process, but neither has yet reached the commercial exploitation stage.

Smelting Reduction Processes

163

TABLE I S Capital Cost of a Typical 300,000 tpa COREX Plant (in Million Dollars (US) — 1987 Estimates) COREX unit 12.6 37.1 7.7

1. 2. 3.

Engineering Equipment (imported) Other equipment

4. 5.

Civil and structural Utilities

6.3 1.4

6. 7.

Erection supervision License fee

9.8 2.1

Oxygen

Power

2.1 25.2 (cost included in power plant) 0.7 (cost included in power plant) 0.7

6.3 75.6 3.5

2.1

—

77.0

28.7

9.

Total cost of CO REX + Oxygen + Power plant Other costs @ 5% on Item 8

=

10. 1 1.

Subtotal of 8 + 9 Contingencies 15%

= 208.6 = 31.5

12.

Total cost

= 240.1

8.

4.2 1.4

93.1

198.8 9.8

TABLE 7,6 Typical Operating Cost of a 300,000 tpa COREX Plant (in Dollars (LS) — 1987 Estimates)

Raw materials Fuel Labor Stores Transfers Credit gas

Blast Furnaces

COREX

74.0 0.1 3.0 1.0 17.0 -(5.0)

64.3 3.0 18.0 -(19.0)

Total cost above

16.2

2.1

Total works cost Credit for slag

106.3 -(3.0)

68.4 -(3.0)

Net works cost

103.3

65.4

conventional boilers. Since the power plant is tied in to the "export gas” generated from the COREX process, one of the major bottlenecks is that when the COREX plant is down, no power is available, unless provision is made to use alternative materials for generating an equivalent quantity of power in such an eventuality. Work on the INRED process was stopped by Boliden as they experienced some problems with control of the flash smelting reactor. Similarly, applicability of ELRED in countries other than those where electricity is relatively inexpensive has now been accepted as being ex­ tremely limited. PEAS MAS MELT also faced some initial problems in the zinc volatilizing unit, a problem that has reportedly been resolved, but the future of this process is still uncertain.

Beyond the Blast Furnace

164

SR has significant potential for the production of ferro-alloys. In fact, the Japanese are working on various smelting reduction processes quite similar to the COREX process, but the main aim of these efforts is not to make hot metal for steelmaking but to use these processes as alternatives for the energy-intensive submerged arc process normally used for making ferro-chrome and other ferro-alloys, which command higher prices. Kawasaki Steel Corpora­ tion claims to have already developed such a process for making ferro-chrome. Considering the relatively high specific energy consumption in the ferro-alloy industry and the rising demand for ferro-alloys in making special steels, this could be an important outlet for SR, particularly in developing countries.

REFERENCES Von Norten, F. and Raymaekers, J., Sei. A m .. June 1988. Eketorp, S. and Wijk, O., Proc. Ini. Conf. on New Smelting Reduction and Near Net Shape Casting Technologies for Steel, RIST, Pohang, Korea, 1990.

Edstrom, J. O., Ma, J., and Scheele, J. V., N o rd ic Steel M in in g Rev., 2(SuppL), 1990. Delport, H. M. W., Iro n m a k in g S teelm a king, 19(3), 1992. Kepplinger, W., Maschlanka, W., and Wallner, F., P ro c. C O R E X S ym posium , Delport, H. M. W. and Holaschke, P. J., Eds., South African Institute of Mining and Metallurgy, Johannesburg, 1990, 17.

Hank, R., P ro c. C O R E X S ym posium , Delport, H. M. W. and Holaschke, P. J., Eds., South African Institute o f Mining and Metallurgy, Johannesburg, 1990, 21. Papst, G. and Flickenschild, J., Iro n S teel E ng.. 63, February 1986.

Kepplinger, W., Maschlanka, W., and Wallner, F., P ro c. C O R E X S ym posium , Delport, H. M. W. and Holaschke, P. J., Eds., South African Institute o f Mining and Metallurgy, Johannesburg, 1990, 14.

Hank, R., P ro c. C O R E X S ym posium , Delport, H. M. W. and Holaschke, P. J., Eds., South African Institute of Mining and Metallurgy, Johannesburg, 1990. Hank, R., P ro c. C O R E X S ym posium , Delport, H. M. W. and Holaschke, P. J., Eds., South African Institute o f Mining and Metallurgy, Johannesburg, 1990, 34. 11. Hank, R., P ro c. C O R E X S ym posium , Delport, H. M. W. and Holaschke, P. J., Eds., South African Institute of Mining and Metallurgy, Johannesburg, 1990, 23. 12. Hauk, R., P ro c. C O R E X S ym posium , Delport, H. M. W. and Holaschke, P. J., Eds., South African Institute o f Mining and Metallurgy, Johannesburg, 1990, 29-30. 13. “The CO REX process: operational experience and potential for the iron and steel industry," Eisenhuttentag 1990, Dusseldorf, Germany, November 15-16, 1990. 14. Delport, H. M. W., Vuletic, B., and Kastner, R. W., P ro c. C O R E X S ym posium , Delport, H. M. W. and Holaschke, P. J., Eds., South African Institute of Mining and Metallurgy, Johannesburg, 1990, 49. 15. Kepplinger, L. W., Proc. Int. Conf. on New Smelting Reduction and Near Net Shape Casting Technologies for Steel, RIST, Pohang, Korea, 1990. 16. Chatterjee, Amit, Misra, B., Datta, K., and Krishnan, S. H., Smelting Reduction Technology — Processes and Possibilities under Indian Conditions, Internal Report, R&D Division, Tata Steel, December 1989. 17. Hatano, M., Miyazaki,T., Yamaoka, H., and Kamei, Y., Proc. 5th Iron and Steel Congress, Washington, D.C., 1986. 18. Innes, J. A., Keogh, J. V., Cunningham, B. C., Philip, D. K., and Batterham, R. J., S E A JS IQ ., April 1990. 19. Iro n Steel E ng., 60, May 1983. 20 . Elvander, H. L, Edenwall, 1. A., and Hellestam, S. C. J., Iro n m a k in g S teelm a king. 5, 1979. 22 . Bengtsson, E. and Widell, B., Iro n S teelm aker. 8, October 1981. 23. Collin, P. and Stickler, H., Iro n S teel In t., 53, April 1980. 24. Pauls, H.-R. and Fritzsche, K.-D., Paper presented at the SL/RN Seminar, New Delhi, India, 1985. 25. Herlitz, H. G., Johansson, B., and Santen, S. O., Iro n Steel E ng., 61, March 1984. 26. Hartwig, J., Neuschutz, D., Radke, D., and Seeling, H. F., Iro n Steel E ng .. 59, February 1982. 27. Hartwig, J., S ta h l U n d Eisen, 100( 10), 1980. 28. Steel Tim es In t., 8, June 1984. 29. Radke, D., Proc. 3rd Int. Iron and Steel Congress, Chicago, 1978. 30. Axelsson, C. L., Torssel, K., Sato, K., and Torneman, B., Proc. Int. Conf. on Alternative Routes to Iron and Steel under Indian Conditions, Jamshedpur, India, 1988.

10.

Smelting Reduction Processes 31. 32.

165

D ire c t fro m M id re x , 16(4), 3rd Quarter, 1991. Keogh, J. V., Hardie, G. J., Philip, D. K., and Burke, P. D., Washington, D.C., 1991.

Iro n m a k in g C onf. P ro c. (IS S ). Vol. 50,

33. Innes, J. A., Moodie, J. P., Webb, 1. D., and Brotzmann, K., Process T e ch n o lo g y C o nference P ro ce e d in g s, 1988. 34. Sugiura, S., Fujita, S., and Demukai, N., Trans. IS IJ , 28, 1988. 35. Steel Tim es In i., 13(6), 1990. 36. Aeron, S. M., Chaudhuri, P. K., Gupta, S. K., and Mukherjee, A. K., Proc. Conf. “Applications of the Latest Technological Innovations and Processes for the Production of Iron, Steel, and High Quality ProductMix”, M a te ria ls W eek, 1992. 37. Steel Tim es In te rn a tio n a l, January, 1992. 38. Inatani, T., Iro n m a k in g C onf. P rocee din gs, ISS, V o l. 50 , Washington D.C., 1991. 39. Kanamori, K., and Saito, N., S A E IS I Q u a rte rly , April 1992.

Chapter 8

IMPORTANT FEATURES OF SR PROCESSES L PROCESS CATEGORIZATION In terms of the source of energy, SR processes can be divided into two groups: 1. 2.

Coal/electricity-based processes (INRED, ELRED, PEAS MAS MELT, COMBISMELT, etc.) which utilize gas generated from coaJ for reduction and electric power for smelting. AIJ coaJ processes (COIN, CO REX) which use coal exclusively, for both prereduction and smelting.

Based on their thermochemicaJ features, alJ SR processes can be characterized as follows:' »

•

‘

Single stage coal-oxygen processes employ the simplest reactor configuration, i.e., a single well stirred vessel in which the ore charged at room temperature, fuel, and oxygen react (Figure 8 .1). These types of processes are extremely inefficient and are not technically or economically attractive for the production of iron unless adequate credits can be obtained for the large amount of surplus gas generated during reduction. Thus, such a process would be viable only in special circumstances, as a synergistic iron and gas producer. Two stage coal-oxygen processes are an improvement on single stage processes and utilize the me Iter offgases for prereduction, as shown in Figure 8.2. These processes can become self-regulating if partial metalization of ore is permitted. To match the perfor­ mances of the two unit processes, namely reduction and melting, proper control of the melting process can ensure that the off-gas fulfills the requirements of producing highly metalized DRl. But the energy consumption in the gas recycle operation is significantly higher than the once-through process, in which the top gas energy can be used effec­ tively. Three stage coal-oxygen processes are those in which the interposition of a reaction zone containing free carbon between the melting and reduction zones provides a means of reducing the temperature of the me Iter off-gas by converting the excess sensible heat in the gas to chemical energy through two highly endothermic reactions: C -H CO, H ,0 -H C

2CO

( 8 . 1)

H, -H CO

( 8 .2 )

The configuration “a” shown in Figure 8.3 is essentially that of the KAWASAKI and the PLAS MAS MELT processes. Configurations “b” and “c” are the COREX and the SUMITOMO processes, respectively, while “d" depicts the BSC Oxy-Coal blast fur­ nace process. The incorporation of the third stage results in a system with potentially the lowest energy consumption, but the major energy inefficiencies in this case are associ­ ated with the production of coke used in process configurations “a ’, “c”, and “d". Electric melting — coal-based reduction processes have been proposed, in which energy is transferred to the melting stage by internal generation of electricity from the reduction off-gases, as illustrated in Figure 8.4. This allows higher flexibility than does the oxycoal processes, but is also more energy intensive than the two- and three-stage pro­ cesses.

Beyond the Blast Furnace

168

Coal

Ore

Gài

I

I Reduction

XT" Slag

FIGURE 8.1.

H etil

Single stage smelting reduction process.

Ore

FIGURE 8.2.

I Melting

Gas

Two stage smelting reduction process.

Coke

Slag

(a)

FIGURE 8.3.

Metal

Slag

Metal

lb}

Slag

Metal

(cj

Slag

Metal

(dj

Three stage smelting reduction process (four alternatives).

Table 8.1 depicts various process developments based on the above mentioned fundamen­ tal classification. ' It will be seen that most of the developments are confined to the groups of two- and three-stage processes, and in both cases, coal-oxygen combustion has been exten­ sively used.

Important Features of SR Processes

169

sits FIGURE 8.4.

Electric melting — coal-based reduction process.

TABLE 8.1 Classification of SR Processes Energy source

Process category Single stage (smelting) Two stage (meltingreduction) Three stage (meltinggasificationreduction)

Electric melting-coaJbased reduction

Coal-oxygen com bustion

External

Electricity (separately or in com bination with coal-oxygen)

PLASMASMELT

PIROGAS

BSC-ICl KRÜPP, BSC,

KOBE, HOOGOVENS,

CIG COREX, KAWASAKI,

SUMITOMO, BSC oxycoaJ blast furnace ELRED,

CSFRED

II. KINETIC ASPECTS OF SMELTING REDUCTION Reduction of iron oxide has been a subject of many studies because of its obvious importance in the iron and steel industry. This reaction, in one form or another, plays an important role in most of the commercial iron and steelmaking systems in use today, and it also forms the basis of a number of alternative process developments which are on the verge of commercial exploitation.

170

Beyond the Blast Furnace

Most of the work done so far, however, concerns solid iron oxides with the reaction temperature only occasionally exceeding 1100°C, and studies on the reduction of iron oxiderich molten slags are limited. Presumably, this is because of experimental difficulties involv­ ing molten iron oxide, since it reacts with practically all types of crucible materials and ultimately gives rise to uncertainties in the results. Further, the extremely fast reaction rates involved lead to difficulties in monitoring the kinetics. Nevertheless, the reduction of iron oxide-rich slags is now being seen in a new perspective with the emergence of SR processes. In all direct iron and steelmaking processes, a substantial proportion of iron oxide enters the slag phase and is reduced by carbon particles (solid carbon or coke) or by carbon dissolved in iron in contact with the slag. Volatiles generated from the coal charged may also help in reduction, and therein lies the basic importance of the reduction of iron oxide-rich slags by various reducing agents. The kinetics of converter-based smelting reduction processes for iron production are to some extent different from those of the blast furnace. One significant difference is the more intense stirring, which together with the slag influences the reduction; another is the heat distribution because of postcombustion, which influences the temperature in the converter. Yet another interesting aspect is the behavior of the iron oxide itself. In a single stage smelting reduction process, iron will exist in different oxidation stages, ferrous and ferric, and this, together with the heat distribution, will lead to different slag conditions within the converter. Stirring together with slag properties such as basicity, viscosity, and FeO activity are the major factors responsible for the rapid reduction obtained. The salient features of reduction of liquid iron oxide or iron oxide-bearing slags by various reducing agents are highlighted below.^

A. REDUCTION BY SOLID CARBON The reduction of pure molten FeO and slags containing FeO above 80 wt% by solid carbon is a zero order reaction. Reduction of slags with 20 to 80% FeO is a first order reaction, whereas second order rate kinetics apply when FeO levels are below 20%. The reduction reactions of molten FeO by solid carbon take place according to the following two steps: 1.

Direct reduction of molten FeO by solid carbon: FeO(l) + C(S) = Fe(S) + CO(g)

2.

(8.3)

Indirect reduction of molten FeO by carbon monoxide and the Boudouard reaction: FeO(l) + CO(g) = Fe(S) + C02(g)

(8.4)

C02(g) + C(S) = 2CO(g)

(8.5)

It has been proposed that Reaction 8.3 occurs along the interfaces among the gaseous, liquid, and solid phases.^ Reactions 8.4 and 8.5 play an important role in smelting reduction. Diffusional transport of FeO and convection of heat have been suggested as the rate control­ ling steps during reduction of slags containing less than 80 wt% FeO;^"^ the high value of activation energy (160 kJ/mol) supports the hypothesis that the Boudouard reaction is the rate controlling step, a view shared by many.^ Recently, Basu^ has shown that the reduction of iron oxide-rich slags by solid graphite powder obeys the nucléation and growth model proposed by Johnson and Mehl:

da

= k" -t"-' •(! - a)

( 8 .6 )

Important Features of SR Processes

171

or, In [l/(l-a)l = (Wn) • r

(8.7)

where a = fractional reduction, k = constant, min n = constant, and t = time, min. The reduction reaction follows a typical sigmoidal pattern, where, after overcoming the initial nucléation delay, the rate remains practically constant, and again toward the end, the rate becomes sluggish.

B. REDUCTION BY CARBON DISSOLVED IN LIQUID IRON The reduction rate of molten FeO by solute carbon in liquid iron plays a major role in SR processes. The total reaction is apparently of first order, which tends to be a second order reaction at FeO levels of below 20 wt%. The diffusion of carbon in liquid iron is the principal rate controlling factor when the carbon content in liquid iron is above 0.2 wt%, and the reaction is partly controlled by fusion and decomposition of FeO. Lloyd et al."^ showed that the decarburization reaction between the slag and the metal takes place mainly at the interface: FeO(l) = Fe(l) + O (slag)

( 8.8 )

C (solute) + O (slag) = CO (g)

(8.9)

Wagner^ postulated that the reaction between carbon saturated liquid iron and an FeObearing slag to evolve carbon monoxide may be an electrochemical half cell reaction at the interface: Fe^'' (slag) + 2e” = Fe°

( 8 . 10)

C (solute) + O^ (slag) = CO (g) + 2e"

( 8 . 11 )

This hypothesis has also been supported by many others.^^ '® It has been shown that the reduction temperature and initial iron oxide content in the slag enhance the reduction kinetics considerably, while basicity (Ca0 /Si 02 ) was found to reach an optimum value. The kinetics of the reaction improve up to a basicity value of around 1.5, beyond which the rate again becomes slower.^ Because all smelting reduction reactions, with the exception of Reaction 8.4, are endothermal, it is very important that heat is supplied continuously to the bath. High heat generation within the molten bath as a result of coinjection of coal and oxygen, or for any other reason, is the most efficient way to accelerate the kinetics of smelting reduction. C. REDUCTION BY CARBON MONOXIDE This reaction plays a major role in smelting reduction. The rate of reduction is proportional to the square root of the flow rate of carbon monoxide passing through the melt. Mass transport of the gaseous phase is generally accepted as the rate-controlling step, but Grieveson'^ proposed interdiffusion of Fe^+ and through the slag as the rate-controlling step. Information regarding reduction of FeO-rich slags by carbon monoxide is also available, but not much has been conclusively proven about the reduction mechanism. Recently Graenzdoerffer'^ proposed that a first order rate law is followed, and arrived at the following rate equation:

172

Beyond the Blast Furnace pre-reduction, •/•

pre-reduction, */•

FIGURE 8.5.

Optimal prereduction in a two stage smelting reduction process.

R/Ao = exp (-32300/RT - 1.37) • (l.O - 0.7

J ( 8. 12)

• (apeo • Pco - ^Fe ' Pco, /k ). tTiol/cm^ • sec The activation energy was found to be 135.0 kJ/mol. The rate of reaction can be increased by having a large reaction surface and a thin boundary layer. Use of fine powdery ore and coal, with the gas well dispersed into small gas bubbles, results in large reaction surfaces, and intense stirring breaks down the boundary layers. Such fundamentaJ studies often lose their relevance if information regarding other impor­ tant parameters, such as degree of postcombustion (PCR), degree of prereduction (PRD) of the iron feed, reduction behavior with different combinations of reductants, slag foaming during smelting, behavior of actuaJ raw materials in the converter, etc., are not available. Direct use of primary energy instead of electrical energy, with simultaneous gasification of coal, requires a combination of final reduction and melting in or directly above the iron bath, with the use of the gases produced for prereduction. The carbon monoxide requirement for prereduction and the evolution of carbon monoxide during smelting reduction can be harmonized in a two stage process, as is shown in Figure 8.5. This leads to a combined smelting reduction process with an optimal degree of prereduction; e.g., in one case, the optimal working condition is 75% prereduction and a carbon consumption of around 450 kg/t Fe is expected.'^ Bath smelting processes are attracting intense attention from process developers. The most proven process is CO REX, where the gas from the melter-gasifier, which is primarily carbon monoxide and hydrogen with no postcombustion, is used to prereduce pellets/lump ore to about 90% metalization. This group of processes, in which there is very little postcombustion and a high degree of prereduction, includes the Kawasaki SR process and the BSC-FIoogovens process. In other bath smelting processes, coal, partially reduced iron ore, and oxygen (or air) are allowed to react in an iron bath. The iron bath and possibly the slag act as the medium for faster reactions, and the carbon monoxide and hydrogen from the reduction reactions are postcombusted to provide up to two-thirds of the total energy requirements. Processes being developed under this category include Hlsmelt and AlSl.

173

Important Features of SR Processes

m\

E

FIGURE 8.6. CoaJ consumption for varying degrees of postcombustion (PCR) and prereduction (PRD) (heat transfer efficiency = 85%, heal loss = 10%).

TABLE 8.2 Material Requirements for Various Combinations of Degree of Prereduction (PRD) and Degree of Postcombustion (PCR) PRD, %

PCR,

Coal,

Oxygen,

kg

Nm-’

C aO , kg

Ore,

%

30 60^ 9Qb

50 30 0 50

537 500 688 697

444 381 451 697

104 101 88 118

1492 1488 1458 1495

0^

kg

Additional 55 kg o f carbon required for reforming for the prereducer. 1% silicon in metal. No preheat o f ore, all others at 800°C.

An optimal choice of the degree of prereduction and the postcombustion ratio significantly affect the process energy requirement. In order to exploit the chemical energy of the me Iter gas, postcombustion is essential; but a high degree of postcombustion increases the oxygen potential of the gas and the reduction is hampered. Therefore, prereduction and the smelting step are best kept separate; sometimes the slag layer itself can act as the separator. One typical plot illustrating the dependence of coal requirement on PCR and PRD is shown in Figure 8.6. It is often felt that an attractive process would be reduction to FeO only in the prereducer with up to 50% PCR. Such a process would be energy efficient, would have almost as much excess energy as a blast furnace, and would require a very simple prereducer. Table 8.2 shows the calculated material requirements for some typical levels of postcombustion and prereduction. The ideal concept of an SR process that directly uses fine ore and coal, and involving only a liquid state reduction to yield a liquid product similar to steel has not yet been realized, in spite of intensive research efforts around the globe. Such a process would be characterized by carbon and oxygen injection into liquid iron oxide so that the upstream carbon monoxide is converted into carbon dioxide, and the iron is almost completely decarburized in the liquid ore. Among other difficulties, process engineering problems lie in the area of refractories and in the design of the reaction chamber with respect to the introduction and transport of energy and materials, as well as in the control of phase contact of carbon and liquid ore and in the phase separation of the slag and the liquid ore.

174

Beyond the Blast Furnace

IIL IMPORTANCE OF SLAG FOAMING IN SMELTING REDUCTION Smelting reduction, in most of its variants, involves reduction of an iron oxide-rich slag in the liquid state by either solid carbon or carbon dissolved in liquid iron. A considerable proportion of unreduced iron oxide along with coal/char descends to the smelting vessel and then is reduced. In bath smelting SR processes, for example, iron oxide and coal are charged into a liquid iron bath. Cracking of coal and reduction of iron oxide-rich slags constitute the major reactions, and because of these reactions, a large amount of carbon monoxide and hydrogen are generated and become primarily responsible for slag foaming. To achieve higher production rates, the rate of gas generation also becomes higher, and hence, violent foaming becomes inevitable. Slag foaming has a crucial role to play in any SR vessel. Although excessive foaming hampers normal operation, foaming helps in many ways if kept under a critical value; it shields the metal from the atmosphere and also acts as a thermal insulator for the process heat. On the other hand, a foamy slag has a direct impact on heat transfer from the postcombustion flame to the metal bath. Also, because of the generation of high surface area in a foamy slag, many of the chemical reactions proceed favorably. Previous studies on slag foaming were confined mostly to the effects of various process parameters (e.g., temperature, basicity, chemical constituents, etc.) on foam stability, foam height, etc. Various systems were studied, including Ca0 -Si02 ,‘^ CaO-FeO,'^ Si02 -Fe0 ,'^ Ca0 -Si02 -Fe0 ,*^ etc. Recently, Ito et al.’^ studied the foaming behavior of the Ca0 -Si02 -Fe0 slag system, the most important in SR processes, and tried to quantify the foaming phenom­ enon. They have shown that (refer to Figure 8.7):

0 §

(8.13)

A

and. 8 =

Ah AV"

(8.14)

where V ’ = superficial gas velocity, cm/sec, 5 = average gas traveling time or foaming index, sec, Qg = gas flow rate, cmVsec, A = cross sectional area of the container, cm^, and Ah = change of slag height, cm. Again, g

= £•V

8

(8.15)

where e = void fraction, the volumetric fraction of gas and V = actual gas velocity, cm/sec. Now, h = £-L

(8.16)

where h = foam height, cm and L = foam layer thickness, cm. So, finally, 8 =

AL AV

(8.17)

Important Features of SR Processes

FIGURE

8 .7 .

175

F o a m in g in a sla g bath.

Here, 5, the foaming index, is equal to the average time of travel of gas through the foamed layer. It is to be noted that the above equations are valid only when £ is independent of the foam height. It has also been shown that, for an ideal system for which £ is constant, the foaming index 8 is equal to the average foam life. In the Ca0 -Si02 -Fe0 system at 1400°C, 8 decreases with increasing basicity to a value of 1.22 but beyond this, 8 increases. P2O 5 increases 8 slightly, whereas CaS marginally decreases it, and Cap 2 and increasing tempera­ ture also decrease 8 . In some recent work, Ito et al.'*^ obtained an empirical equation to predict the foaming index for Ca0 -Si02 -Fe0 slag using dimensional analysis: 8 = 5.7 • 102- (p) • (y • p)"

(8.18)

where p = viscosity of slag. Pa • sec, y = surface tension, N • m and p = density, kg • m"^. The 8 values for bath smelting slags were estimated, and the foam height in actual practice was calculated. They also studied foaming in bath smelting processes in a 200 t BOF vessel and found that the foam height increases with increasing rate of production. The amount of gas generated from the reactor decreases with an increasing prereduction degree, which causes a decrease in the superficial gas velocity in the vessel and the amount of slag foaming. Since the addition of coke or coal suppresses slag foaming by breaking the bubbles physically, the production rates limited by foaming will be more in actual bath smelting operations compared with the calculated values. Pressurized operation may be effective in increasing the production rate, because the reduced off-gas volume decreases the foam height. Moreover, the foam height is proportional to the production rate, which can be controlled by the rate of reduction of iron oxide in slag. Since the reduction rate of iron oxide in slag with solid carbon is proportional to ap^o^^ and reduction with carbon in molten iron is also expected to be proportional to ap^Q. the foam height, h, can be expressed by the following equation: h = C • 8 • ap,

(8.19)

where C = sum of constants including reaction surface, furnace diameter, etc., which are not a function of FeO concentration.

lY. PRACTICAL CONSIDERATIONS The electrical power required for melting in coal-plus-electricity-based processes is gen­ erated internally by waste heat recovery from the off-gases of the reduction step; otherwise, it must be supplied from an external source. The main limitations are that waste heat/power

176

Beyond the Blast Furnace

requirements are dependent on the quality of coal, and the economics of hot metaJ production in this system depend primarily on the cost of coal and electricity. On the other hand, in alJcoal processes, direct gasification of coal takes place, which generates heal for coal drying, heating, prereduction, and melting. Table 8.3 shows the energy consumption figures expected in various SR processes,^' which indicate that almost all of them (except COIN) use energy ranging from 15.5 to 19.0 GJ/thm, which is slightly higher than that of a blast furnace (between 14 to 15 GJ/thm). It should be noted that the fixed carbon content of coal has a major influence on the energy requirements; coal with low levels of ash, sulfur, and VM can be expected to give higher fuel efficiency. It should also restrict the input of energy-consuming components into the system, such as fluxes, desulfurizers, etc. In addition, credits derived from the off-gas for power generation or for heating purposes in other parts of the steel works also play a major role in accounting for variations in the energy requirements. In contrast to other processes, which rely primarily on coal, PLASMAS MELT is highly power intensive because of the plasma heater that generates both the reducing gas and the heat for smelting; it may be suitable only in locations where hydroelectric power is relatively inexpensive. One of the main requirements for the development of SR processes is the flexibility in processing a wide range of iron ores in terms of chemical characteristics and granulometry, with little or no requirements with respect to their treatment or preparation. The expected iron ore requirements for various processes are given in Table 8.4, which shows that while ELRED and PLAS MAS MELT claim to be well suited for relatively inexpensive, finely sized ore, the other three processes — COIN, COREX, and COMBISMELT — depend almost exclusively on the use of pellets or size-graded lump ore. Based on the description given, the salient features of the various new processes can be summarized as follows: •

• •

•

All the SR processes for hot metal production comprise two basic functions of the blast furnace process, namely prereduction in the stack region and smelting in the hearth region; these are carried out in separate units in almost all the new processes. Some of the processes are based on the use of iron ore fines while others are based on lump ore, pellets, sized sinter, or mixtures thereof. Some of the processes are based on noncoking coal and purchased power, others are based on noncoking coal and power generated within the process module itself; the remainder use noncoking coal alone. An important feature of many of these processes is the concept of recovering waste energy for generation of electric power. In fact, in many SR process flow sheets, facilities for the recovery of the waste gas energy are far more elaborate than the SR reactor itself.

For SR processes to compete successfully in small installations with (mini) blast furnaces and DR processes, the hot metal produced should be such that no costly downstream treat­ ments like desulfurization, dephosphorization, etc., become necessary before conversion to steel. Further, for a successful new process system, the investment cost must be well below that of the blast furnace system (including sinter and coke plants).

Important Features of SR Processes

177

REFERENCES 1. Smith, R. B. and Corbett, M. J., Iro n m a k in g S teelm a king, 14(2), 1987. 2. Yusheng, Z. and Ting, D., P roc. S h en y a n g Int. S y m p . o n S m e ltin g R e d u ctio n ,

S h e n y a n g , P.R . o f C h in a ,

1986. 3.

Davies, M. W., Hazeldean, G. S. F., and Smith, P. N.,

Proc. R ich a rd so n C o n f. o n P h y sica l C h e m istr y ,

L o n d o n , J u ly, 1973.

4. Fun, F., M et. T rans., 1(9), 1970. 5. Sasaki, Y. and Soma, T., M et. T rans. B., 8 B ( 1), 1977. 6. B ase, P., O n the C arbon R e d u c tio n o f Iron O x id e R ich

S la g s , P h .D . th e s is , ITT, K haragpu r, India, Janu ary,

1991.

Lloyd, G. W., Young, D. R., and Baker, L. A ., Iro n m a k in g S teelm a king, 2 (2 ), 1975. Wagner, C., The P h y s ic a l C h e m istry o f S teelm aking, E llio t, J. F., E d., John W ile y & S o n s , N e w Y ork , 1 958. 9. Tarby, S. K. and Philbrook, W. 0 „ T rans. M et. Soc. o f A IM E , 239, 1967. 10. Rawling, J. R. and Elliot, J. F., Trans. M et. Soc. o f A IM E , 233, 1965. 11. Grieveson, P. and Turkdogan, E. T., Trans. M e t. Soc. A IM E , 2 3 0 , 1 9 6 4 . 12. Graenzdoerffer, G., Kim, W. M., and Fine, H. A ., Proc. P r o c e ss T e c h n o lo g y C o n f., 1 9 8 8 . 13. Steffen, R., Steel Res., 60(3^ ), 1989. 14. Fruehan, R. J., Proc. Int. C o n f. on N e w S m e ltin g R e d u c tio n and N ea r N et S h a p e C a stin g T e c h n o lo g ie s for 7. 8.

S te e l, R IS T , P o h a n g , K orea, 1990.

Cooper, C. F. and Kitcher, J. A ., TIS I, 193, S ep tem b er, 1 9 5 9 . 16. Swisher, J. H. and McCabe, C. L., T rans. T M S -A JM E , 2 3 0 , D e ce m b er , 1 9 64. 17. Hara, S., Ikuta, M., Kitamura, M., and Ogino, M., T e tsu -to -H a g a n e , 69(9), 1983. 18. Ito, K. and Fruehan, R. J., Part I, M et. T rans. B, 2 0 B , A u g u st, 1 9 89. 19. Ito, K. and Fruehan, R. J., Part II, M et. T rans. B, 2 0 B , A u g u st, 1 9 89. 20. Sugata, M., Sugiyama, T., and Rondo, S., T e tsu -to -H a g a n e , 58, 1972. 15.

21.

Jana, S. and T e o h ., L .,

SEAJSI Q .,

14(4), 1985.

Power -3.56 generation credit (300 kWh)

15.48

Coke Oil

Total

17.32

—

Electricity generated (680 kWh)

—

100 —

136 —

Auxiliaries. kWh Byproduct gas, Nm-^ 50 kg 0-140 kg

130

700

194

Oxygen. Nm^

1100

680 kWh consumed

370

200-0

17.91-18.41

1.42 0-5.82

11.30

5.19

PLASMASMELT

17.32

617

620

Electricity. kWh

19.04

INRED

680

ELRED

Coal.* kg

Process

TABLE 8 3

— —

525

—

350

—

440

COIN

10.67

^ .3 9

2.80

12.26

— —

2000

—

700

1000

15.48

-17.57

5.02

28.03

COREX

Energy Consumption per Ton of Hot Metal in Different SR Processes (in GJ)

(CFB)

Energy recovery

(SAF)

Smelting

(SL/RN)

Reduction

18.83

-16.32

9.20

25.94

COMBISMELT

Si

2

r

S03

;s

03

QO

Remarks

*Coal analysis

Fixed carbon. % Ash, % Volatile matter. %

Sufficient electrical energy generated from off-gas d.c. arc furnace and supply 300 kWhh of credit

81 4 32-37

Anthracite, bituminous coal or lignite

Sufficient electric power generated from off-gas to operate electric arc furnace

5 6 51 14 3&35

Cheap gas. coal. anthracite 88.0 9.0 3.0

Coke

Plasma arc heater used to superheat a portion of process off-gas which is injected with coal dust and iron ore fines at slag metal interface. Small amount of coke used in the shaft

75.9 8.6 15.5

Coal 75 9.2 14.9

A surplus of CO-rich gas (byproduct) as generated and for gas exponlsale. The necessary quantities of oxygen can be generated by internal energy utilization

Lignite coal (0.3% S) 2-3% moisture 15 (Size: - I mm)

Lean coal (1.0%)

CFB plant allows recovery of energy in the off-gas and from coal discharged at the end of kiln end: this amounts to 16.3 GJh

44 27 29

Lignite, subbituminous coal, bituminous coal, anthracite, coke breeze

180

Beyond the Blast Furnace TABLE 8»4 Iron Ore Requirements and Composition of Hot Metal Expected in Different SR Processes Hof metal composition,

Process

Capacity,^ tpa

ELRED

120,000400,000

INRED

200,000400,000

PLASMASMELT COIN

250,000450,000 —

COREX

300,000

COMBISMELT

200,000

Iron ore Fine ore (grain size 0.1 mm, av., Fe^ : 65%) Fine ore, calcined pyrite, dust containing Zn, and Pb from gas cleaner Fine ore Pellet and lump ore, Fe^j : 65%

Pellets, Fe^j : 65% and size: l/2"-7/8" Pellets, Fe^ : 65%, Lump ore

s

C

Si

Mn

P

3.0-4.0

0.05

__

__

0.5

3.7 1.2 (with pyrite ore)

—

—

—

2.4 0.3 (with magnetite ore) 4.0 1.0

—

—

—

—

—

—

1.2 — (with lean coal containing 1.0% S) 2 .7 ^ .4 — (with lignite containing 0.3% S) 3.7 1-4

2.0 0.5 (high carbon iron) O.l 0 .0 1 (low carbon iron )

No commercial plant has actually been built except in the case of COREX.

-

0.1-0.2

0.02-0.20

0.1

0.05-0.20

—

0.1

0.3

—

0.03

0.02

Chapter 9

ALTERNATIVE AVENUES OF IRONMAKING L INMETCO PROCESS IntemationaJ Metals Reclamation Company (INMETCO), a subsidiary of INCO in Canada, has introduced a new ironmaking process based on a rotary hearth furnace (RHF).' The process utilizes iron oxide fines and coaJ fines for producing hot briquetted iron, hot metal, cold pig iron, or crude steel, depending upon the specific requirements. The RHF in combi­ nation with a submerged arc furnace (SAF) offers a competitive method for the production of hot metal. The test results obtained from a 25,000 tpa commercial plant at Ell wood City, Pennsylvania, have strengthened the claim of the following advantages in the process: • • • • • •

Both iron ore fines and coal fines can be used, with very few quality restrictions. The process consists of completely independent systems for agglomeration, reduction, and melting, thus enabling easier operation and improved process optimization. Excellent process flexibility can be achieved with respect to independent control of the atmosphere, temperature, and residence time in the RHF. Very short residence times in the RHF ensure rapid adjustment of process parameters and minimize production loss from process upsets. Continuous feeding of hot reduced iron to the SAF involves lower capital costs as well as reduced electricity consumption, compared to those in a conventional EAF. The module size of the plant can vary between 100,000 and 400,000 tpa for economic operation.

A process flow sheet for the production of hot metal using this process is shown in Figure 9.1. Iron oxide fines and coal fines are mixed together, along with a binder, before being pelletized. The green pellets containing the oxide and the reductant are placed on a rotary hearth (Figure 9.2) in an even layer, one to three pellets deep. As the hearth rotates around the circular furnace, the pellets are heated to 1250 to 1350°C and are reduced to metallic iron. The intimate contact between carbon and iron oxide in conjunction with the high reduction temperature results in very fast reduction rates. The residence time in the hearth is typically about 15 to 25 min, during which 90 to 95% metalization is achieved. The RHF is heated by multiple burners located at the inner and outer periphery of the furnace, and a wide variety of fuels, such as pulverized coal, natural gas, oil, coke, oven gas, or coal-oil mixtures, can be used in the burners. The flue gas from the RHF is used to preheat the combustion air to about 600°C, and after leaving the preheater, the flue gas is quenched, scrubbed, and released to the atmosphere through a stack. The hot reduced pellets are discharged at 1000 to 1200°C from the RHF to refractory lined transfer bins, and then continuously fed to a stationary SAF operating with an open slag bath (Figure 9.3). Carbon and sulfur contents of the hot metal are controlled by adding requisite amounts of lime and coke, and carbon and sulfur levels of 0 .1 to 3.5% and 0.05 to 0.15%, respectively, can be obtained depending upon the raw materials. Hot metal is tapped in batches, at intervals dictated by the optimum cycle for downstream steelmaking and casting. A wide range of reductants, like noncoking coal, char, coke breeze, charcoal, etc., can be used in this process. The total thermal requirement of an RHF is approximately 16 GJ/thm, and typical power consumption for melting is in the range of 300 to 350 kWh/thm at a melter

181

182

Beyond the Blast Furnace Iron oxide Reductant I

Exhaust gas

Combustion nuu»i air

Binder

Fan

Scrubber Stack

'[ 3 — 1

1

Heit exchanger

1

F,n

Mixer

t]

i

Pelletizer

I Rotary hearth furnace

l

FIGURE 9.1.

g

^

^

INMETCO process flow sheet.

Burner fuel

FIGURE 9.2.

Rotary hearth furnace used in the INMETCO process.

FIGURE 9.3.

Open hearth operation o f electric metter.

^

Burner fuel

183

Alternative Avenues of Ironmaking Coal

3.78

Electricity coke electrodes

0.30

0.19

O.Oé

gas ^ E xExhaust hi

FIGURE 9.4.

Energy balance of INMETCO process (gross GJ/thm).

energy efficiency of 80%. The overall yield of iron in the process is about 95% (losses are 2% and 3% in the RHF and me Iter, respectively). A typical process energy balance shown in Figure 9.4 illustrates the energy requirement at various stages in the system. This process is claimed to offer a relatively simple and economic approach to hot metal production. Production cost is reduced considerably because of the use of ore and coal fines, as well as by achieving a high iron yield (95%). Fines and dust generated during processing can also be recovered and recycled. Combustion of volatiles from the reductant is completed inside the RFIF, and the sensible heat in the exhaust gases is used to maximize the energy efficiency. Capital costs are also low as a consequence of the direct use of coal and ore fines and of the complete avoidance of induration of pellets, generation of steam, electricity, and oxygen as well as the need for a complex gas recycling plant. Typical operating figures for a 300,000 tpa plant are as follows: Items Iron ore fines CoaJ fines Electricity Labor Electrode paste Carburizer Lime Binder

Requirement/thm 1.55 t 0 .7 0 I 450 kW b

0.8 manhours 8 kg 4 0 kg 5 0 kg 15 kg

Potential applications of this process include the following: • • •

Supplementary hot metal for existing steel mills. Small capacity, integrated mini-mills based on noncoking coals and ore fines. Recycling of waste or byproduct oxide fines.

The process appears to be quite attractive from the point of view of energy consumption and its flexibility with regard to raw materials requirement and product chemistry.

184

Beyond the Blast Furnace

IL MIDREX FASTMET PROCESS Based on extensive market studies on the application of direct reduction processes, MIDREX concluded that in countries lacking adequate availability of natural gas, a need existed for a coaJ-based ironmaking technology, which should have the following characteristics:^ Low capital cost for plants producing 100,000 to 500,000 tpa of DRl Use of iron ore fines and pulverized coal Low electricity consumption Operating flexibility Low environmental contamination Continuing efforts to develop such a process ultimately culminated in the MIDREX FASTMET Process (similar to INMETCO), converting iron oxide fines and pulverized coal into DRl suitable for the production of pig iron or steel, depending upon the specific application. The focal point of the process is an RHF similar to that used conventionally in the iron and steel industry for heating tubes and billets. Use of an RHF in direct reduction was facilitated for the first time by MID REX Corporation for the “Heat Fast” process in the 1950s.^ This process consisted of mixing and pelletizing iron ore fines and coal fines, drying the pellets on a grate, prereduction on an RHF, and finally, cooling of the pellets in a shaft furnace. A 2 tph pilot plant was successfully run at Cooley, Minnesota, from 1964 to 1966. In 1974, INCO used an RHF to reduce stainless steel mill wastes, and pilot plant tests at Port Colboume, Ontario, Canada, later led to the development of the INMETCO process. The advantages claimed for the MIDREX owned FASTMET process include the follow­ ing: The process does not demand any significant technological breakthroughs; nearly all the equipment can be manufactured, even in countries having a nominal industrial base. Iron ore fines and coal fines, which are otherwise seldom used directly in the iron and steel industry, are used as feed materials. In-plant dusts and fines can be recycled. Completely independent units for agglomeration, reduction, and melting make the process easier to operate and to optimize. There is a high degree of flexibility with respect to independent control of temperature, atmosphere, and residence time, which makes the process suitable for materials with specific characteristics. Rapid adjustment of process parameters and minimal production loss from process upsets are achieved because of the low residence time in RHF. Continuous feeding of hot reduced iron into the me Iter reduces the operating cost. The process can be used for the production of hot briquetted iron, cold pig iron, hot metal, or steel. Process economy can be achieved at unit module sizes varying between 150,000 and 450,000 tpa. Compared to other coal-based ironmaking processes, it has a low environmental impact. The export gas generated is rather small and need not be utilized to justify the process economics (this is quite unlike all SR processes, where utilization of excess export gas is a deciding factor). A PROCESS DESCRIPTION In the FASTMET process (Figure 9.5), iron oxide fines, pulverized coal, and binder are mixed together and pelletized.^ The resulting green pellets are fed either to a dryer or directly to the rotary hearth furnace, where the pellets are placed on the rotating hearth in an even layer.

Alternative Avenues of Ironmaking

185

Iron oxide

FIGURE 9.5.

FASTMET process flow sheet for making steel.

FIGURE 9.6.

Cross section and plan view of rotary hearth furnace used in the FASTMET process.

one to three pellets deep (Figure 9.6). As a result of the rotation of the hearth, the pellets are heated to 1250 to 1350°C and reduced to metallic iron. Reduction is accomplished by the intimate contact between the in situ carbon contained inside the pellets and the iron oxide at a high temperature. The residence time on the hearth varies between 8 to 30 min, depending upon the raw material characteristics being used, the number of pellet layers, and other associated factors. During this time, 90 to 95% of the iron oxide is reduced to DRI, which is discharged continuously from the furnace into refractory lined hot transfer cans at 900 to 1000°C. Afterward, lime is mixed with DRI and the cans are sealed and transported to the melt shops for hot charging into the melter. The ability to hot charge DRI from the RHF into a melter plays a key role in the overall economics. Hot charging of DRI at 850°C enables an electrical energy savings of about 230 kWh/t of DRI plus an additional savings in electrode consumption, refractory consumption, and tap-to-tap time. The hot product from the RHF can be melted by any of the following methods:

186 • • • •

Beyond the Blast Furnace Electric Arc Furnace (EAF) Submerged Arc Furnace (SAF) Energy Optimizing Furnace (EOF) Smelting Reduction Vessel (SRV)

Of these options, the RFiF/EAF combination seems to be the most attractive.

B. RAW MATERIALS AND ENERGY A wide range of iron-bearing materials such as hematite, magnetite, ilmenite, iron sand, waste oxide, etc., have been tested in the laboratory and found to be generally suitable for processing in the FASTMET process.^ While using earthy hematites or fine concentrates, it is necessary to include a dryer in the flowsheet, and a grinding circuit is required if the ore is coarser. Obviously, the gangue content in the ores should be as low as possible, to minimize melting costs. The FASTMET process can accept oxide fines (80% should be o f - 3 2 5 mesh size) as pellet feed. Because of in-plant recycling of fines and dust, very little loss of the undersized materials occurs, resulting in higher yields. This is an advantage over other direct reduction processes (including those using fluidized beds), which require coarser ores and therefore do not have as much flexibility in recycling in-plant fines. Any coal, coke, or char containing less than 20% ash and at least 50% fixed carbon can be used as reductant if a dryer is included in the flow sheet, whereas coals with fixed carbon above 70% may not require a dryer. Sulfur content of the reductant should not be above 1%, and the preferred size range of the reductant is 70 to 80% below 200 mesh. The heat supply to the RFIF is ensured by burning pulverized coal, natural gas, propane, or fuel oil fed through direct-fired burners. Pulverized coal firing increases the capital cost of the plant but at the same time provides a better radiant flame than that obtained with natural gas and is more cost effective if the price of other fuels is significantly higher. Pulverized coal used as the burner fuel should be sized at -200 mesh (70 to 80%) with volatile matter 30% (minimum) and ash 20% (maximum). C. PROCESS APPLICATIONS Figure 9.7 illustrates some of the possible applications of the FASTMET process.^ These include the following; • ® • • •

Supply of iron units containing low residual contents to supplement scrap in existing EAF melt shops for the production of high quality steels Supply of low residual iron units to new flat product mini-mills featuring thin slab casting facilities Supply of iron units as a replacement of costly imported scrap in scrap-lean countries Source of iron for the supplementary production of hot metal at existing integrated (BFLD) steel works Replacement of old coke oven/sinter plant/blast furnace capacities in existing integrated steel mills

Along with all the above-mentioned applications, the FASTMET process may find appli­ cations in the following nonironmaking areas:• • • • •

Reduction of chrome, nickel, manganese, or titanium ores Reduction and treatment of stainless or carbon steel mill wastes Calcination of limestone, phosphate rock, gypsum, lignite, etc. Pyrolysis of industrial wastes, tar sands, oil shell, etc.

Alternative Avenues of Ironmaking

187

HBI Granulated ’ iron or pig iron -S te e l

- Steel

-S te e l

• Steel

FIGURE 9.7.

FASTMET process applications.

Plant capacity. M t / y

FIGURE 9.8.

FASTMET capital cost as a function of rated capacity (with dryer).

D. ECONOMICS In most of the ironmaking processes, 80 to 90% of the production cost is tied to the cost of raw materials, energy, and capita] investment.^ Thus, any significant breakthrough in ironmaking economics is likely to be made in these three areas. Because of the use of iron ore fines and efficient recycling of fines and dust generated during processing, the iron ore cost is reduced to a minimum in the FASTMET process. Energy costs can be minimized by using available coal fines. Use of an ore-coal mixture and direct firing of fuel using air preheated with sensible heat from the exhaust gases also help to reduce energy consumption. Hot charging of DRJ further reduces required electrical energy as well as refractory and electrode consumption. The estimated capital cost for a turnkey 400,000 tpy FASTMET plant is 140 dollars (US)/ annual ton of DRJ (as of 1991), and its variation with rated capacity is shown in Figure 9.8. The facilities include raw material bins, reductant pulverizing system, mixer, pelletizing disk, dryer, rotary hearth furnace, flue gas treatment system, process and machinery cooling water systems, hot transfer system, instrumentation, buildings, and cranes. Typical FASTMET process operating parameters for estimating production costs of D RI are summarized in Table 9.1. Table 9.2 contains the typical operating parameters for producing steel using hot D RI from the FASTMET process as 50% of the charge in a 500,000 tpa EAF-ladle furnace (LF) melt shop. Likewise, specific consumption figures for the production of hot metal using

188

Beyond the Blast Furnace TABLE 9A

Typical O perating Parameters for Estimating Production Costs of DRI Using the FASTMET Process Item Iron ore concentrate Reductani coal Burner coal Binder Lime for flue gas desulfurization Fuel for pulverizing coal and start-up Electricity Water Nitrogen Maintenance and supplies Labor

Units/t of DRI 1.25 t 0.27 1 (9.92 GJ) 0.12 t (4.06 GJ)^ 22 kg 6 kg 0.33 GJ 60 kWh 1.1 m^ 2.5 Nm-^ 6.00 dollars (US) 0.2 manhours

If natural gas firing is used instead of coal firing, then the burner fuel would be 4.48 GJ/i of DRI.

TABLE 9.2 Typical Operating Parameters for Producing Steel Using Hot DRI from the FASTMET Process as 50% of the Charge

in a 500,000 tpa EAF-LF Melt Shop Item DRI (78% Fe, 11.3% FeO. 3% C and 7.6% gangue) Scrap Electricity for EAF and LF (DRI at 850°C. scrap at 25°C) Oxygen Fuel Carbon Lime for EAF Flux for LF (ladle furnace) (CaO, AÌ20^, CaF 2 ) Alloys for LF (Fe-Mn, Si-Mn, Fe-Si) Electrodes for EAF and LF Refractory for EAF and LF Maintenance and supplies for EAF and LF Labor for EAF and LF

Units/t of liquid steeF 0.59 t 0.59 1 525 kWh 12 Nm^ 0.21 GJ 1.2 kg 75 kg 40 kg 7 kg 3.5 kg 6.50 dollars (US) 8.00 dollars (US) 0.7 manhours

For estimation purposes, the tap temperature and carbon content of steel produced in this facility are assumed to be 1600°C and 0.1%, respectively.

FASTMET hoi DRJ in a 250,000 tpa SAF including hot metal desulfurization facility are presented in Table 9.3. The MIDREX FASTMET process promises a simple and economical approach to hot DRI production to be used for making pig iron or steel. It combines proven unit operations and equipment into a reliable ironmaking system and offers greater flexibility. These, however, are still claims, as the process has yet to be commercialized. In particular, in countries where the availability of coal with less than 20% ash is limited, the process may not merit serious consideration.

Alternative Avenues of Ironmaking

189

TABLE 9 J Specific Consumption Figures for the Production of Hot Metal Using FASTMET Hot DEI in a 250,000 tpa SAF Including Hot Metal Desulfurization Facilities Item

Units/t of hot m etaF

D RJ

Coke Electricity Electrode paste Lime Flux for hot metaJ desulfurization (caJcium carbide and diamide lime) Refractory for SAF and hot metaJ desulfurizer Maintenance and supplies Labor

1.22 t 40 kg 370 kWh 8 kg 122 kg 8 kg 5.00 dollars (US) 5.00 dollars (US) 1.0 manhours

^ The hot metal produced in this facility would have a temperature of 1400°C and chemistry consisting of 95.5% Fe, 3.5% C, 0.6% Si, 0.02% S, and 0.06% P.

IIL IRON CARBIDE MANUFACTURE In this novel method, iron ore fines are converted to iron carbide in a fluidized bed using carbonaceous gases and hydrogen."^ The process produces iron carbide powder which is hard, nonfriable, resistant to oxidation (up to 300°C), and has the potential to replace DRl, scrap, or molten iron to produce steel at a lower cost. The process was invented in the United States by Frank M. Stephens, Jr., who later formed a company named Iron Carbide Holdings, Ltd., which possesses the right to license this technology. In this process, iron carbide containing 7% C and 93% Fe is produced in a fluidized bed by treating preheated iron ore fines with an appropriate mixture of hydrogen, carbon monox­ ide, and hydrocarbons at a temperature of 600°C and pressure of 1.8 atm. The reactions leading to the formation of iron carbide starting with hematite and magnetite iron ores are as follows:^ 3Fe203 + 11H2 + 2CO = 2Fe3C + 11H2O Process requirement:

(9.1)

H2 =0.515 Nm-^/kg Fe203 CO = 0.094 Nm^/kg Fe203 Fe3C = 0.7495 kg/kg Fe203

Fe304 + 5 H2 + CO = Fe3C + 5 H 2O Process requirement:

(9.2)

H2 = 0.484 Nm^/kg Fe304 CO = 0.097 Nm^/kg Fe304 Fe3C = 0.7754 kg/kg Fe304

The hydrogen and carbon monoxide requirements must be carefully matched to arrive at iron carbide.

190

Beyond the Blast Furnace

Another reaction involving methane may also take place:^

3Fe,0, + 5H, + 2CH. = 2Fe,C + 9H,0

(9.3)

Thus, the only byproduct of iron carbide manufacture is water vapor, and there is, therefore, no pollution threat. Hydrogen is the principal reducing agent and carburization of the iron is effected by carbon generated from the carbon deposition reaction:

2CO = CO, + C

(9.4)

The high thermal efficiency of the process is inherently associated with the fact that the temperature of formation of iron carbide in the fluidized bed reactor is only 600°C, versus around 900 to 1100°C for DRJ production and about 1800 to 2000°C to produce hot metal in the blast furnace and smelting reduction processes. The lower operating temperatures in this case prevent the formation of ferrous oxide, thus eliminating sticking problems normally encountered in other fluidized bed reduction reactors. Approximately 1.35 t of fine grained hematite ore, with the minimum possible gangue content, and 12 GJ of natural gas are required to produce one ton of iron carbide, and 500 liters of distilled water is generated concurrently. Regarding the iron oxide feed preparation,^ the particles should be reasonably uniformly sized: typically the ore feed falls into two ranges, -1 mm, +0.1 mm, and -0.1 mm. In a recent study conducted at Hazen Research, Inc. with Indian blue dust containing over 66% iron, the product analyzed:^ = 97.5%, Fc304 = 0.2%, Fe = 0.0%, FeO = 0.2%, and gangue = 2.0% (assumed). The reaction takes place in a highly energy efficient fluidized bed reactor. Screened and classified iron ore is preheated (often in a rotary kiln) to around 500 to 600°C and introduced into the reactor, where the hot gas under pressure is forced upward through the bed of finely divided iron ore at sufficient velocity to fluidize it, thereby bringing the solid charge into intimate contact with the gas. This promotes efficient chemical reactions, leading to the formation of iron carbide and water vapor. The reactive elements in the gas stream are hydrogen and carbonaceous gases. The off-gas is then cleansed from its dust burden by a cyclone at the top of the reactor, before passing to a heat exchanger for cooling and finally into the scrubber to condense the water vapor and to remove the last traces of dust."^ Ultimately, the scrubbed gas is reconstituted with make-up gas, passed through the heat exchangers to boost the temperature, raised to the reactor working pressure, heated to a temperature of over 600°C, and reintroduced into the reactor (Figure 9.9). Some of the important process features are highlighted below:^ • • • • •

The process is not exothermic; rather it is slightly endothermic, and hence poses no risk of catastrophic temperature rise. All the feed materials and the product are far away from their softening/melting points, and the materials remain solid. Decripitation of neither the feed nor the product occurs in the reactor because of the low temperatures and the fine sizes involved. No volatile iron compound is formed in the carbiding reaction. No corrosive gases or solids are involved in the process.

Iron carbide can be used as a substitute for hot metal, scrap, and DRJ in EAF and LD steelmaking. The combined carbon in iron carbide forms a latent source of energy. While being oxidized in the fumace/converter, this carbon releases a significant quantity of heat into the steel bath, thereby reducing the heat otherwise required from other sources. It has also been shown theoretically that preheated (1200°C) iron carbide can constitute 100% of the charge

Alternative Avenues of Ironmaking

FIGURE 9.9.

191

Process flow sheet for the production o f iron carbide.

in oxygen steelmaking. If pure iron carbide is used and the carbon in the iron carbide is burned to 90% CO and 10% CO 2, the additional energy requirement to produce 1 ion of steel would be about 0.72 GJ, while with iron carbide preheated to 1 100°C, steelmaking would become autothermal.^ While using in an EAF with oxygen lancing, it may be used as the sole charge material without any preheating. A typical comparative cost analysis between the standard blast furnace ironmaking practice and various modes of use of iron carbide is illustrated in Table 9.4.^ Table 9.5 summarizes the superiority of iron carbide over many other methods of making iron.^-^ With its relatively fine consistency, the free-flowing iron carbide permits close control over its introduction into a LD converter. It may be charged at the bottom of the vessel, injected through tuyeres below the liquid level, or injected below the slag level. Similarly, it can be added easily into an EAF by means of gravity feed pipes or lances. Iron carbide can also be used as a ferro-alloy for carbon addition in the ladle. Another interesting possibility is the use of sintered iron carbide as cutting tool tips. Besides these technological features, the economic advantages of iron carbide include the following: Excellent substitute for the blast furnace route of ironmaking, eliminating coking, sintering, and sizing steps. Capital cost per annual ton of iron claimed to be one-third to one-fourth that of blast furnace-coke oven combination and only one-half to two-thirds that of current DRI or HBI plants. Attains optimum production efficiency at a production capacity of 0.33 mtpa only. Substantially lower production cost claimed, based on the availability of natural gas.

Total Cost reduction. US$/t steel Hot metal savings, %

100

30 25 3/GJ

130

DRl

100

150

Price, US$/t

FC3C Lime Oxygen Fuel gas

Hoi MetaJ Scrap

Feed

0.072 0.074

0.851 0.332

168

2.16 1.85

0.072 0.126

0.236

10.9

181 13

2.16 3.15

30.68

145.376

us$ 0.944

us$

1

Cost,

t 131.054 33.2

Cost,

Amount,

Standard practice

Amount,

Direct Reduced Iron (0.5% C 1.0% Si02)

0.111

0.509 0.072

0.667

Amount, t

38.3

21.6

2.16 2.85

66.1

80.85

us$

Cost,

152 16

0.661 0.072 0.114

0.525

Amount, t

1.191 0.072 0.126 1.2 GJ

0

t

Amount,

100

128 40

1 19.1 2.16 3.15 3.6

0

us$

Cost,

1100°C Iron

Carbide (6.6% C - % S 1O2)

480°C Iron

Carbide (6.6% C - 0% Si02)

159 9

50.9 2.16 2.775

102.718

us$

Cost,

C - 0% Si02)

Cold Iron Carbide (6.6%

TABLE 9A Steelmaking Cost Comparisons: Standard Practice/Direct Reduced Iron/Iron Carbide

na

s

S' Co

533

\o

Alternative Avenues of Ironmaking

193

TABLE 9«5 Comparison of Selected Ironmaking Processes

Process

mtpa

Capital cost, US$ per annual ton

Blast furnace and coke ovens MIDREX HyL FIOR SL/RN ELRED Fe,C Hot Fe,C Hot Fe3 C [NRED COREX

3.65 0.60 0.60 0.60 0.15 1.30 1.27 1.27 1.27 0.45 0.30

Capacity,

•

•

Fuel used

Iron oxide feed

Fossil fuel equivalent, GJ/l

41 1

coke

lump/sinter

3.67

115 115 166 165 305 106 115 154 290

natural gas natural gas gas coal coal natural gas natural gas coal gas coal coal

pellet pellet fines lump fines fines/conc. fines/conc. fines/conc. fines lump/pellet/

2.70 2.70 3.30 4.34 3.39 2.52 2.78 3.27 4.17

Cheaper source of metallics compared to hot metal, scrap, and DRl, leading to lower cost of steelmaking. Since iron carbide does not contain any tramp elements, which is also the case with DRl, quality steel can be produced. It contains practically no sulfur, which means that an EAF could be operated with acid practice. Gives a stable price range and, again like DRl, is not subjected to the frequent cost fluctuations witnessed in the scrap market.

The estimated production cost of iron carbide (62 dollars (US)/t) is much lower compared with the MID REX DRl cost (about 102 dollars (US)/t). The capital cost for a plant of 1,000 tpd capacity (0.33 mtpa) has been estimated to be 105 dollars (US)/annual ton, compared to 115 dollars (US)/annual ton for a MIDREX plant. The total investment cost would be about 34 million dollars (US) for a 0.33 mtpa iron carbide plant (all 1991-1992 figures). Over the past 15 years, more than 100 different iron ore samples from various sources have been tested at Hazen Research, Inc. in Golden, Colorado.^ The initial test work was performed in a 2 in diameter batch reactor capable of processing about 3 kg iron ore at a time. Later on, the process was scaled up to a 24 in diameter reactor, producing iron carbide continuously at a rate of about 22.5 kg/h. The product was tested successfully in a basic oxygen converter at MEFOS in Lulea, Sweden, to produce tonnage steel.^ In 1989, an idle vanadium plant in Wundowie, Western Australia, was converted to a pilot plant to produce tonnage iron carbide. The plant was run for about three months, and the materials produced were tested in various steel plants. The plant used a 1.83 m diameter reactor with a bed depth of 3.66 m. North Star Steel Co., in the United States, was the largest single consumer of the iron carbide product from this plant, and several tests were performed successfully using a 115 t EAF.^ '® Although no commercial production facilities are in operation now, commercialization is anticipated by the end of 1993 in one/two 1,000 tpd unit(s) located in the United States.^ It will be of great relevance to many countries like India, endowed with vast reserves of high grade iron ore fines and concentrates and natural gas. This process could also be economical in certain locations, at a nominal plant capacity of 60,000 to 90,000 tpa, but in a high labor cost country like the United States, the minimum economic size of the plant has been projected to be 350,000 tpa.

194

Beyond the Blast Furnace

Charge conveyor { or skip ) Charge distributor (Optional)

Gas outlet Double bell type charging system B la st furnace top

Stock line

Stack Shell cooling system

Bosh - Hot air ring Tuyere

Hearth Hearth bottom

BLAST FURNACE » Section

F I G U R E 9.10.

SectionaJ view of a typicaJ mini blast furnace.

lY. IRONMAKING IN MINI BLAST FURNACES Besides the classical route of ironmaking in blast furnaces, which had a 97% share of global iron production in 1990 (527 mi), many other alternative processes of ironmaking have arisen over the years. Although most of the processes are in their nascent state, a handful have been able to establish their industrial importance, e.g., shaft furnace DR processes (MIDREX, HyL), rotary kiln DR processes (SL/RN, CODIR, TDR), and very recently, the COREX smelting reduction process. None of these processes, except COREX, can produce liquid iron; a growing need over the years to produce liquid iron with inferior grade coal or even with charcoal has led to the development of alternative technologies for liquid iron production. The mini blast furnace (MBF) is the most proven technology, as revealed by the global iron production figures: while gas-based as well as coal-based DRJ production routes produced 2.7% of the total iron production in 1990-1991, the corresponding share held by MBF, operating practically only in Brazil, China, and India, was 3.4%.

A. SPECIAL FEATURES OF MBF An MBF,which can be viewed as a miniature version of the conventional large blast furnace (Figure 9.10)," also has a few additional characteristic features known for their simplicity and economy. These include: 1. 2.

3.

Working volume: The working volume of an MBF can vary between 150 and 370 m^, corresponding to production capacities of hot metal between 60,000 and 200,000 tpa. Top pressure: MBFs are usually operated at a top pressure of about 0.3 kg/cm^, which is just enough to ensure an efficient gas cleaning by a double venturi scrubber system. No pressure release turbine is necessary, and a normal double bell type top closing system gives satisfactory performance. Gas cleaning system: A wet, double venturi scrubber ensures a solid content less than 10 mg/Nm^ in the cleaned gas. The performance is nearly comparable to that of the much costlier electrostatic filter systems of gas cleaning in conventional blast furnace operation.

Alternative Avenues of Ironmaking 4.

5.

7.

8.

195

Blast compression: Unlike blast furnace operation with a huge turboblower and a standby, MBFs often use a set of fans, coupled in series with a provision for their easy replacement and repair. Blast preheating: Usually metallic blast preheaters (MBPs) are used for a steady, easily controllable temperature in the MBF blast, without requiring any hot valve or compen­ sation cycle as are used in the costlier stoves. Conventionally, the blast preheating temperature is raised to 1100°C or more in blast furnace operation. In an MBF, designed for operation even with 100% inferior grade raw materials, the maximum hot blast temperature required, in the absence of any blast additives, is about 850°C, and often 750°C can suffice. Lining and cooling: The MBF uses high duty medium alumina (62%) bricks, with shell spray cooling. Lining life is about 3 to 5 years, and downtime for relining is less than 30 days. Copper plates and an expensive water circuit are not required for furnace shell cooling. Instrumentation and control: Large blast furnaces require expensive and sophisticated instrumentation and control, whereas MBFs can do with only two 6-point temperature recorders, a few pressure indicators, and a blast volume and flow recorder. Raw materials: The MBF can easily accept 100% natural lump ore and coke of a poorer quality than large blast furnaces.

K CHARCOAL VERSUS COKE MBF Operation of MBFs with charcoaJ as well as coke has been extensively tried and tested in Brazil, and in spite of many similarities, a few differences have been reported, which arise primarily because of the character of charcoal compared to coke.^^ These are summarized below: • • • • • •

Temperature in the thermal reserve zone (800°C)is about 150°C lower than that in the coke MBF (950°C), due to the higher reactivity of charcoal. Residence time of the ore in the thermal reserve zone is about half that of the coke MBF. Volume occupied by ore in the charcoal MBF ( 15%) is about half that of the coke based counterpart (30%). Maximum blast temperature attained in charcoal MBF is about 850°C. The charcoaJ MBF has a much lower slag volume (120 kg/thm) with lower binary basicity (CaO/Si02 = about 0.8) and sulfur content in hot metal below 0.020%. Top gas from a charcoaJ MBF has a much higher heating value (4.0 to 4.2 MJ/Nm^) than that obtained while operating the furnace with coke (less than 3.1 MJ/Nm^).

Over the years, a steady rise in production, along with a gradual depletion of natural rain forests, have resulted in a shortage of charcoaJ and an appreciable increase in its price. This gave rise to attempts to operate charcoaJ MBFs with coke, and extensive trials started in early 1989 have now enabled the complete replacement of charcoal by coke. A comparative study conducted at the 250 m^ MBF No. 2 of Mannesmann S/A at Bello Horizonte, Brazil, using 100% coke and charcoal with the same ore, revealed the following: A coke rate of 570 kg/thm (carbon rate 498 kg/thm) was achieved, which was quite satisfactory, taking into account the relatively low blast temperature. An improved productivity of about 1.62 t/mVd could be achieved with associated decrease in coke rate (6%), while using the enriched top gas in the blast preheater, thereby boosting the blast temperature to 770°C. As expected, the pressure drop inside the furnace increased from 0.3 to 0.6 kg/cm^, whereas the top pressure remained the same (0.33 kg/cm^), along with an increased blast pressure (0.95 kg/cm^). Operation with coke was smoother and not hamnered by hangings and slippings.

196

Beyond the Blast Furnace

C» WORLD SCENARIO Small blast furnaces having volumes ranging from less than 100 m^ to a few hundred cubic meters are operating in some parts of the world, the majority of which are found in Brazil and China and a limited number of which in Germany, Bulgaria, Vietnam, Russia, and IndiaJ There are more than 100 charcoal blast furnaces of less than 800 m^ volume, accounting for about 30% (approx. 7 mtpa) of the total iron production of Brazil, while more than 50 MBFs with volumes of less than 300 m^ are in operation in China. In India, Kalinga Iron Works is successfully operating three low shaft furnaces with volumes of less than 100 m^, and an MBF of 175 m^ capacity was commissioned in Goa in 1992.

1. Brazilian charcoal blast furnace The Brazilian design incorporates a few major changes in the cooling arrangement, blast generation, and preheating equipment and in the operating practices and parameters.'^ Some of the features of this design are use of a thin shell, simple external cooling, use of low top pressure, and use of a recuperator instead of a single turboblower. The high reactivity of charcoal gives similar reaction kinetics at a lower temperature than that of metallurgical coke used in a blast furnace, which leads to the requirement of a highly reducible ore. Because of the low bulk density of charcoal (200 to 280 kg/m^), the ore occupies only about 15% of the burden volume, and excellent burden permeability, almost independent of the size and the decripitation properties of the ore, is achieved in this way. Some important dimensions of a charcoal blast furnace of 180 tpd capacity are as follows: Hearth diameter Working height Working volume Number o f tuyeres Tuyere diameter Wind volume Blast pressure Top pressure Blast temperature Top gas temperature

2.75 m 14.45 m I 10 m^

6 95 mm 10,500 Nm^/h (max.) 8.00 mm WG 1.80 mm WG 780°C 90 to 120°C

The normal requirement of raw materials per ton of molten iron is as follows: Raw m aterial Charcoal Lump ore Flux Silica Manganese ore

Requirement 650 kg 1600 kg 100 kg 40 kg 30 kg

2. Chinese mini blast furnace Presently about a fifth of China’s total iron production (55 mtpa) is accomplished by this m ean s.T h ese furnaces closely resemble their larger counterparts, with effective volumes and productivity ranges on the order of 100 to 300 m^ and 1.3 to 2.0 t/mVd, respectively. Extensive use of sinter (85 to 100%) is a usual practice, with a basicity (CaO/Si02) of 1.3 to 1.5. Unlike the Brazilian furnaces, the furnaces in China use metallurgical coke, and the coke rate varies between 550 and 630 kg/thm (with coke containing 14% ash). Presently, about 50 such MBFs are in operation in China, and extensive innovations have been introduced. Some of them are Injection of pulverized anthracite coal, as much as 60 kg/thm, to bring down the coke rate by about 40 to 50 kg/thm,

Alternative Avenues of Ironmaking • • •

197

Heat recovery from stove waste gas at 250 to 300°C for increasing the hot blast temperature by about 80°C, Incorporation of self preheating process stoves, enabling the generation of hot blast with a temperature of more than 1200°C, Dry cleaning of furnace gas, allowing better cleanliness than that of electrostatic precipitators.

Plant availability, coupled with the perfection achieved in technology, make MBFs a wellaccepted ironmaking route in China. While the productivity ranges between 1.70 and 2.80 t/ mVd, the average coke rate is 550 kg/thm. It is to be mentioned that, for the productivity calculation, the Chinese practice is to consider the volume between the tap hole and the stock line, whereas the standard practice elsewhere is to take the volume between the tuyere line and the stock line. Thus, on a like-to-like basis, the Chinese MBF productivity would be higher by about 10 to 12%. The Chinese experience shows that, compared to an MBF operating with only lumpy ore, an MBF on 80% sinter and 20% lump ore would gain in productivity by about 25%.’^ Some typical characteristics of raw materials used in Chinese MBFs are shown below: Iron ore

Sinter

Chemical analysis, % Fe FeO CaO SiO^ A IP 3 MgO S P Tumbler Index, % Basicity

5 3-54

40-45 —

1 1 -1 2

8

13-14

10.5-12.0 8.5-9.0

2

—

1

0.03-0.04

2.5 0.03-0.04

0 .0 2

—

81-82 1 .2 -1 .4

A typical size range of the iron oxide feed is as follows: Size

-7 0 -6 0 ^0 -25 -1 0 -5

and and and and and mm

+70 +60 +40 +25 +10 +5

Percent mm mm mm mm mm mm

2 2 8

12 49

20 7

The coke characteristics are: Ash V.M. Sulfur Moisture MIO index M40 index Size

13.5 - 14.0 1.1 - 1.4 0.25 - 0.75 7.5 - 8.0 7 75 2 5-60 mm

Some typical operational details (as of April 1992) of the SHUTA ZHUANG Iron and Steel Works, in the People’s Republic of China, typify the Chinese MBF practice:

Beyond the Blast Furnace

198 Facilities Available Blast furnaces

1 X 150

m ^,

1 X 300 Sinter plant

2

X

24

Production Details

150 m-i MBF Hot blast temperature, °C Production Productivity, t/mVd Coke rate, kg/thm Hot blast volume (excluding loss 15%), NmVmin Thermal value o f gas, MJ/Nm^ Slag rate, kg/thm CO2ACO + CO2 ) ratio, %

900-1000 330 tpd (0.110 mtpa) 2.2 560-570

300 m-i MBF

N.A.

900-1000 660-670 tpd (0.220 mtpa) 2.2 550 700

3.56 500 30

3.56 500 35

Analysis of Pig Iron,

C

4.0 1.4 (special foundry grade) 0.7 (normal) 0.12 0.07 0.03

Si Mn

P S

Analysis of Slag, 38-39 12.5 13.5 35.5 0.5 0.7

CaO MgO

A1,0, SiO^ FeO S

Analysis of Coke F.C.

85%

V.M.

1.8-

1.9% 13% 0.7% 2 5-70 mm

Ash S Size

Other details Sinter in burden, % Ore in burden, % No. of tap holes No. of slag notches No. of tuyeres Use of top gas, Nm^/h • Sinter Plant • Boiler • Balance Dust content in top gas, mg/Nm^

95 5 1 2 12 12000 5000 X 2 Stoves 20

Alternative Avenues of Ironmaking No. of stoves Manpower

199

600 for 2 MBFs

Besides Brazil and China, other industrially developed countries, such as Germany, Russia, and Japan, also offer MBF technology through various collaborators. German and Russian technologies offered by Demag and Gipromez, respectively, are similar to large blast furnaces. Even though not much activity is published in this area, Japan is reported to be running an MBF of 389 m^ useful volume to produce ferro-manganese.

3. Indian furnaces Three small shaft furnaces, one with a working volume of 37.19 m^ and the other two with volumes of 61.21 m^, each of German design, have been in operation at Ralinga Iron Works, Orissa, for the last 30 years. A few important design data for the 61.21 m^ furnaces are given below: Total volume Useful volume Hearth volume Hearth diameter Total height Useful height Stack height Number of tuyeres Tuyere diameter Number of tap holes

72.73 m-^ 61.21 m^ 6.43 m^ 3.00 m 14.60 m 9.75 m 5.80 m 8 ¡00 mm 1

The initial poor performance of these furnaces has improved remarkably through a few innovative operating practices, some of which are highlighted below: Use of cold bonded pellets in the burden Use of low temperature carbonized coke (trial) Formed coke in the burden (trial) Sponge iron in the burden (trial) Sinter in the burden Operation with low ash coke Lime dust injection External desulfurization of hot metal By incorporating some of these features in furnace operation, a productivity as high as 2.4 t/mVd has been achieved. Presently, rejected coke from integrated steel plants (10 to 25 mm) is blended with nut coke and hard coke and is used with lump ore (10 to 30 mm) and sinter. The slag basicity is kept at 0.7 to 0.8 with an MgO content of 8% in the slag, and the alumina level in the slag remains at about 22 to 26%. The first Tata-Korf MBF of 175 m^ size was commissioned by SESAGOA Ltd., Goa, in March, 1992. Table 9.6 highlights some of the salient features of the SESAGOA plant. Within 2 weeks after its commissioning, the plant attained the rated capacity. The capital cost per ton of installed capacity of such an MBF is estimated to be on the order of 160 to 180 dollars (US). Presently (1992), many new MBFs are coming up in India and entrepreneurs in the medium-scale private sector are also showing interest in this technology, particularly keeping in mind its suitability under the existing conditions. Two current projects in India are utilizing the Tata Korf technology, and another MBF (0.189 mtpa capacity) using German technology is in advanced stages of implementation, again in Western India.

im

Beyond the Blast Furnace TABLE

Some Features of the MBF at SESAGOA Basic, detailed engineering, supervision o f erection Project engineering coordination Capital cost Blowing system

Heating system

Top gas temperature Gas cleaning plant Pig casting machine Specific power consumption Gas holder Manpower MBF slag Productivity Coke rate Metal quality

Korf Techno logia Siderurgica Ltda, Brazil Tata-Korf Engineering Services Ltd., India 10.72 million dollars (US) (I dollar (US) = Rs 28) 6 electrically operated centrifugal fans in series, giving 20,000 Nm^/h air flow Design pressure = 1 . 3 kg/cm^ Actual blast pressure = 0.7 kg/cm^ Top gas pressure = 0.3 kg/cm^ Metallic recuperator (Glendon) fired by BF gas, 2 in parallel, designed for 900°C blast temperature (actual blast temperature for 600 kg/thm coke rate = 750°C) 250°C Wet system. Dust content in clean gas = 5 mg/Nm^ (max.) Carousel PCM to cast 5 kg pigs 165 kWh/thm 5000 Nm^ 175 275-300 kg/thm. used in cement plant 1.5 t/m^d 600 kg/thm Temperature: 1450°C. typical chemical composition: 2.0% Si. 0.6% Mn, 0.03% S. 0.06% P

REFERENCES 1. Lepinski, J. A., MIDREX Corporation, private communication. 2. D ire c t fro m M iD R E X , 14(1). 4th Quarter, 1990. 3. T ennies, W. L., Lepinski, J. A., and Kopfle, J. T., M P T , 2,1991.

4. Stefens, F. M., Jr., Steel Tim es In t., 13(3), 1989. 5. Evans, D. D., Stefens, F. M., Jr., and Williams, W. E. J., Iron Carbide Holdings Ltd., private communi­ cation.

6. Stefens, F. M. and Williams, W. E. J., Steel T echnol. In t., 1992. 7. 8. 9. 10. 11.

Report on the Conversion of Indian Blue Ore to Iron Carbide, Iron Carbide Holdings Ltd., January 29, 1992.

Stephens, F. A., Iron Carbide Holdings Ltd., private communication, April 16, 1990.

“Alternate routes of ironmaking,” T IE A C Rep.. Delhi, June 15. 1990. Garvey, R. A., M B M , March 1991. Lakshmanan, V. K., Jayaram, P., and Date, P. K., Paper presented at the Int. Conf. on Secondary Steel Sector: Problems and Prospects in the Nineties, New Delhi, India, 1991. 12. Pfeifer, H. C., Paper presented at the Int. Conf. o f Iron and Steel Technology in Developing Countries, Sao Paulo, Brazil, November 1986.

13. Patel, N. K., M et. N ew s, IIM , 13(5), 1991. 14. Panchapakesan, S., Report on the Visit to the People's Republic o f China, Tata Kothari Steels Ltd., India, 1992. 15. Das, A. K., Tata Steel, private communication. 16. Lakshmanan, V. K., G ro u p B u ll. T ata Steel, 3(2), 1992. 17. The H in d u , India, 24 April, 1992.

Chapter 10

USE OF DRI/HBI AND SR/MBF HOT METAL IN IRON^ AND STEELMAKING L USE OF SR/MBF HOT METAL AND DRI/HBI IN STEELMAKING The steel industry can be broadly classified into two categories, integrated steel plants and mini mills. The hoi metal produced in conventional blast furnaces is predominantly used for oxygen sleelmaking in integrated plants, and the route is well established. Similarly, hoi metal produced in MBFs or SR processes can be processed in oxygen steelmaking vessels in mini/ midi steel plants. A typical steelmaking process route based on SR (COREX) hot metaloxygen converter (LD)-secondary metallurgy (SM)-continuous casting (CC) is shown in Figure 10.1. The specific consumption figures of input materials at different stages are also shown in this figure. Apart from the use of MBF hot metal in mini steel plants, it can also be used in foundries. The products of solid state direct reduction, DRJ and HBI (hot briquetted iron), can be used successfully in the following melting units: Electric arc furnaces, as a substitute for scrap Basic oxygen converters, as a coolant or as a substitute for scrap Open hearth furnaces, as a charge material Induction furnace for melting Ladle metallurgy, as additions Cupola charge, as a partial substitute for cold pig iron Low metalized sponge iron, as a charge material for blast furnaces under certain conditions

II. USE OF SPONGE IRON IN ELECTRIC ARC FURNACES As far as the use of sponge iron for steelmaking is concerned, the predominant reactor is the EAF, where sponge iron is mainly used to substitute scrap, either partially or wholly. This particular route has become extremely popular in mini mills with increasing usage of sponge iron and is, in fact, growing at a faster rate than the conventional BF-LD route. The use of sponge iron in electric arc furnaces has the following advantages:• •

•

Sponge iron does not contain any undesirable impurities, thereby permitting the use of some low grade scrap as a part of the charge without adversely affecting the final steel quality. This can have significant economic implications if scrap prices vary markedly. Since sponge iron does not contain tramp elements like copper, tin, arsenic, chromium, nickel, antimony, lead, etc., its utilization in place of poor quality scrap allows electric steelmakers to meet the more stringent quality requirements being imposed by the end users, such as the wire drawing industry. Figure 10.2 shows the decrease in residuals which can result as an increasing proportion of sponge iron is used in the charge.^ This ultimately leads to improved cold workability of the product, which is especially significant in cold rolling and deep drawing operations. Since the chemical composition of sponge iron is known before it is used, it enables accurate prediction of the end point analysis, beginning with continuous feeding of

201

202

B e y o n d

Lump

Iron ort Sinter

Pellets-

t h e

B l a s t

F u r n a c e

1600 kg

F IG U R E 10.1. Steelmaking processes based on SR (COREX) hoi metal-oxygen converter-secondary metallurgy and continuous casting.

Sponge iron In charge , V®

F I G U R E 10.2.

Influence o f amount of sponge Iron on the level o f residual elements in steel.

Use of DRl/HBI and SRIMBF Hot Metal in Iron- and Steelmaking

•

•

•

2D3

sponge iron. This can result in superimposition of a major portion of the refining requirements on the melting period, thereby contributing to an increase in productivity in electric furnaces. Better cost control can be achieved as the cost of the input materials becomes less sensitive to fluctuations in scrap price. Scrap is a byproduct, and its price depends on market conditions; on the other hand, sponge iron is a product and its price is influenced primarily by the cost of manufacture. The more uniform size of sponge iron compared with scrap lends itself to mechanical handling and continuous charging. This reduces handling operations significantly, which again can contribute to improvements in productivity. Elimination of the need to recharge an EAF twice or three times, as is the case with light scrap, also decreases the heat losses associated with the removal of the roof, thereby improving the thermal efficiency of the process. The unreduced iron oxide present in sponge iron reacts with the carbon present in the bath, resulting in a vigorous “boil” during the continuous charging period. This im­ proves bath heat transfer and slag-metal mixing, resulting in an acceleration of the metallurgical reactions relative to normal sciap melting. This also improves the homo­ geneity of the bath and results in lower hydrogen and nitrogen contents in the steel made, a definite advantage in the case of manufacture of quality steels.^

A. RECENT DEVELOPMENTS IN ELECTRIC ARC FURNACES WITH PARTICULAR REFERENCE TO THE USE OF SPONGE IRON Some recent developments in electric arc furnaces have encouraged the use of sponge iron."

1. Furnace size The size of an electric arc furnace is defined in terms of its shell diameter and transformer rating. Although the shell diameter for a given furnace is chosen largely on the basis of the tap weight required and the steelmaking practice to be adopted (as is the case of DRJ-fed furnaces), the modem trend is to select a larger shell diameter in proportion to the tap weight. This provides adequate room for taking care of the slag making constituents which invariably accompany sponge iron.

2. Furnace classification A method of comparing and classifying electric arc furnaces that has gained international recognition is based on the ratio of the nominal transformer rating in kVA to the tap weight of the furnace in tons. Thus, a 100 ton furnace fitted with a 25 MV A transformer would be rated at 250 kVA/t. On this basis, the Committee on Technology of the International Iron and Steel Institute has proposed a new classification of electric arc furnaces, given below: • • ® •

Low power, 100 to 200 kVA/t Medium power, 200 to 400 kVA/t High power (HP), 400 to 700 kVA/l Ultra high power (UHP), 700 kVA/t and greater

Many existing electric arc furnaces in developing countries such as India, for example, fall in the category of low power, and only some are medium power. This automatically limits the proportion of scrap that can optimally be replaced by sponge iron to 20 to 50%. Furnaces in developed countries generally belong to the last two categories and have been known to consume as much as 100% sponge iron; 70 to 80% sponge iron usage is a routine affair.

20 4

Beyond the Blast Furnace

3. Furnace design The electrical design specifications of a furnace were formerly limited by refractory wear considerations. With the introduction of water cooled panels, now considered essential parts of modem electric furnaces, such restrictions have been substantially relaxed. As graphite electrodes are derived from oil in most cases, their cost has risen sharply. Electrode costs, together with the cost of electricity, are now major factors in the conversion cost of an electric arc furnace. Minimum electrode consumption has, therefore, become a major objective in the design of any new furnace. High secondary voltages and high power factor operations that yield long arcs markedly reduce electrode consumption, particularly when melting the charge. Sponge iron usage offers advantages in these areas.

4. Electrical design features The operating characteristics of an electric arc furnace are determined by the design of its secondary circuit. Before water cooled linings were used, secondary circuits were designed to achieve minimum reactance. As a result, very high electrode currents and very high short circuit currents were employed, which gave rise to high electrode consumption. The high electrode currents also generated large electromagnetic forces, which strained the furnace stmcture. The modem concept of high voltages to achieve long arcs necessitates high reac­ tance values to achieve stable arcing conditions. In this way, not only are electrode currents reduced, thereby minimizing electrode consumption, but so are the short circuit current levels, which serve to reduce the forces acting on the electrode arms. Melting of sponge iron, which is always more difficult than melting of scrap because of its lower thermal and electrical conductivity, has thus become easier in modem arc furnaces.

5. Process control and instrumentation Although instmments used in early arc furnaces were limited to the monitoring of a few parameters, the advent of the modem UHP furnace has entailed a marked increase in the need for instmmentation and diversification in its use. Microprocessors and high power computers are now widely employed in process monitoring and process control. This by itself has made it possible to use large quantities of sponge iron continuously charged through the roof.

6. Electrodes Since electrode consumption cost currently accounts for between 10 and 15% of the conversion cost of steel in an electric arc furnace in most plants, electrode performance is of great importance. The development of UHP furnaces has led to the development of high performance electrodes, the quality of which is largely determined by the raw materials used and the method of manufacture. The mechanism of electrode consumption must be recognized before methods of reducing it can be reliably established. In this connection, it is noteworthy that tip consumption accounts for about 50% of the overall electrode consumption, surface consumption for 40%, and intermediary consumption for the remaining 10%. As tip consump­ tion is directly related to electrode current, raised to a power between 1.5 and 2.0, savings can be effected by operating with lower electrode currents (and longer arcs) at unchanged power input levels. Surface consumption can be reduced by coating the electrode surface and/or by careful operation of the furnace. The level of intermediary consumption, arising from break­ age and stub losses, is reduced by preventing scrap cave-ins, avoiding contact with noncon­ ducting materials, minimizing thermal shocks to the electrodes, increasing the stability of the electrode arms and furnace masts, and improving the efficiency of the electrode regulator. Sponge iron usage definitely has a beneficial influence on at least two out of these three areas of electrode consumption. It must be mentioned that the evolution of electrodes will also depend on arc furnace development. Thus, continuously fed furnaces, as is the case when sponge iron is used, may

Use of DRJIHBl and SRIMBF Hot Metal in Iron- and Steelmaking

205

require electrodes with properties different from those used in scrap-charged furnaces. This aspect has not yet received adequate attention but should be examined, as increasing amounts of sponge iron in the charge will be used in a larger number of arc furnaces in the future.

7. Refractories The beginning of the 1970s witnessed the introduction of water cooled linings, and although this development did not lead to the complete elimination of refractories, it permitted economically acceptable use of high quality products in those small areas which remained refractory-lined. The introduction of water cooled panels has necessitated the development of improved refractories for use in close contact with the panels and has culminated in the development of a new class of magnesite carbon products characterized by high thermal conductivity and high resistance to thermal shock. The presence of unreduced iron oxide, mainly in the form of FeO in sponge iron, can have a harmful effect on the furnace bottom and side walls, but with better quality refractories now available, this aspect can be taken care of to a larger extent.

8. Raw materials The performance of an electric arc furnace is markedly influenced by the physical and chemical characteristics of the materials charged. Some of the iron units now available for steelmaking are in the form of DRl, the real value of which to the steelmaker depends on the comparative price of scrap and, in particular, scrap containing low levels of residual elements. Any arc furnace should be capable of using the most economically available raw materials, thus providing flexibility to the steelmaker. Cyclic shortages of high grades of scrap led to the development of methods for upgrading poor quality scrap, such as fragmenting, cryofragmenting, chipping of large-size scrap, and deoiling and briquetting turnings. The products of these beneficiation techniques can be continuously fed. Although the traditional raw material used in electric steelmaking is scrap, shortages of good quality scrap have caused steelmakers to consider other sources of iron units, such as cold iron in granulated pig or plate form, to supplement scrap. Simultaneously, the development of direct reduction processes has resulted in DRl being used in electric arc furnace steelmaking. Thus, some of the most important technical developments in electric arc furnace steelmak­ ing in recent years have been the introduction of DC arc furnaces, water cooled linings, high quality refractory materials, oxy-fuel burners, long arc practice during melting coupled with larger internal diameter furnaces for a given tap weight, charging of raw materials through the furnace roof, computer control, and secondary steelmaking. Not all of these developments may be applied to any one furnace, nor are they necessarily complementary. It is for each steelmaker to establish which development or combination of developments would be benefi­ cial. By exploiting new technologies, steel production costs in electric arc furnaces can approach those of integrated plants, which is one of the principal reasons why EAFs are steadily becoming more popular for the production of quality steels.

CHARGING AND MELTING PRACTICES Physical and chemical properties of DRl lead, in a logical way, to the continuous charging of this raw material in electric arc furnaces. Several methods of continuous charging of sponge iron are available. Figure 10.3 shows the CONTIMELT system developed by Stelco and Lurgi.^ The beginning of a high percentage (70 to 80%) sponge iron heat is the same as that of an all scrap heat. The scrap portion of the charge and predetermined amounts of carbon and lime or limestone are charged into the furnace. As soon as the roof has been swung back into the operating position, full power is applied to melt the initial charge and bring the bath to a

Beyond the Blast Furnace

206

FIGURE 10.3.

..

FIGURE 10 4

Continuous charging of sponge iron through an EAF roof.

Practice used in high-percentage continuously charged sponge heats.

temperature of about 1565 to 1590°C. The time required for the initial meltdown is directly related to the energy available and, of course, to the size of the initial charge. It has been reported that about 450 to 475 kWh of power per ton of charge is usually required during this stage. Once the desired bath temperature is reached, continuous feeding of sponge iron is immediately started, with full power being maintained throughout the feeding period even though open bath conditions prevail. This high power input is decreased only after DRJ feeding is completed. If the heat has been properly worked, this coincides with the ready-totap time. The progress of a typical heat using a high percentage of continuously charged DRI is shown in Figure 10.4. The best method available at the moment to utilize 100% sponge is to operate on the ‘‘hot heel” principle illustrated in Figure 10.5. In practice, 15 to 25% of the melt from the previous heat is retained in the furnace; this heel is carburized before continuous sponge feeding is started. Maximum power is again applied throughout the continuous charging period, and, with appropriate feed rate adjustments and periodic carbon checks, the furnace is ready to be

Use of DRllHBI and SRIMBF Hot Metal in Iron- and Steelmaking

2D7

Time from stirt of heat, h

FIGURE 10.5.

Practice used for all-sponge heats.

tapped shortly after the completion of continuous charging. A portion of this heat is again retained for the next, and the cycle is repeated. Such a cycle, of course, cannot be continued indefinitely, as the furnace refractories require periodic attention. Therefore, after every three to six such back-to-back heats, depending on the general condition of the lining, the furnace is drained for fettling. The succeeding heat would then be made with a high percentage of sponge iron, and the all-DRJ practice resumed thereafter. The principal variables in all sponge heats are the size of the hot heel and the quality of sponge iron that is charged. The optimum size of the heel is that amount which is just sufficient to provide a suitable working bath into which continuously charged materials can be fed. The quality of sponge iron used will affect the melting rate and, consequently, the productivity because of the iron content of the sponge and also because of the necessity of charging lime to neutralize the acidic gangue components. Productivity would depend, to some extent, on the number of heats that can be made in succession without a pause for fettling.

C. FURNACE PERFORMANCE WITH CONTINUOUSLY CHARGED DRI L Heat time and productivity If sponge iron is charged continuously into any EAF, it reduces the heat time considerably compared with all scrap heats, owing to the elimination of the conventional refining period and better control over the heats, which increases the furnace productivity. Trials conducted with 70 to 80% sponge iron in the charge have typically shown up to 45% increase in productivity on a tap-to-tap basis.^ Not only are the heat times considerably shortened, but they also become more predictable and reproducible. A very interesting correlation between the amount of sponge iron used, melting rate, and productivity is shown in Figure 10.6.^ It can be seen that a minimum tap-to-tap time corre­ sponding to a maximum productivity exists at 30% sponge iron. Beyond this limit, the refining period cannot be further eliminated, and the productivity decrease arises out of an analogous increase in slag weight per ton of steel resulting from the gangue present in the DRI. However, the heat time compared with all-scrap heats is always lower and productivity higher, even up to 100% sponge iron in the charge. As shown in Figure 10.7, the gangue content of sponge iron has a pronounced effect on the productivity of electric arc furnaces.^ The gangue content of sponge iron should always be low (generally less than 5%). Low gangue content results in the maximum increase in productivity, and this remains essentially valid up to 100% sponge iron in the charge. On the other hand, use of high gangue-bearing sponge iron leads to a lower maximum and a rapid decrease in

Beyond the Blast Furnace

208

^

FIGURE 10.6.

20

Correlation between percentage o f sponge iron, melting time, and productivity of an EAF.

25

FIGURE 10.7.

50 Sponge iron in charge . V®

75

100

Electric furnace productivity as a function of the amount and gangue content o f sponge iron used

in the charge.

productivity in the case of charges with a high percentage of sponge iron. Figure 10.7 also explains conditions under which the productivity relative to all-scrap practice cannot be exceeded and is, in some cases, lowered, when sponge iron is used. Figure 10.8 depicts the existence of an optimum level for residual oxygen in sponge iron, corresponding to a maximum productivity; the minimum residual oxygen content is approxi­ mately 0.9%. Nevertheless, excellent results have been obtained over a range of residual oxygen levels varying from 0.6 to 1.2%. As far as the effect of metalization on heal time is concerned, with a degree of metalization higher than 92%, the heal time normally remains constant, as is depicted in Figure 10.9. With lower degrees of metalization, however, heat

Use of DRIIHBI and SRIMBF Hoi Metal in Iron- and Steelmaking

F IG U R E 10.8.

..

F IG U R E 10 9

209

Effect of residual oxygen in sponge iron on productivity (power-on to tap).

Heat time as a function of metalization of sponge iron.

times increase considerably,^ and use of DRJ with less than 88% metalization is best avoided.

2. Energy considerations The power consumption in DRJ-charged heats can either be higher or lower than in all­ scrap heats, depending on the relative importance of various opposing factors, which include the following:

3.

The reduction in heat time, when sponge iron is used, leads to lower radiation losses, resulting in some saving in power. The boiling action caused by the FeO present in sponge iron improves heat transfer in the bath and thus has a favorable influence on power consumption. However, additional power is required for the reduction of this FeO, which is normally more than compen­ sated for by the heat transfer if the degree of metalization of sponge iron is high (more than 90%). A “balanced DRJ'’ is one that has the correct combination of carbon content and residual oxygen linked to iron so that the final reduction by carbon is completed in the EAF. The extra gangue input into the bath through sponge iron requires a larger quantity of fluxes to be added, which tends to increase the power consumption. For low ganguebearing sponge iron (4% gangue max.), the power consumption is lower than in all-scrap

Beyond the Blast Furnace

210

F IG U R E 10.10.

Effect of sponge iron percentage and its gangue content on energy consumption.

86

87

88

89

90

91

92

93

94

95

96

Metallisalion of sponge Iron , V*

. .

F IG U R E 10 11

4.

5.

Meltdown energy as a function of metalization of sponge iron.

heats when the percentage of sponge iron in the charge is less than 85%, as is shown in Figure 10.10.*^ On the other hand, for a high gangue sponge iron, lower power consumption is realized for only about 20 to 4 0 % of sponge iron in the charge. Figure 10.11 shows the comparative meltdown energy data plotted as a function of the metalization of sponge iron. The slope of the curve indicates 20 to 25 kWh/t energy change per percent metalization of sponge iron.’^ Sponge iron has a lower thermal and electrical conductivity than does scrap, which has an adverse effect on power consumption and can result in the formation of “clusters” (icebergs) of unmelted material on the bath surface if the rate of charging is not in harmony with the melting capacity of the bath. The wider power fluctuations experienced when melting scrap are a direct result of the nonhomogeneous nature of scrap and of the continuously varying arc length between the electrode and the scrap on which the arc strikes. These fluctuations increase the electri­ cal reactance of the furnace system and thus reduce the effective power input to the

211

Use of DRJ/HBI and SR/MBF Hot Metal in Iron- and Steelmaking

c

8.

Sponge iron in charge, %

F IG U R E 10.12.

Slag weight as a function of the gangue content of sponge iron.

furnace. The useful power input to the furnace can increase by 10 to 14% as a result of reduced electrical losses when sponge iron is continuously introduced.*' The use of hollow electrodes can also improve furnace operations because the arc flare pattern is primarily downward from the center of the electrode, instead of the outward flaring arcs found with solid electrodes. The improved heat transfer under open bath conditions appears to make hollow electrodes desirable when continuous charging is used. Further, there are indications that the increase in the energy requirements for a given increase in slag volume is less when hollow rather than solid electrodes are used. 3. Slag volume The amount of slag generated in a sponge iron heat is directly related to the amount of gangue present in the DRl, its basicity, and the percentage of sponge used in the heat.*^ Figure 10.12 depicts the extent to which the slag weight increases as a function of the gangue content of sponge iron and its total proportion in the charge. A comparison of the slag amount encountered during all-scrap practice, which is also included in the figure, indicates that with continuous charging of sponge iron, the slag weight does not exceed the value in all-scrap practice unless a high percentage of sponge iron with a large amount of acidic gangue is melted. Figure 10.13 shows that the quantity of additives like CaO and MgO required to maintain a certain slag basicity depends on the amount and nature of the gangue present in sponge iron, which in turn gives rise to an increase in slag weight.'^ It would appear natural that the consumption of lime should increase in a DRl charge in view of the gangue content. However, reduced sulfur in sponge charges often makes it possible to carry out the refining operations with a lower Ca0/Si02 ratio than is normally found in all-scrap heats. Moreover, there is a better utilization of lime when it is fed continuously along with sponge iron. Since energy is applied at a constant rate throughout the melting period, the use of “acidic” sponge requiring greater lime additions leads to higher energy requirements and longer heat times compared with heats made with a “neutral” sponge. It has been found that the change in energy consumption is about 40 kWh/t of steel for each 45 kg change in slag volume. Nevertheless, large slag volumes can be tolerated in continuously charged heats without a penalty in productivity and energy consumption, compared to all-scrap heats, because of the continuous application of full power and the virtual elimination of the conventional refining period.

Beyond the Blast Furnace

21 2

Gangufi conttnt of DRI , */•

F IG U R E 10.13. Percentage of (CaO + M gO) to be added as a function of the gangue content of sponge iron and basicity of the gangue.

4. Electrode consumption The electrode consumption in heats when sponge iron is charged continuously depends on various opposing factors: L

2.

3.

Oxidizing wear at the tip of the electrode depends mainly on the level of power input to the electrode: in this respect, electrode consumption tends to increase with sponge iron. Side wear of the electrode column depends on the atmosphere inside the furnace, the melting time, and the length of time the electrodes are raised and exposed to the atmosphere outside the furnace. During sponge iron usage, the atmosphere inside the furnace is high in carbon monoxide, thus preventing the oxidation of graphite: the overall melting time is shortened, and the period when the electrodes are outside the furnace is decreased because of fewer recharges. All these help in decreasing the electrode consumption. Electrode breakage is considerably reduced when sponge iron is used, compared with the breakages that occur in all-scrap heats that are caused by scrap falling on the electrodes or even by electrical overloading.

Electrode consumption has been found to decrease, typically by 4 to 5%, with high percentages of sponge in the feed when hollow electrodes are used. These electrodes appear to run cooler, possibly because of better heating are under open bath conditions, thus contrib­ uting to decreased electrode consumption. When solid electrodes are used, an increase in consumption has been observed in many cases, while a reduction has been found in those furnaces where electrode breakage is a major problem with all-scrap charges.

5. Refractory consumption The side wall refractory brick consumption for continuously charged sponge heats has been reported to be about 25% less than the consumption during conventional heats. Burning in the critical refractory areas adjacent to the electrodes has been found to be considerably reduced;

Use of DRUHBI and SRIMBF Hot Metal in Iron- and Steelmaking

213

in particular, burning opposite the center or the hottest place decreases by more than 40%. Continuous charging of sponge iron directly into the arc flare regions of the bath apparently creates a “heat sink" effect, thereby reducing the extent of radiation burning of the side walls. The immersion of the arcs in a large volume of foamy slag and the greatly decreased refining times are believed to be the principal reasons for the decreased refractory consumption in the case of continuously charged heats, in spite of the extended periods of high power operation under open bath conditions.

6. Charging rate The rate at which sponge iron can be continuously fed in any EAF is directly related to the rate of energy input. This rate varies slightly with the type of sponge, since the gangue components and neutralizing lime additions have heat capacities that differ from that of the iron portion of the sponge. A so-called “feed index" has been established, which determines the rate at which DRJ can be charged. This feed index is defined as the kilograms of sponge iron per minute which can be fed to the furnace for each megawatt of electrical circuit power input, without causing any change in the bath temperature. The use of hollow electrodes gives rise to a slight increase in the rate at which sponge iron can be fed to any furnace because of improved heat transfer from the predominantly downward flaring arcs. A feed index ranging between 27 and 30 kg/min/MW has been reported for furnaces operating with solid electrodes and using high quality sponge iron with a neutral gangue. 7. Yield The iron yields for all-scrap heats and sponge iron heats are about the same; however, the charge yield (weight of steel produced as a percentage of the weight of metal-bearing charge material) is directly related to the iron content of the DRI. The use of high percentages (70 to 80%) of sponge containing 95% iron has been found to increase the charge yield, typically from 90% in all-scrap heats to 94.2%.'^ DRI can also be batch charged, sandwiched in between layers of scrap in the commercial scrap buckets, but in this case the amount must be restricted to 15 to 25% of the total metallic charge and no possible change in yield is to be expected.

D. FEATURES OF DRI USAGE IN EAFs Based on the above considerations, the following can be stated: Continuous charging of sponge iron in amounts up to 100% of the electric arc furnace charge is possible in HP and UHP furnaces. The best method of producing steel from 100% DRI is to maintain, at all times, a “hot heel" from the previous heat, amounting 20 to 25% of the furnace capacity. The major advantage of using continuously charged sponge iron is the almost complete elimination of the conventional refining period. This is achieved by simultaneous melting and refining, made possible by the use of a known charge material and by the ability to monitor and correct, if necessary, temperature and chemistry while charging is still in progress. The heat can be worked long before the complete charge is in the furnace; this considerably decreases the heat times. Since the heat times are reduced, the rate of steel production increases. An increase in productivity by as much as 45% compared with all-scrap heats has been reported in some cases. The reproducibility and predictability of heal times are fairly high in DRlcharged heats. Energy consumption in continuously charged heats varies with the amount and type of gangue present in the DRI charged. For a low gangue-bearing sponge iron (4% gangue max.), the power consumption is lower compared with all-scrap heats when the percentage

214

•

•

•

•

•

•

•

•

•

Beyond the Blast Furnace of sponge iron in the charge is less than 85%. The use of hollow electrodes appears to be desirable due to improved heat transfer under open bath conditions. With every 1 kg/t increase in the slag weight, the energy requirement increases by about 0.9 kWh/t of steel. However, large slag volumes can be tolerated while still maintaining productivity improvements over all-scrap operations. The optimum range of residual oxygen as iron oxide in sponge iron appears to be 0.6 to 1.2% of the total weight of sponge. In any “balanced” DRJ, the carbon to oxygen ratio should be 3 to 4. For electric furnace smelting, sponge iron should have a consistent analysis, particularly with respect to its carbon and residual oxygen contents. Variation in the amounts of these elements defeats the major advantage of predictable performance otherwise achieved by continuous charging of sponge iron. The extent of decrease in the refractory consumption as reported is more than 40% in certain critical regions. Electrode consumption is identical to that in all-scrap heats, but the frequency of electrode breakage is definitely lower. The use of hollow electrodes can decrease electrode consumption slightly. 15 to 25% sponge iron can also be batch charged along with scrap in EAFs if DRJ is used intermittently, but this is not an optimum usage. Ferro-alloy requirements can be slightly higher due to the virginal nature of sponge iron. The optimum charging rate should be in the range of 27 to 30 kg/min/MW circuit power, and the inclination of the feed pipes should be between 35 and 50 degrees from the horizontal, depending on which electrode must be fed with the material. Because of the continuous nature of the operations, the process can be automated. Energy cost decreases because of the improved electrical characteristics of the arc furnace. The phosphorus content of sponge iron should, preferably, be less than 0.05%. How­ ever, this is difficult to achieve in some cases of coaJ-based DRJ as both iron ore and noncoking coaJs can contain up to 0.08 and 0.2% phosphorus, respectively. This is a definite drawback of sponge iron produced with high phosphorus raw materiaJs. SuJfur content in the sponge iron should be below 0.030%, preferably; this is not at all difficult to fulfill even in all-coal processes and is very easily achieved in gas-based DRJ, where sulfur levels can be as low as 0.008%. The presence of tramp elements can sometimes be a major issue, and the use of even up to 30% sponge iron in the charge has been found to bring down the final contents of nickel, chromium, copper, arsenic, and tin in steels.

The use of sponge iron up to 100% in the charge in HP and UHP electric arc furnaces is a technologically viable proposition, and at times of normal steel demand, it can also be economically attractive. It is interesting to note that in anticipation of a boom in steel demand and the resulting shortage of scrap in the coming years, arc furnace operators all over the world are already thinking in terms of or are already using larger proportions of sponge iron.

III. USE OF HOT DRI IN STEELMAKING Research and development programs for using the sensible heat of DRJ in steelmaking have aJready been undertaken. Methods and systems for linking the DR reactor and the electric arc furnace are being or have been developed so that DRJ can be discharged hot and taken immediately to an EAF or a submerged arc furnace. The heat normaJJy lost in cooling the DRJ can be used in the steeJmaking furnace, thus reducing the electrical energy consumption. One such system is the HYTEMP iron feeding system, developed by HyL.'^ Jt involves an HyL IIJ reactor connected to an adjacent electric arc furnace shop by means of pneumatic

Use of DRJIHBl and SRIMBF Hot Metal in Iron- and Steelmaking

215

HOT DISCHARGE

F I G U R E 1 0 .1 4 .

H y L 's

HYTEMP s p o n g e

iron fe e d in g sy ste m .

222 GCal Iron ore 1150 kg_ Iron ore losses 17 kg ■ Carbon from natural gas 17 kg

LJT

67 kWh

DR Plant

_T — 837 kg DRI = 753 kg iron 0.04 ^ C a l

Scrap 322 kgTotal iron lo sse s 75 kg ^

J I—

5 8 » kW h

EAF

~ r Liquid steel 1000 kg

F I G U R E 1 0 .1 5 .

M a te ria ls and e n e rg y req u irem en t in a c o n v e n tio n a l E A F plant u sin g DRJ.

transporters, as is illustrated in Figure 10.14. Hot DRJ is temporarily held in an insulated inert storage bin at the melt shop, prior to being fed to the furnace using continuous injection, which introduces the material directly onto the metallic bath surface. The temperature of the hot DRJ charged to the furnace is around 650°C. From the time the iron ore enters the charging bins of the reduction reactor until it is discharged hot and transported to the melt shop for charging into the furnace, the material is handled in an enclosed environment. This reduces both material and energy losses and eliminates environmental problems. Such a system for EAFs provides a number of advantages over the conventional system of continuously feeding DRJ through the furnace roof. DRJ is fed in very close to the molten iron bath and pushed toward the hottest zone in the furnace, which is the central portion between the arcs. Fines and small DRJ particles are injected along with the coarse material, and thus the yield is increased. The operation of the EAF also becomes more stable and controllable. Figure 10.15 shows the material and energy require­ ments for a typical EAF using a metallic charge of 70% cold DRJ and 30% steel scrap. The data used in this figure include iron ore of 66.5% Fe to produce DRJ with a total iron of 90%, metalization 92%, and 2% carbon. It can be seen that in a cold DRJ practice, for every ton of liquid steel, the energy input is 9.28 GJ as natural gas and 656 kWh as electricity. Of the latter, 589 kWh are required by the EAF to melt the above mix of cold DRJ and steel scrap. The iron losses from DRJ to the melt shop amount to 75 kg/t of liquid steel. In the case of hot charging of DRJ, using the same assumptions as in the cold DRJ case (i.e., that the charge to the EAF has the same proportion of DRJ and steel scrap), the materials and energy requirements are given in Figure 10.16. Because the iron losses to the EAF in this enclosed system are fewer, the iron required as metallic charge to the furnace is approximately

Beyond the Blast Furnace

216 2.15 GCil Iron ore 1136 kg

LJT

Iron ore losses 17 kg —

‘

Carbon from natural gas 17 kg

”T ~

Total iron losses

OR Plan827 kg. DR I r 744 kg iron

Scrap 31f kg-

63 kg

66 kWh

j j—

§.@4 0 Cil 477 kWh

EAF

T

Liquid steel 1000 kg

FIGURE 10.16.

Materials and energy requirement for hot charging of DRJ in EAF.

FIGURE 10.17.

Effect of temperature of pneumatically-charged hot DRi on energy consumption in EAF.

1% less than when operating with cold DRJ. The energy savings is also quite significant compared with the conventional arrangement. The EAF may now require only around 480 kWh versus 590 to 600 kWh for the conventional configuration, a reduction of around 110 kWh. Thus, there is a savings of about 20% in electrical energy per ton of liquid steel, while the energy requirements for the DRJ producing unit remain unchanged. The higher the temperature of the DRI charged to the EAF, the larger the benefits, as shown in Figure 10.17. The main advantages of such a configuration are summarized in Table 10.1. The reductions in both electrical energy and electrode consumption are on the order of 20%, while the productivity measured in terms of the tap-lo-tap or the heal time can be improved by 15%, along with a significant reduction in refractories consumption.

IV. REPLACEMENT OF SCRAP BY SPONGE IRON IN LD STEELMAKING In an integrated steel plant, if the steel output must be maximized for a given hot metal availability, the only option is to increase the amount of the solid charge in steelmaking. With the adoption of continuous casting in most integrated units all over the world, the availability of internal scrap has decreased considerably. As far as purchased scrap is concerned, the price has increased significantly and its quality has deteriorated. Thus, increased usage of purchased scrap, even if possible, would result in higher operating costs and could give rise to quality problems. As far as LD operation is concerned, productivity decreases as a result of lower yield, mainly because of higher incidence of slopping, increased flame, and fume generation, all of which lead to interruption of hot metal charging and/or the blow proper, if poor quality scrap is charged into the vessel. In addition, the temperature accuracy at the end point is affected; in some cases, use of poor quality scrap is associated with heavy wear in the waste heal boilers in the gas cleaning system, which lowers the quality of LD gas.

Use of DRIIHB/ and SRIMBF Hot Metal in Iron- and Steelmaking

111

TABLE lOJ

Comparison of Cold Charged and Hot Charged DRI Using 70% DRI and 30% Steel Scrap Charging temperature, °C

25 Consumption Power, kWh/t Electrodes, kg/l Productivity Heat time, min Tap-to-tap time, min

589 2.9

650

Benefit

%

477 2.3

112 0.55

19-20 19-20

102

86

16

126

106

20

15-16 15-16

Therefore, in order lo increase production and productivity of LD converters, two possible alternatives are the use of iron ore and/or the use of DRI. Compared to that of purchased scrap, iron ore has a more consistent quality; its increased usage, however, is normally associated with increased tendency toward slopping, which affects the converter yield. Further, since iron ore is at least 3.0 to 3.5 times more effective than scrap as a coolant, the iron units available decrease as the quantity of iron ore is increased. Sponge iron, on the other hand, is a pure source of metallic iron without any tramp elements, and its cooling effect is only slightly (10%) more than that of in-plant scrap. Additionally, it is a high purity material of consistent quality and price, and it combines the metallurgical advantages of iron ore and the process advantages of cooling scrap. The use of DRI in LD steelmaking is, therefore, a technoeconomic possibility. DRI has been successfully used in LD vessels in many parts of the world. Based on the results obtained during a typical trial the following have been reported:'^• •

•

•

• •

•

•

DRI having a high degree of reduction, ranging between 91 to 96%, can be used to replace completely the scrap normally charged in oxygen steelmaking processes. Whether it is charged before or during the early stages of the blow, it does not influence converter performance, but to avoid reoxidation it is important to reduce to a minimum the length of the period for which the DRI is exposed prior to charging. In the manufacture of low carbon steels, the final carbon, phosphorus, manganese, oxygen, and nitrogen contents of the steel tapped would be approximately the same for the DRI as for the scrap charged heats. The use of DRI necessitates the use of an additional amount of lime to maintain the slag basicity normally required for efficient refining. This increases the volume of slag in the converter, and space must be available for the same. Charging of sponge iron is accompanied by an increase in converter slopping, and the resulting loss in the yield is about 5%. DRI tends to remain undissolved fairly long after the commencement of the blow because it clusters together to form aggregates insulated by molten slag. The presence of such undissolved clusters lowers the refining capacity of the slag, which may affect the removal of the metalloids unless the converter is operated with slag containing a higher percentage of iron than in the case of all scrap heats. This would have a deleterious effect on the converter yield and the lining life of the vessel. For the production of very low sulfur steels, DRI constitutes an ideal substitute for scrap. In such cases, however, it is essential to use low sulfur hot metal to extract the maximum benefits from sponge iron, since at high sulfur levels the sulfur partition coefficient can be adversely affected when sponge iron is used. At times when scrap prices are high, the metallurgical benefits arising out of DRI charging could also be supplemented by a savings in the cost of steel production.

21 8

Beyond the Blast Furnace However, at alJ times, it is more economical to use DRJ with as low a content of gangue as possible.

Trials with 100% replacement of scrap by DRJ were also carried out at Tata S te e l.T h e characteristics of the sponge iron used and the test results obtained are summarized in Table 10.2. From the data presented the following can be seen: •

•

•

• ®

•

•

With the use of DRJ, the first turndown conditions in terms of bath chemistry, temperature, and iron percentage in the slag were comparable with those obtained with scrap. Since the turndown condition dictates the lining life of the vessel, it can be inferred that use of DRI did not affect the vessel life markedly. It should be possible to achieve similar vessel lining lives with sponge iron and scrap. The turndown temperatures and iron percentages in the slag were more favorable with DRI, probably because of the more consistent quality of sponge iron and the smoother blow resulting from its use. The gross yield obtained with DRJ was comparable with that of normal scrap-charged heats. The sulfur levels in the steel produced with DRJ in the charge were higher due to high sulfur in some batches of DRJ used in these trials. However, the degree of dephosphorization was adversely affected, probably because of insufficient lime input as well as a distinctly higher phosphorus input into the converter through DRJ. To achieve adequate dephosphorization when using DRI, the lime rate had to be increased by about 5 kg/t of steel, since DRJ made from most Indian ores normally contains at least 0.060% phosphorus, compared to a maximum of 0.040% phosphorus in scrap. The reason for the lower tonnage of DRJ charged in each heal compared with scrap was that the DRJ used contained around 90 to 92% total iron and 80 to 82% metallic iron; its cooling capacity was therefore about 1.10 to 1.15 times higher than that of scrap. Hence, to maintain the same thermal balance, the amount of DRJ charged was decreased by 20% compared with the amount of scrap used in all-scrap heats. The fact that most of the temperatures at turndown were under control showed that this assumption was valid in practice. The slag basicity was comparable in all cases and commensurate with the flux addition. There was no difference in scrap and DRJ heats as far as the slag basicity was concerned. It could not be determined whether there was any influence in furnace refractory wear because of the limited number of heats with DRJ under any one set of conditions, but the MgO content of the final slag, which should increase if there is any increased lining wear, remained almost the same as in scrap-charged heats. The manganese contents at turndown were on the lower side because of the unreduced oxides present in DRJ, and accordingly, ferro-manganese additions in the ladle had to be increased slightly when using DRJ. Nothing unusual was detected in the final product routed either through the conventional ingot route or through continuous casting. On the positive side, the level of residuals in steel was lower in the case of the DRJ heats.

The trials clearly showed that even with relatively poor quality lime, and higher silicon and phosphorus contents in hot metal, sponge iron can be used to replace scrap in LD steelmaking, and even the extent of slopping under these adverse conditions can be controlled. The replacement of scrap by sponge iron in LD converters can offer the following advantages:• • •

A better coolant compared with scrap and/or iron ore (in the form of HBI, it penetrates the slag layer and melts quickly, resulting in a predictable and consistent cooling effect) Elimination of mechanical shocks on the lining during charging

Use of DRIIHBI and SRIMBF Hot Metal in iron- and Steelmaking

219

TABLE 10,2 Results of Trials at Tata Steel with Sponge Iron Vis-à-vis Scrap as the Coolant in LD Converters Physical and Chemical Characteristics of Sponge Iron Used Bulk M etalization,

88-90

Phosphorus,

0.06 (max.)

Sulfur,

Size,

density.

%

mm

kg/m^

0.0250.040

3-25

1.6-2.1

Results Material consumption Hot metal. l Scrap/sponge iron. t Lime, t Dolomite, i Manganese ore, t

With scrap

With sponge iron

138.30 14.09 14.57 5.06 2.04

138.30 12.00 13.75 5.30 2.00

139.04 8.90 14.98 4.44 1.91

27.66 72.34 56.25 43.75 70.00 15.00 15.00 0.16

16.67 83.33 50.00 50.00 40.00 60.00 0.13 —

26.67 73.38 46.80 53.19 85.36 7.32 7.32 0.12

10.20 26.53 36.74 26.53

16.67 50.00 33.33

12.76 6.38 42.55 38.29

71.43 10.20 16.37

66.67 16.67 16.67

59.10 20.45 20.45

65.31 20.41 14.28

50.00 50.00

48.94 25.79 21.28

First turndown condition Carbon up to 0.06% o f heats Over 0.06% ÌT up to 1640 C, % IT over 1640 C, % Slag Fe up to 15%, % 15-18%, % Over 18%. % Manganese in Steel, %

Gross yield, % of heats Up to 80% 81-85% 85-90% Over 90%

Degree of dephosphorization, % of heats LSA “P" beJow 0.030% Between 0.030-0.035% ”P" over 0.035%

Degree of desulfurization, % of heats LSA “S" below 0.030% Between 0.030-0.035% ”S" over 0.035%

Extent of slopping, % Mild Heavy

N otes:

L.S.A. = Ladle sample analysis; N.

4.08 14.28

N.A. N.A.

: No observations available.

6.36 4.25

220 • • • • • •

Beyond the Blast Furnace Good flowability, enabling continuous charging during the blow using the overhead bins Lower slag temperature, resulting in a higher lining life Lower sulfur content and hence a better coolant in the production of electrical steel sheet, ball bearing steel, deep drawing steel, etc. Steel yield comparable to those of all-scrap heats Lower levels of sulfur, nitrogen, and copper in the steel tapped Addition during the final stages of blowing enabling better control of temperature and end point analysis

DRl/HBl is an ideal coolant for combined blown oxygen converters, particularly for the production of high carbon steels where viscous low FeO slags exist. It prevents violent and uncontrolled reactions typical of trim cooling with iron ore. This results in less slopping, higher yield, and easier operation with suppressed combustion hoods and a high percentage of hot metal in the charge. Thus, the problem of scrap availability in integrated steel plants using LD converters can be alleviated by replacing the scrap charged completely by DRJ, which has a consistent chemistry. Other important applications of DRl/HBI in LD steelmaking in smaller amounts are trim cooling in the converter and control of metal temperature in the ladle.

V. ADVANTAGES OF USING HBI/DRI IN STEEL LADLES •

•

•

HBI can be added along with ferro-alloys during tapping or during ladle treatment. It dissolves rapidly in the metal and helps give precise temperature control without delays in operation. Precise temperature control minimizes the consumption of bubbling gases and refrac­ tories and allows operation with the lowest possible temperature during continuous casting. Since HBI/DRI contains a balanced amount of carbon, it often does not change the steel composition.

VI. USE OF SPONGE IRON IN OPEN HEARTH FURNACES DRJ can be used in open hearth furnaces with the following advantages: • • • • •

Increased charging rate because of uniform size of the material and, hence, better utilization of the volume of the charging boxes Better control of melting Improved heat transfer Lower sulfur input into the furnace Better bath agitation

VII. USE OF DRI/HBI IN FOUNDRIES Because of its chemical and physical properties, sponge iron is becoming an important base material in foundries and is being used increasingly to replace pig iron and scrap in cupolas and induction furnaces. In order to reduce the production cost, many foundries replace pig iron with poor quality and less expensive steel scrap in increasing amounts. Because of the contaminants in scrap, however, this creates problems in meeting the quality requirements of the product, especially in the case of ductile iron grades. One of the major disadvantages of using pig iron in some cases is that it contains titanium (from 0.10 to 0.20%) whenever there

Use of DRJIHBI and SRIMBF Hot Metal in Iron- and Steelmaking

221

is an appreciable amount of titania in the ore and coke is used for the production of hot metal. On the other hand, sponge iron does not contain any titanium, even from titania-bearing iron ores, since the reduction potential in all direct reduction processes is such that titania is not reduced to titanium. Although the effect of titanium can be neutralized by the addition of cerium or misch metal, this leads to extra cost. It has been estimated that the additional requirement of inoculant for every 0.1% titanium would be approximately 0.5 to 1.0 kg/t. In addition to the absence of tramp elements, sponge iron normally contains extremely low levels of sulfur, which allows the saving of magnesium alloys in the production of S.G. iron. The applications of DRl include • • •

Replacing part of foundry grade pig iron in cupola operation to produce quality cast iron, Controlling manganese and tramp element levels in the production of ductile iron, Continuous charging of DRl into the induction furnace.

A, SPONGE IRON USAGE IN CUPOLAS Trials have been conducted at a number of sites to establish the fact that sponge iron can replace pig iron/steel scrap as a feed stock for cupolas. The amount of sponge iron in the charge depends on the type of cupola operation. With acid cupola operation, a charge with as much as 70% sponge iron can be used, since the gangue in the sponge iron is mainly acidic; with basic cupola operation, the sponge addition must be correspondingly lower, 15 to 30%, to avoid any additional lining wear problems. In both cases, cupola operation with sponge iron leads to an increase in the coke rate, an increase in the slag volume, and a decrease in the carbon content of cast iron, the exact amount of each depending on the percentage of sponge iron in the charge and the gangue content in sponge iron. At the Rheinstahl Foundry, sponge iron briquettes, making up 40% of the charge, were used, and regular grade cast iron was produced without any change in its metallurgical structure or strength. Based on some trials conducted at Tata Steel, the following was reported: •

•

•

• •

Up to 25% sponge iron can be used easily in a cold blast cupola (to replace pig iron) without causing any significant operationaJ difficulties. With higher percentages of sponge iron in the charge, however, the permeability of the charge tends to decrease; as a result, the gas flow experiences greater resistance. This calls for increased air pressure in the air supply system and decreased air flow. With 25% sponge iron, the air pressure can increase by about 30%. In such cases, care must be taken to consider this aspect before using significant percentages of sponge iron in cupolas. Sponge iron should not be included in the first six to eight charges. If added earlier, it causes difficulty in tapping out the liquid iron. Optimal charging ensures that when the sponge iron reaches the melting area, its temperature is sufficiently high. The slag volume of the cupola increases with an increasing amount of sponge iron in the charge, so care should be taken in slagging. With 25% sponge iron, the slag volume can increase by about 15 to 20%. The liquid iron temperature is not affected by sponge iron addition. The smelting capacity of the furnace is expected to go down with increasing amounts of sponge iron in the charge. The exact effect depends on its gangue content and the degree of metalization. It is, therefore, necessary to choose sponge iron with a low gangue content and as high a degree of metalization as possible for use in cupolas.

The effect of sponge iron addition on the composition of cast iron is shown in Figures 10.18 and 10.19. In general, a decrease in carbon, silicon, and titanium levels was noticed, while sulfur levels remained almost unaffected. The decrease in carbon and silicon levels was due to the fact that sponge contained very little carbon and practically no silicon, and some of the

Beyond the Blast Furnace

22 2

0.090

0.080

3 0.070

-

0.060

FIGURE 10.18.

Effect o f sponge addition on carbon and sulfur contents of cast iron produced.

FIG URE 10.19.

Effect of sponge addition on silicon and titanium contents of cast iron produced.

Use of DRJ/HBI and SR/MBF Hot Metal in Iron- and Steelmaking

223

carbon from coke and silicon from the ferro-silicon charge was consumed in the reduction of residual iron oxide in sponge iron. The lower titanium content of the cast iron produced should make the production of SG iron substantially easier and more economical, which is a distinct advantage of using sponge iron in many cupolas. The advantages of using DRl/HBl in cupola are: • • • • •

Replacement of 15 to 30% of pig iron without significant changes in operating proce­ dures. Lower titanium and aluminum levels and, hence, reduced formation of gas bubbles causing pinholes in the castings. Replacement of ferrosilicon as inoculation material, because of the presence of FeO. Lower consumption of inoculants for spheroidization, when necessary. Ease in handling because HBI is dense, resistant to reoxidation, generates very little fines, and can be stored outdoors, uncovered if necessary.

B. USE OF DRl/HBI IN THE INDUCTION FURNACE An induction furnace with its well known characteristics of uniform power supply, clean melting, and flexibility of operation is better suited for melting sponge iron for the production of quality steel castings. However, the melting of sponge iron in induction furnaces is affected by the difficulties associated with the requirement of a molten pool and quick slag formation. Since the slag is electrically nonconductive, it can be melted only by the absorption of heat from the molten pool of metal. Sponge iron, being much lighter, remains on the top of the pool, leading to the formation of a solid crust, which prevents it from coming in contact with the pool. The National Metallurgical Laboratory (NML) in India^® has made an attempt to: • •

Optimize the melting of sponge iron in the molten iron pool to produce quality low phosphorus and low sulfur iron, and Optimize the melting of sponge iron in low carbon melts to produce high quality low carbon steels. Based on the results obtained, it has been reported that:

• •

The best usage and melting of sponge iron in an induction furnace can be achieved if the particle size of the sponge iron is in close range. The preferred metalization is 90 to 92%, and the density of the particle should be as high as possible (as in the case of HBI). The operating advantages of using HBI in the induction furnace are as follows:

• • • •

Up to 30% HBI can be continuously charged without changing the normal operating procedures. Up to 10% HBI can be batch charged. HBI penetrates the slag and melts rapidly at the slag-metal interface. The low sulfur in HBI can represent significant savings in ladle desulfurization when used in the induction furnace.

Co USE OF DRI FOR BLAST FURNACE IRONMAMNG It should be mentioned that sponge iron, particularly low metalized DRI, can also be used for blast furnace ironmaking.^’ The effect of the degree of metalization on the coke rate and productivity is illustrated in Figure 10.20. It shows a remarkable increase in productivity and

Beyond the Blast Furnace

22 4

FIGURE 10.20.

Use of DRJ in blast furnace ironmaking.

decrease in coke rate, with an increasing degree of metalization. However, use of highly metalized DRJ in ironmaking is unlikely to be economical.

VIII. REOXIDATION OF SPONGE IRON DRl is generally stored and transported in bulk to cater to the needs of steelmakers at places often far from the DRl production units. It is well known that DRl is a nonequilibrium product and has an inherent tendency to be reoxidized in a self-sustaining manner. The rapid reoxidation rate of DRl is attributed to its extremely high surface-to-volume ratio. Most of the chemical reactions involved in the reoxidation process are exothermic, as is shown in Table 10.3.^^ Since DRl has a poor thermal conductivity (approximately 2.09 kJ/mh °C), the exothermic heats of reaction lead to a temperature build-up in any DRJ storage pile, in relation to the oxygen picked up by the material, the heat released amounts to approximately 17 kJ/g of oxygen. Thus, an oxygen pickup of 0.1% would give rise to a temperature increase of about 35°C under adiabatic conditions, which improves the reoxidation kinetics.^^ The heat gener­ ated, coupled with the poor thermal conductivity, can lead to self-ignition of DRl piles. Some fire occurrences during shipment of DRJ are presented in Table 10.43'* The rate of reoxidation when atmospheric oxygen is present (see Appendix lOA), is favored by moisture and temperature, and these two determine the course of the reaction. The three principal reactions contributing toward the reoxidation phenomenon are as follows: 4Fe + 3 O 2 = 2Fe203

(10.1)

2Fe + 3 H2O = Fe203 + 3 H2

(10.2)

2Fe + 2 H2O + 02 = 2Fe (OH )2

(10.3)

While Reaction 10.1 is the usual air oxidation of iron, which is rather slow at low temperatures and becomes prominent only beyond 500°C, the other two are corrosion type

Use of DRUHBl and SRIMBF Hot Metal in Iron- and Steelmaking

225

TABLE 103 Heat Effects During Reoxidation of DRI Exothermic reactions 0.95 3 Fe 2 Fe Fe + Fe + Fe +

Fe + 1/2 0 , + 2 O2 + 3/2 O 2 1/2 O 2 + 1/2 H2 O 1/2 02 + H2 O 3/4 O 2 + 3/2 H2 O

—^ ^ 0 95^ = ^ ^ , 0, = Fe2 0 3 = FeO (OH) = Fe (0H)2 = Fe (OH)^

Endothermic reactions 3Fe + 4 H2 O Fe + 2 H 2O

= + 4 H2 = Fe (OH). + H.

Heat evolved, kj/kg DRI -4 ,6 9 0 -6 ,6 6 0 -7 ,9 5 0 -7 ,9 3 5 -10,175 -1 4 ,7 4 0

Heat absorbed, kJ/kg DRI -M80 +60

TABLE WA Some Fire Hazards Encountered During Shipment of DRI Date

Port

Ship

September, 1977 May, 1979 December, 1979 June, 1981

Lisbon Houston Barcelona Visakhapatnam (India) Santander (Spain)

Cape Crest Medstar Jay Ambika Sanix Ace

March, 1982

Reines

reactions and take place in the presence of moisture at low temperatures. It has been found that gas-based DRI is more prone to reoxidation (it is more pyrophoric), because of its higher porosity and the fine grained structure imparted by hydrogen reduction. Moreover, the nature of deposited carbon also has a role to play in enhancing the sensitivity of gas-based material to be more pyrophoric. Coal-based DRI has less carbon (0.20% max.) compared with the gasbased product (1 to 2%), and because of reduction at higher temperatures, has an outer protective layer around each sponge iron particle. Reoxidation of coal-based sponge iron was studied under the following conditions 1. 2. 3.

Keeping the sponge iron inside a fully enclosed room. Keeping the sponge iron inside a shed without any side cover, so that it was not directly exposed to rain, etc. Exposing the sponge iron directly to the atmosphere, where it was subjected to various changes in weather conditions.

The sponge iron in all cases was kept for 45 days and analyzed at regular intervals of 3 days for changes in degree of reoxidation. Figure 10.21 explicitly shows that there was no signifi­ cant reoxidation in situations 1 and 2, above, while a decrease in metalization occurred in case 3. Sponge iron is, therefore, best not stored outdoors in large piles on the ground. Oxidation is much more severe in the presence of water and during transportation in ships where moist air enhances the kinetics significantly. Constant monitoring of temperature within any DRI pile is essential, and occasional purging with CO 2 or nitrogen can be a temporary solution. Transportation of HBl, on the other hand, is relatively safe because of its high density (or low porosity) and low reactivity.

Beyond the Blast Furnace

226

•s \A s

mc 51

«1r u.

« .S