Resource Utilization of Solid Waste by Thermometallurgy in Steel Processes 9819956544, 9789819956548

Citation preview

Kai Dong Hongyang Wang Zhenqiang Jiang

Resource Utilization of Solid Waste by Thermometallurgy in Steel Processes

Resource Utilization of Solid Waste by Thermometallurgy in Steel Processes

Kai Dong · Hongyang Wang · Zhenqiang Jiang

Resource Utilization of Solid Waste by Thermometallurgy in Steel Processes

Kai Dong Institute for Carbon Neutrality University of Science and Technology Beijing Beijing, China

Hongyang Wang Institute for Carbon Neutrality University of Science and Technology Beijing Beijing, China

Zhenqiang Jiang Institute for Carbon Neutrality University of Science and Technology Beijing Beijing, China

ISBN 978-981-99-5654-8 ISBN 978-981-99-5655-5 (eBook) https://doi.org/10.1007/978-981-99-5655-5 Jointly published with Metallurgical Industry Press The print edition is not for sale in China (Mainland). Customers from China (Mainland) please order the print book from: Metallurgical Industry Press. © Metallurgical Industry Press 2023 This work is subject to copyright. All rights are solely and exclusively licensed by the Publisher, whether the whole or part of the material is concerned, specifically the rights of translation, reprinting, reuse of illustrations, recitation, broadcasting, reproduction on microfilms or in any other physical way, and transmission or information storage and retrieval, electronic adaptation, computer software, or by similar or dissimilar methodology now known or hereafter developed. The use of general descriptive names, registered names, trademarks, service marks, etc. in this publication does not imply, even in the absence of a specific statement, that such names are exempt from the relevant protective laws and regulations and therefore free for general use. The publishers, the authors, and the editors are safe to assume that the advice and information in this book are believed to be true and accurate at the date of publication. Neither the publishers nor the authors or the editors give a warranty, expressed or implied, with respect to the material contained herein or for any errors or omissions that may have been made. The publishers remain neutral with regard to jurisdictional claims in published maps and institutional affiliations. This Springer imprint is published by the registered company Springer Nature Singapore Pte Ltd. The registered company address is: 152 Beach Road, #21-01/04 Gateway East, Singapore 189721, Singapore

Preface

Since the industrial revolution, the measurement of national power is the amount of steel production. For more than 200 years, with the great development of productivity, the scale of human steel production has also been growing, and the expansion of the steel industry has also become a booster of productivity development. By 2022, the global crude steel production has reached an astonishing 1.84 billion tons. The World Iron and Steel Association predicts that the world steel demand will continue to increase by 2.2% to 1.8814 billion tons in 2023. With such huge demand and capacity for steel, it is hard to find a replacement for steel in human production and life in a few future decades. Although steel is the most indispensable material for modern industry production and human life, with the deepening of sustainable development and environmental awareness, the traditional iron and steel metallurgy industry with large pollution and waste has become an increasingly huge challenge for human development. According to incomplete statistics, the iron and steel metallurgical industry produces more than 3 billion tons of metallurgical solid waste by the year, but only a little amount of the waste can pass the solid waste detoxification and reforming regeneration system completely and enter back into the cycle of iron and steel production. A large amount of solid waste has become a pollutant that damages the environment and endangers human life. Dealing with the resources in solid waste effectively has become an important development problem. In recent years, people have begun to think about the feasibility of collaborative treatment of metallurgical solid waste and domestic solid waste in the ironmaking and steelmaking process. According to the process requirements, steel production has an excellent reaction environment such as high temperature, oxidation, reduction and sealed reaction equipment space and has been equipped with complete environmental protection facilities. Based on the iron and steel manufacturing process, the difficulty of solid waste treatment can be greatly reduced. It can not only make complete uses of iron, carbon and other resources in solid waste directly, but also save a lot of iron ore, coal and other resources. It gives great improvements on the comprehensive utilization rate of mineral resources.

v

vi

Preface

The purpose of this book is to summarize the current situation of the treatment and consumption of various metallurgical wastes, domestic waste fertilizers and their products through metallurgical processes at home and abroad. The process characteristics of the whole metallurgical process from sintering, blast furnace, electric furnace, converter and other aspects are collected, as well as with the process parameters for various solid waste treatments, the problems that may be encountered in the relevant process and corresponding solutions, including of changing the process flow or adding some pretreatment links. In addition, some important production and digestion processes are also evaluated in this book, and relevant suggestions for readers to think more comprehensively are put forward. University teachers, graduate students, researchers and engineers engaged in iron and steel metallurgy process, especially those who use iron and steel process to absorb metallurgical solid waste and domestic solid waste together, will get great benefit. This book introduces and summarizes the processes and technologies of metallurgical solid waste and municipal solid waste treatment and recovery of valuable elements in detail. Not only detailed process is described, but also the principles of thermodynamics and dynamics are profound analyzed. Therefore, this is a book that students and scholars need to read. Beijing, China

Kai Dong Hongyang Wang Zhenqiang Jiang

Contents

1 General Situation of Iron and Steel Smelting and Solid Waste Treatment . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1.1 General Situation of Iron and Steel Production . . . . . . . . . . . . . . . . . . 1.2 Advantages of Iron and Steel Process in Treating Solid Waste . . . . . 2 Direct Reduction Iron Process . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.1 Pilot Test Case of Copper Slag Reduction Magnetic Separation in Rotary Hearth Furnace . . . . . . . . . . . . . . . . . . . . . . . . . . 2.1.1 Thermodynamic Analysis . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.1.2 Experimental Scheme of Rotary Hearth Furnace . . . . . . . . . . 2.1.3 Analysis of Experimental Results . . . . . . . . . . . . . . . . . . . . . . 2.1.4 Mineral Phase Analysis of Reduction Products . . . . . . . . . . . 2.1.5 Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.2 Direct Reduction-Melting Experiment of Sulfuric Acid Slag in Rotary Hearth Furnace . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.2.1 Experimental Materials and Methods . . . . . . . . . . . . . . . . . . . 2.2.2 Temperature and Calcination Time Parameters . . . . . . . . . . . 2.2.3 Basicity Parameter Control . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.2.4 Transformation and Micromorphology of Iron Oxides During Reduction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.2.5 Reduction Melting Theory and Experiment of Carbon-Containing Pellets of Sulfuric Acid Slag . . . . . . . 2.2.6 Melting Test Results . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.2.7 Behavior Analysis of Sulfur Element in Melting Process . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.2.8 Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3 Digestion of Solid Waste During Sintering-Pelletizing Process . . . . . . 3.1 Return Sintering of Steel Slag as Raw Material . . . . . . . . . . . . . . . . . 3.1.1 Research Status of Steel Slag Utilization . . . . . . . . . . . . . . . . 3.1.2 Properties of Raw Materials . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.1.3 Sintering Pot Experiment . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

1 1 3 5 6 7 8 13 19 20 21 22 23 27 29 30 32 36 37 39 39 39 40 41 vii

viii

Contents

3.1.4 Experimental Results . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.2 Application of Waste Incineration Fly Ash in Sintering Process . . . 3.2.1 Physicochemical Parameters of Fly Ash . . . . . . . . . . . . . . . . . 3.2.2 Sintering Experimental Procedure . . . . . . . . . . . . . . . . . . . . . . 3.2.3 Proportioning Design of Fly Ash Addition in Sinter . . . . . . . 3.2.4 Sintering Results . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.2.5 Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.3 Formation Mechanism of Binder Phase in Sintering Process with Fly Ash Instead of Lime . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.3.1 Experimental Design . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.3.2 Thermal Dynamic Calculation . . . . . . . . . . . . . . . . . . . . . . . . . 3.3.3 Results and Discussion . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.3.4 Conclusions . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

42 45 47 48 49 52 58 58 59 59 62 74

4 Case Analysis of Blast Furnace Consumption Technology and Process . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 81 4.1 Production Case of Metallurgical Solid Waste Disposal in Tunnel Kiln-Blast Furnace Process . . . . . . . . . . . . . . . . . . . . . . . . . 81 4.1.1 Process Introduction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 82 4.1.2 Major Equipment Technology and Production Capacity . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 84 4.1.3 Energy Consumption of Production Revenue and Expenditure . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 87 4.2 Disposal of All Solid Waste in Copper Slag-Steel Slag Blast Furnace Process . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 88 4.2.1 Solid Waste Raw Materials . . . . . . . . . . . . . . . . . . . . . . . . . . . . 90 4.2.2 Design and Preparation of Artificial Mineral . . . . . . . . . . . . . 91 4.2.3 Reducibility of Artificial Mineral . . . . . . . . . . . . . . . . . . . . . . . 95 4.2.4 Charge Melting Dropping Characteristics . . . . . . . . . . . . . . . . 97 4.2.5 Reduction Melting Results . . . . . . . . . . . . . . . . . . . . . . . . . . . . 100 4.2.6 Material Energy Balance in Blast Furnace Melting Process . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 102 4.3 Prospect of Solid Waste Treatment by Blast Furnace Process . . . . . . 103 5 Disposal of Solid Waste in Steelmaking Process . . . . . . . . . . . . . . . . . . . 5.1 Reuse and Zinc Enrichment Technology of Zinc-Containing Dust in Electric Arc Furnace . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 5.1.1 Brief Introduction of Process . . . . . . . . . . . . . . . . . . . . . . . . . . 5.1.2 Zn Recycling and Enrichment . . . . . . . . . . . . . . . . . . . . . . . . . . 5.1.3 Separation of ZnCl2 by High Temperature Roasting . . . . . . . 5.2 Direct Reduction of Red Mud for Refining Desulfurization . . . . . . . 5.2.1 Simple Method for Na Removal by Red Mud Pretreatment . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 5.2.2 Deep Reduction of Red Mud After Sodium Removal . . . . . . 5.2.3 Desulfurization Design of Steelmaking Refining Slag Prepared from Reduction Products . . . . . . . . . . . . . . . . . . . . . .

105 105 105 106 107 121 121 125 127

Contents

ix

5.2.4 Calculation of Desulfurization Ability of CaO–Al2 O3 –SiO2 –TiO2 Slag System . . . . . . . . . . . . . . . . . . . 130 5.2.5 Desulfurization Experiment of Refining Slag Prepared from Red Mud Reduction Product . . . . . . . . . . . . . . . . . . . . . . 137 References . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 141

Chapter 1

General Situation of Iron and Steel Smelting and Solid Waste Treatment

Abstract The article provides an overview of the iron and steel production process and the treatment of solid waste generated by it. Iron and steel production is a fundamental industry that provides key materials for other industries and plays a significant role in economic development. The ironmaking process extracts crude metal from ore using coke as a reducing agent, while the steelmaking process removes excess elements and impurities to produce high-quality steel. The production process has evolved to include multi-level operations and specialization of functional division. Solid waste generated during production includes slag, dust, and sludge, which require appropriate treatment and disposal to avoid environmental pollution. Various treatment methods, including recycling, reuse, and disposal in designated areas, are employed to manage solid waste. Keywords Development of steel and metallurgy · Long-term steelmaking · Sustainable development · Resource utilization of solid waste · Joint consumption of the metallurgical industry

1.1 General Situation of Iron and Steel Production Iron and steel material are not only the most important structural material used in human society, but also the functional material with the largest output and the most extensive application, which plays an important role in economic development. Although iron and steel are facing the competition from ceramic materials, polymer materials and non-ferrous metal materials (such as aluminum and magnesium) in recent years, the important position of iron and steel in various materials will not change in the foreseeable future due to its obvious advantages in ore reserves, production costs, recoverability rate and comprehensive properties [1]. Iron and steel industry is one of the basic industries of the country, and all industrial departments such as machinery, transportation, construction, national defense and civil use are inseparable from iron and steel materials. Iron and steel output is often the main symbol to evaluate the industry level and production capacity of a country, and the

© Metallurgical Industry Press 2023 K. Dong et al., Resource Utilization of Solid Waste by Thermometallurgy in Steel Processes, https://doi.org/10.1007/978-981-99-5655-5_1

1

2

1 General Situation of Iron and Steel Smelting and Solid Waste Treatment

quality and variety of iron and steel have a great influence on the quality of products in other industrial sectors of the national economy [2]. Iron and steel metallurgy is a process in which metals are extracted from iron ore or other valuable waste resources according to the basic principles of physical chemistry, thermodynamics, kinetics, transmission principle, reaction engineering and metallurgy, and then refined and then made into iron and steel materials with certain properties by various processing methods. Currently, iron and steel production adopt pyrometallurgical process, which has always used two-step smelting method, namely steelmaking and steelmaking main processes. Among them, ironmaking process provides key raw materials for steelmaking, and steelmaking process can also be regarded as further processing of iron and steel. (1) Ironmaking. Coke is mainly used as fuel and reducing agent to extract crude metal from ore or concentrate. Under the reduction condition in blast furnace, ore is reduced to obtain crude metal-pig iron, in which carbon (4–5%) from reducing agent and impurities in ore and gangue are dissolved, such as Si, Mn, S, P and other elements. (2) Steelmaking. Excessive elements (C, Si, Mn) and impurities (S, P) are removed by oxidation and chemical reaction involving slag to reach the limit of no harm to steel grade properties. Meanwhile, the dissolved gas (H, N) in liquid steel and oxygen (deoxidation) introduced into liquid steel by oxidation should be removed, and the composition of liquid steel should be adjusted. Finally, the liquid steel with qualified composition is poured into ingots or billets, which is convenient for rolling. In the main process of traditional iron and steel production, the core smelting tasks of ironmaking and steelmaking are completed by blast furnace and converter respectively, and the final qualified liquid steel is poured into ingots by die casting. With the expansion of iron and steel production scale and the improvement of smelting efficiency, it is impossible to complete many complex smelting tasks in a single metallurgical furnace. The changes of modern iron and steel production process are characterized by multi-level operation and specialization of functional division. With the change of iron ore raw material supply and the operation requirements of concentrate production in blast furnace, blast furnace production has developed from the earliest direct feeding of iron ore to the feeding of clinker such as sinter and pellet, and gradually developed into a series of process steps before ironmaking such as coking, sintering, pelletizing and pulverized coal; in order to get rid of the dependence of iron and steel production on coal, especially coke, many non-blast furnace ironmaking methods have appeared one after another, among which the smelting reduction ironmaking process represented by COREX is more successful; after “pretreatment” (desilication, dephosphorization and desulfurization), the molten iron from the blast furnace is supplied to the converter for steelmaking, so as to reduce the heavy slag burden of removing these elements from the converter and compensate the deficiency of desulfurization ability of oxidized slag of the converter, thus realizing the blowing process with less slag and high productivity. In order to improve the productivity of general steelmaking process

1.2 Advantages of Iron and Steel Process in Treating Solid Waste

3

and the quality of liquid steel (further reduce the content of impurities and gases), some refining processes of steelmaking process are transferred to steel drums outside the furnace or special reaction furnaces for continuous or deep completion. At the same time, the large-scale application of liquid steel continuous casting has greatly changed the production efficiency and benefit of iron and steel enterprises. Therefore, the traditional “long process” of iron and steel production with iron ore as the initial raw material to produce qualified liquid steel gradually evolved into: process before iron making—blast furnace—hot metal pretreatment—converter—secondary refining-continuous casting. On the other hand, with the abundance of iron and steel materials accumulated in human society, the “short process” of iron and steel production using scrap iron and steel materials (scrap steel) to produce qualified liquid steel is booming. The traditional electric arc furnace smelting task is divided and handed over to electric arc furnace and external refining respectively, and gradually developed into: raw materials (scrap steel)-electric arc furnace-secondary refining-continuous casting. Meanwhile, the direct reduction method in non-blast furnace ironmaking uses gaseous fuel, liquid fuel or non-coking coal as energy, and reduces iron ore (or iron lump) below solid softening temperature to obtain direct reduced iron (DRI or sponge iron), which is usually added into electric arc furnace as solid metal burden. Qualified liquid steel with low content of impurity elements P, S, O, H and N and certain amount of C or some alloy elements is obtained by smelting in “long process” or “short process” of iron and steel production, which is poured into semi-finished slab by continuous casting machine, and then heated, hot rolled, pickled, cold rolled, heat treated, annealed and surface processed to become the final steel product.

1.2 Advantages of Iron and Steel Process in Treating Solid Waste Solid waste generally includes municipal solid waste and industrial solid waste. Urban solid waste is increasing at an annual rate of 8–9%, and the annual output of urban per capita domestic waste is about 450–500 kg. With the increase of population and urbanization, the pressure of urban garbage disposal gradually increases. Currently, 80% of urban garbage in China is still treated by landfill, and the comprehensive utilization rate of urban construction garbage resources is less than 30%, far lower than 90% in Japan, South Korea and the European Union. In the early stage, landfill was just a simple landfill without considering the gas and liquid pollution in the waste, resulting in serious environmental pollution. Since the twenty-first century, with the progress of related technologies in China, the problem of landfill seepage control has been effectively solved. Advanced HDPE membrane seepage control technology has been adopted in many newly built sanitary landfills in Guangzhou and Shenzhen. In addition to landfill treatment, another major treatment measure is incineration. Municipal solid waste incineration can effectively reduce volume

4

1 General Situation of Iron and Steel Smelting and Solid Waste Treatment

and weight, but dioxin produced during incineration will cause irreversible harm to human body. With the development of incineration technology and improvement of incineration equipment, dioxin emission has been effectively controlled. Any single treatment process cannot make municipal solid waste fully up to standard. However, due to the cost limitation of multi-process collaborative treatment, there are still a considerable number of landfill anti-seepage requirements until 2011. In recent years, with the emphasis on environmental protection and the improvement of technology, as well as the implementation of waste classification pilot measures, it is expected to solve the problem of urban solid waste treatment in China. Compared with urban solid waste, industrial solid waste is more polluting and contains many heavy metal elements, which will cause far-reaching negative effects if not handled properly. Industrial solid waste is mainly composed of dust, slag, waste and other wastes produced in industrial production, mainly from mining, chemical industry, non-metallic mineral processing, power gas production, ferrous and nonferrous metallurgy and other industries. As far as the ferrous metallurgy industry is concerned, the crude steel output in China will reach 1.053 billion tons in 2020, and the steel output ranks first in the world. With the increase of iron and steel production year by year, the environmental pollution caused by various wastes is also deepening day by day. Generally, the amount of dust produced by iron and steel enterprises is 8–15% of the steel production. Based on this calculation, the annual amount of dust produced by iron and steel enterprises in China is about 100 million tons. The necessity and urgency of metallurgical dust treatment put forward high-quality requirements for scientific researchers. Pyrometallurgical treatment of solid waste is a kind of treatment method, which uses carbon or pulverized anthracite in dust as reducing agent to reduce metal oxides in dust and recover some or all valuable elements at a certain high temperature by using the difference of reduction temperature and melting boiling point of metal oxides. The main treatment processes are as follows: sintering method, pelletizing method, dust injection method, direct reduction method and briquetting return steelmaking method, etc. According to the process requirements, iron and steel production has excellent reaction environment such as high temperature, oxidation and reduction, and sealed reaction equipment space, which has been equipped with perfect environmental protection facilities. Based on the steel manufacturing process, the difficulty of solid waste treatment can be greatly reduced, which can not only directly make full use of iron, carbon and other resources in solid waste, but also return to the steel production process for use, save a lot of resources such as iron ore and coal, and greatly improve the comprehensive development and utilization level and comprehensive utilization rate of mineral resources.

Chapter 2

Direct Reduction Iron Process

Abstract The article examines the impact of alkalinity, particle size, temperature, and mineral analysis on direct reduction roasting. R = 0.6 is best for alkalinity, but 0.4 suits magnetic substances. Finer particle size yields higher iron grade but reduces metal recovery. Optimum calcination temperature is 1175 °C. Copper’s trend in magnetic material follows iron grade, and Cu and iron phase separate when crushing to 42 µm. Metallic iron is dispersed in the glass phase, while copper is embedded in iron as pyrite and matte. The paper studies recycling sulfuric acid slag and extracting iron through rotary hearth furnace ironmaking. Optimal metallized pellet direct reduction parameters are: carbon 1.2, calcination temperature 1200 °C, calcination time 60 min, yielding TFe% = 73.2%, MFe% = 61.8%, and metallization rate up to 91.5%. Metal recovery reaches 92.6% under the optimal melting experiment parameters of alkalinity 1.2, temperature 1450 °C, carbon 1.2, and melting time 20 min. Metallized pellets’ poor desulfurization ability leads to sulfur content of about 1%, making it ideal for producing sulfur-based free-cutting steel. Keywords Mineral analysis · Direct reduction roasting · Sulfuric acid slag · Direct reduction · Metal recovery

Direct reduction ironmaking is a technological process of smelting ore into metallic iron under the condition of solid ore or solid–liquid mixed state [3]. Compared with traditional blast furnace ironmaking, direct reduction process has the characteristics of strong adaptability of raw materials, flexible and simple operation and low investment in the field of solid waste treatment. This process can be realized by rotary hearth furnace, rotary kiln, shaft furnace and other equipment.

© Metallurgical Industry Press 2023 K. Dong et al., Resource Utilization of Solid Waste by Thermometallurgy in Steel Processes, https://doi.org/10.1007/978-981-99-5655-5_2

5

6

2 Direct Reduction Iron Process

2.1 Pilot Test Case of Copper Slag Reduction Magnetic Separation in Rotary Hearth Furnace As a major copper producing country in the world, the amount of copper slag discharged in China is about 8 million tons per year. The slag contains many valuable metals such as Fe, Cu, Zn, Pb, Co and Ni, and a small number of precious metals such as Au and Ag, among which the iron content is much higher than the recoverable grade of iron ore in China [4]. The utilization rate of copper slag in China is very low, and most of copper slag is piled in slag yard, which not only occupies land, but also pollutes the environment and causes huge waste of resources. At present, copper slag is mainly used to recover non-ferrous metals such as Cu, Zn, Pb and Co from copper slag by different methods, except for a small amount of copper slag used as cement concrete raw material and antirust abrasive [5, 6]. This chapter is a roasting reduction-magnetic separation process in rotary hearth furnace [7]. In the experiment of 500 kg rotary hearth furnace, the temperature and holding time in the furnace were controlled by frequency converter and temperature feedback. In this experiment, copper slag supplied by a smelter in Yunnan was used as raw material, pure calcium carbonate (CaCO3 ≥ 99.0%) was used as additive, and anthracite pulverized coal was used as reducing agent. The composition of copper slag is shown in Table 2.1, and the composition of pulverized coal is shown in Table 2.2. Mineral phase analysis of powdery copper slag shows that the granularity is mostly between 0.002 and 0.6 mm, the morphology is mostly flaky, and the minerals are mainly silicate liquid slag and a small amount of olivine minerals. The content of silicate liquid slag is between 75 and 80%, and the content of magnetite grain is more, reaching about 15 and 20%, as shown in Fig. 2.1a, and the grain size is mostly between 0.1 and 10 µm. 2–3% of chalcopyrite and magnetite are scattered among silicate liquid slag, and the particle size is basically equal to magnetite grains, as shown in Fig. 2.1b. Copper slag is mainly magnetite, chalcopyrite and silicate liquid slag. Table 2.1 Chemical composition of copper slag (mass fraction) Composition

TFe

FeO

SiO2

CaO

MgO

Al2 O3

Cu

S

Pb

Zn

MnO

Content, %

38.9

44.9

32.1

3.34

2.28

4.68

0.61

1.04

0.26

1.95

0.24

Table 2.2 Chemical composition of pulverized coal Composition

Fixed carbon

Ash content

Volatile matter

Analytical water

Content, %

85.85

5.44

8.30

0.46

2.1 Pilot Test Case of Copper Slag Reduction Magnetic Separation in Rotary …

7

Fig. 2.1 Micrograph of copper slag. a 1-magnetite, 2-silicate liquid slag; b 1-chalcopyrite, 2-olivine

2.1.1 Thermodynamic Analysis The existing form of iron element in copper slag is complex. According to literature, XRD and mineral phase analysis of raw materials, it can be seen that iron element in copper slag mainly occurs in magnetite and olivine [8, 9]. The reduction behavior of Fe2 SiO4 and Fe3 O4 is different. When the temperature is higher than 843 K, Fe3 O4 is reduced step by step in the following order: Fe3 O4 → FeO → Fe. Fe2 SiO4 is decomposed into FeO at 298–1600 K, and then reduced to metallic iron. Gibbs free energy and oxygen potential of each element in the reduction process of copper slag are shown in Table 2.3 and Fig. 2.2. As shown in Fig. 2.2, the higher the direct reduction temperature, the smaller ΔGº , indicating that the reduction reaction is more likely. Fe3 O4 in copper slag can be easily reduced to metallic iron, and Fe2 SiO4 can also be reduced to metallic iron Table 2.3 Thermodynamics data of main reduction reaction No.

Reaction formula

ΔGº , kJ/mol

Critical temperature, K

1

ZnO(s) + C(s) = Zn(g) + CO(g)

344,555–281.2T

1225.3

2

PbO(l) + C(s) = Pb(l) + CO(g)

87,665–165.9T

528.4

3

Cu2 O(s) + C(s) = 2Cu(s) + CO(g)

51,045–156.3T

326.6

4

Fe3 O4 (s) + 4C(s) = 3Fe(s) + 4CO(g)

615,880–676.1T

910.9

5

Fe2 SiO4 (s) + 2C(s) = 2Fe(s) + SiO2 (s) + 2CO(g)

354,140–341.6T

1036.7

6

Fe2 SiO4 (s) + CaO(s) + 2C(s) = CaSiO3 (s) + 2Fe(s) + 2CO(g)

235,347–310.7T

757.4

8

2 Direct Reduction Iron Process

Fig. 2.2 Oxygen potential diagram

when the direct reduction temperature is higher than 1036.7 K. If CaO is added in the direct reduction process, it can react with Fe2 SiO4 to release simple iron oxides [10], thus reducing the direct reduction temperature and promoting the reduction of Fe2 SiO4 .

2.1.2 Experimental Scheme of Rotary Hearth Furnace Raw materials (iron concentrate, anthracite and binder) are batched by electronic scales and spiral weighing machines. After the mixture is milled by a mixing mill and evenly mixed, the raw balls are dried and screened out by a disc feeder, and then uniformly distributed on a rotary hearth furnace by a vibrating feeder. After rotating for one week at the furnace temperature of 1300–1380 °C, the metallized pellets are pushed out by a spiral discharge machine, and the finished products are cooled indirectly by a water jacket, and finally stored in a finished product warehouse for the next working procedure. The rotary hearth furnace body has five sets of burners, two sets of thermocouples (central high temperature area), flue gas outlet and sampling point (Fig. 2.3). The operation of rotary hearth furnace shall be carried out according to the following process. (1) Heat up and dry the oven. Turn on the gas and air burners, and raise the temperature in the furnace to 1300–1350 °C through combustion. (2) When the temperature rises to the target temperature, measure the flue gas outlet and gas composition online. Record the atmosphere in the furnace (CO–CO2 ratio). Keep the atmosphere in the furnace stable before formal feeding.

2.1 Pilot Test Case of Copper Slag Reduction Magnetic Separation in Rotary …

9

Fig. 2.3 Main part of rotary hearth furnace

(3) Place all pellets on customized trays in batches and put them into the furnace in batches. Take it out after rotating once with the furnace body. Here, it has been identified as a Shi Ying tray, and the laboratory is currently exploring whether it can be reused. (4) Water quench, collect and dry the semi-molten pellets in the tray (drying is also needed here) to the magnetic separation process. In order to simulate continuity and combine with industrial production as much as possible, continuity and stability need to be considered. In order to simulate continuity in the experimental process, it is combined with industrial production as much as possible. Continuity is mainly reflected in continuous feeding and discharging of pellets. Feeding, discharging and pellet pallet methods are shown in the following schematic diagram (Fig. 2.4). The experimental rotary hearth furnace does not have the condition of continuous feeding and discharging. During the experiment, customized Shi Ying was used to pre-load pellets. Put the pellets on the plate, push them into the rotary hearth furnace in batches, and take out the quartz plate after rotating for one week and reducing. Because the rotary hearth furnace runs slowly, there is enough time for feeding and discharging. By using this method, continuous feeding and discharging can be realized as much as possible. The effective area of rotary hearth furnace is 19,300 cm2 , and the effective area of single tray is 25 cm × 30 cm and 500 cm2 . The furnace can hold 25–30 pallets at the same time. The effective height of the tray is 7 cm. It can hold 10–15 pellets. Up to 250–300 pellets can be loaded in a single furnace. For the experimental amount of 500 kg (experimental time is 20 h), 40 trays need to be prepared for recycling. After the pellets are loaded into the quartz tray, the quartz tray is put into the rotary hearth furnace, because the number of trays in the furnace is insufficient in the early stage and the late stage. The scope of experimental investigation is that when the first tray is released from the oven, the tray is the first experimental group. The tray taken out after the last tray was put in was the last experimental group. Calculated

10

2 Direct Reduction Iron Process

Fig. 2.4 Rotary hearth furnace test and distribution diagram

according to 3 h circle. 3 h after the first tray is placed and 3 h before the last tray is taken out, the pellets are not magnetically separated. The pellets in the middle time period are considered to be carried out under uniform and stable conditions. Temperature control in rotary hearth furnace is very important. The temperature distribution of rotary hearth furnace is as Fig. 2.5 (previous experimental measurement data and numerical simulation results). There are three thermocouples in the rotary hearth furnace, which can measure the temperature in the core high temperature area (Fig. 2.6). Measure the temperature before starting the furnace, and after the temperature measured by thermocouple reaches the set requirements and remains stable. Start the material feeding operation.

Fig. 2.5 Use intention of temperature range in rotary hearth furnace

2.1 Pilot Test Case of Copper Slag Reduction Magnetic Separation in Rotary …

11

Fig. 2.6 Installation position of spray gun and thermocouple

During the experiment, if the temperature fluctuates, it is adjusted by adjusting the gas and air supply volume. The furnace temperature range of rotary hearth furnace is 1200–1300 °C, and the temperature in high temperature zone is 1300–1320 °C (the temperature range is too large, and the temperature target in high temperature zone is 1300 °C plus or minus 20 °C). According to the experimental requirements, the temperature can be adjusted by thermocouple feedback data. The main problem is. Furnace temperature is uneven. The holding time should be extended accordingly. The recommended time is 3.5–4 h. The atmosphere control in rotary hearth furnace has a direct influence on the reduction result. The fixed end of the rotary hearth furnace is equipped with a flue gas exhaust device, and the side edge is equipped with a gas detection window (Fig. 2.7). Used for detecting flue gas components, hand-held flue gas analyzer is adopted, and the response is fast. By analyzing the flue gas results, adjust the atmosphere in the furnace. Before the experiment begins, the experimental temperature and atmosphere are preset. During the experiment, try to keep the atmosphere unchanged, and the fluctuation in a short time should be compensated by the operation of air supply. The atmosphere is mainly coordinated by the combustion ratio of gas, the amount of gas and the flow rate of combustion-supporting air. If the CO2 content is too

12

2 Direct Reduction Iron Process

Fig. 2.7 Smoke sampling port

high, the gas flow rate will be increased and the air flow rate will be appropriately reduced. Under the condition of ensuring stable temperature, the CO–CO2 ratio can be adjusted in an appropriate range (when the CO content is too high, the gas flow rate will be appropriately reduced and the air flow rate will be increased to achieve the same effect). Gas equipment and gas control device are shown in Fig. 2.8. In order to ensure the reduction stability of the atmosphere, a layer of graphite powder or carbon powder is placed at the bottom of the plate to maintain the stability of the atmosphere and prevent the fluctuation of the atmosphere from having a great influence on the reduction rate. Combined with the previous laboratory experiments, in the reduction stage of copper slag, the effects of reduction temperature and alkalinity on reduction products were mainly investigated, and in the magnetic separation stage, the effects of different crushing particle sizes of reduction products on magnetic separation were mainly investigated under corresponding conditions. The contents and plans of the investigation are as follows:

Fig. 2.8 Gas equipment and gas control device

2.1 Pilot Test Case of Copper Slag Reduction Magnetic Separation in Rotary …

13

(1) The influence of alkalinity change (R = 0. 4, 0.6, 0.8, 1.0) on reduction calcination, calcination time T = 75 min, calcination temperature T = 1150 °C, carbon content C/O = 1.2; (2) The influence of temperature change (T = 1150, 1175, 1200 °C) on reduction roasting, roasting time 75 min, carbon content C/O = 1.2, alkalinity R = 0.4; (3) The corresponding reduction products are crushed to different particle sizes for magnetic separation, and its influence on the magnetic separation products is analyzed.

2.1.3 Analysis of Experimental Results The influence of different basicity on copper slag treated by coal-based direct reduction is shown in Fig. 2.9. The experimental conditions are as follows: reduction time t = 75 min, temperature T = 1150 °C, carbon content C/O = 1.2. As shown in Fig. 2.9, with the increase of alkalinity, metallization first increases and then decreases. When alkalinity is 0.6, metallization rate is the highest, which is consistent with previous theoretical analysis; The total iron content in the reduction product decreases gradually with the increase of alkalinity, which is due to the increase of CaO and the increase of non-iron content with the increase of alkalinity, which leads to the decrease of iron grade; The changes of metallic iron content and metallization rate were observed. When the alkalinity was 0.6, the content of metallic iron was relatively high, and the metallization rate showed a downward trend when the alkalinity was increased, which indicated that the proper increase of alkalinity was beneficial to reduction, but too high alkalinity would hinder reduction.

Fig. 2.9 Effect of alkalinity on reduction products

14

2 Direct Reduction Iron Process

The reduction products with different basicity were separated by magnetic separation, and the influence of magnetic separation powder particle sizes of 154 and 71 µm on the results was investigated, as shown in Fig. 2.10. With the increase of alkalinity, the iron grade, Cu content and metal recovery rate of magnetic substances after magnetic separation show a gradual downward trend. It is due to the increase of basicity and the increase of slag amount, which is not conducive to the aggregation and growth of reduced iron grains and the dispersion and distribution of metallic iron, and the iron-containing grains in the original slag are generally less than 50 µm (Fig. 2.10), and the selected small particles have more gangue components and lower iron grade; Cu may be dissolved in iron phase, the particle size of crushing is not enough, and magnetic separation can not separate iron from copper, so its change law is consistent with iron. Further grinding the reduction product can improve the iron grade. Compared with Fig. 2.10a and b, the iron grade of magnetic separation product with particle size of 71 µm is greatly improved compared with that with particle size of 154 µm, but the metal recovery rate is reduced. XRD patterns of copper slag before and after magnetic separation with different basicity are shown in Fig. 2.11. As shown in Fig. 2.11a, in the process of alkalinity from 0.4 to 1.0, silicate iron-containing compounds in calcination products gradually disappear, calcium silicate phase in non-magnetic products obviously increases, and ferrous silicate phase greatly decreases. When alkalinity R = 1.0, ferrous silicate phase basically disappears. Due to the addition of calcium carbonate, iron in most iron-containing silicates can be replaced in the form of simple iron oxides, which

Fig. 2.10 Effect of alkalinity on magnetic substances after magnetic separation. a 154 µm; b 71 µm

2.1 Pilot Test Case of Copper Slag Reduction Magnetic Separation in Rotary …

15

greatly reduces the reduction temperature and improves the reduction rate. As shown in Fig. 2.11b, XRD analysis results of magnetic substances show that only when the alkalinity is 0.6, the magnetic substances are basically metallic iron, which proves that low or too much alkalinity is not conducive to the separation of slag iron in magnetic separation. As shown in Fig. 2.11c, the XRD analysis results of slag phase after magnetic separation show that with the increase of alkalinity, the content of calcium silicate increases and the content of ferrous silicate decreases, indicating that increasing alkalinity has obvious effect on destroying the structure of ferrous silicate. Based on the material index of magnetic separation of the final product and the principle of less introduction of foreign flux, it is considered that the alkalinity of 0.4 is the best for the comprehensive effect of reduction and magnetic separation, and the obtained magnetic material has higher high iron content, and the reduction effect and metal recovery rate are also ideal. Temperature has a significant influence on chemical reaction, and the higher the temperature, the faster the chemical reaction. Figure 2.12 shows the effect of temperature on reduction products. The reduction temperatures are T = 1150, 1175 and 1200 °C. Other experimental parameters are: reduction time t = 75 min, alkalinity R = 0.4, carbon content C/O = 1.2, magnetic separation particle size 71 and 42 µm. As shown in Fig. 2.12, with the increase of temperature, the total iron content in direct reduction products first increased and then decreased slightly, while the metallic iron content and metallization first increased and then decreased. Although high temperature is beneficial to chemical reaction, too high temperature will melt ferrite, a low melting point substance in raw materials, which will be wrapped on the surface of minerals and hinder the reduction from continuing. In addition, too high temperature will accelerate the consumption of pulverized coal, which will rapidly weaken the reducing atmosphere in the furnace. In the later stage, the reduction atmosphere was weaker, and the reaction basically stagnated, so that the best reduction opportunity was lost, and finally the metallization rate, total iron and metallic iron content decreased [11, 12]. XRD patterns of copper slag before and after magnetic separation at different temperatures, as shown in Fig. 2.13. With the increase of temperature, the content of Fe3 O4 in the reduction product (Fig. 2.13a) gradually decreases. XRD results verify the theoretical analysis that high temperature is beneficial to iron reduction. Among the magnetic materials (Fig. 2.13b), only 1175 °C is basically full of metallic iron, which shows that the magnetic separation effect is the best at 1175 °C. The reduction products at different temperatures were separated by magnetic separation, and the effects of magnetic separation particle sizes of 71 and 42 µm on the results were investigated, as shown in Fig. 2.14. The change trend of the two diagrams is basically the same. With the increase of temperature, the metal recovery first decreases and then increases, but both remain in a relatively stable range, and the change trend is not obvious. However, the iron and Cu contents in magnetic products increased first and then decreased, which was closely related to the direct reduction process. The main reason for this trend was that temperature had two sides on the reduction. Temperature increase was beneficial to chemical

16 Fig. 2.11 XRD patterns before and after magnetic separation with different alkalinity. a Direct reduction of the product prior to magnetic separation; b Magnetic material after magnetic separation; c Non-magnetic material after magnetic separation

2 Direct Reduction Iron Process

2.1 Pilot Test Case of Copper Slag Reduction Magnetic Separation in Rotary … Fig. 2.12 Effect of temperature on reduzate

Fig. 2.13 XRD diagram before and after magnetic separation at different temperatures. a Direct reduction of the product prior to magnetic separation; b Magnetic material after magnetic separation

17

18

2 Direct Reduction Iron Process

Fig. 2.14 Effect of temperature on magnetic separation products. a 71 µm; b 42 µm

reaction kinetics and thermodynamic conditions, but it also made low melting point substances encapsulate minerals and hindered the reaction from continuing. The total iron grade of magnetic products is the highest at 1175 °C (in which, Fig. 2.14b can reach 74.7%, and the metal recovery rate is 73.90%). The change of copper content is consistent with that of iron, indicating that some copper elements may exist in iron phase. Comparing Fig. 2.14a with b, the changes in metal recovery and total iron grade are basically similar to those in Fig. 2.10, but the changes in Cu are different. In Fig. 2.10, the Cu content with large particle size is lower, while in Fig. 2.14, the Cu content with small particle size is lower. It is considered that Cu changes with the content of Fe when the particle size is larger, mainly because Cu inclusions can not be separated in Fe phase; When the particle size is small enough, Cu can no longer attach to the iron phase, and magnetic separation can separate the two. When the particle size reaches 42 µm, the content of Cu in magnetic materials decreases, which shows that Cu has obvious separation ability from iron phase at this particle size. The change trend of the influence curve of temperature on magnetic separation products in Fig. 2.14 not only shows the two sides of the influence of temperature on magnetic separation products, but also shows that small crushing particle size has a positive effect on magnetic separation. However, in the actual process, it is difficult to crush the particle size to a smaller size, which will lead to increased energy consumption and equipment loss. It is necessary to determine the appropriate

2.1 Pilot Test Case of Copper Slag Reduction Magnetic Separation in Rotary …

19

crushing particle size according to the actual process capacity and product index requirements. In conclusion, the optimum roasting temperature of coal-based direct reduction treatment of copper slag is 1175 °C, the iron grade can reach 74.7%, the metal recovery rate is 73.90%, and copper and iron phase begin to separate under the condition of 42 µm crushing particle size.

2.1.4 Mineral Phase Analysis of Reduction Products The reduction samples of solid copper slag (T = 1175 °C, R = 0.4, C/O = 1.2, t = 75 min) with the best experimental results were taken out, bonded with epoxy resin glue, and after consolidation, the samples were ground four times and polished one time to make optical films, and the mineral phase was observed. Figure 2.15 and Table 2.4 are the result of scanning electron microscope-energy spectrum analysis. The samples are mainly glass phase, metallic iron, residual carbon, pyrite, matte, floater and so on. As shown in Fig. 2.15a and c, metallic iron is mostly round granular or arc-shaped, and some arc-shaped interconnections, with a metallic iron content of about 25–30%. As shown in Fig. 2.15b, there is a small amount of wustite (Fex O) at the edge of the ore block, and the fine-grained connection is lath-shaped, with a content of 1–2%. As shown in Fig. 2.15c, 5–7% of residual carbon particles exist in ore blocks in irregular granules. As shown in Fig. 2.15d, about 2% of pyrite and matte are distributed among metallic iron particles. Cryptocrystalline pyroxene glass and olivine are precipitated from the glass phase, and more olivine is precipitated near the floating body area. The content of olivine and pyroxene is 3–5%, and the content of glass liquid slag is 55–60%. As shown in Fig. 2.15, there are a large number of pores in the reduced sample, and the metallic iron particles are dispersed in the sample without obvious aggregation phenomenon, and a large number of glass phases encapsulate the metallic iron. Because of the dispersion distribution of iron particles, the particle size crushed before magnetic separation should be fine enough to achieve the ideal magnetic separation effect, and the magnetic separation experiment also verified this analysis. The results of magnetic separation of reduced products with different particle sizes show that the iron grade of magnetic separation materials increases with the decrease of particle size when the particle sizes are 154 µm, 74 µm and 42 µm respectively. In addition, as shown in Fig. 2.15d, copper exists in the form of pyrite and matte, which is mixed in metallic iron and will be magnetically separated together with metallic iron particles, so it is difficult to separate, and it is possible to separate only when the particle size is small enough.

20

2 Direct Reduction Iron Process

Fig. 2.15 Micrograph of reduced sample. a 1-metallic iron, 2-glassy phase, 3-cryptocrystalline pyroxene glass; b 1-metallic iron, 2-floss, 3-calcium iron olivine; c 1-metallic iron, 2-residual carbon, 3-glass phase; d 1-metallic iron, 2-pyrite, 3-matte

2.1.5 Summary (1) The alkalinity has a great influence on the experimental results. In the direct reduction roasting process, the alkalinity R = 0.6 is the best alkalinity condition, but compared with the comprehensive index of magnetic substances in the final product, the alkalinity 0.4 is more suitable. (2) The finer the particle size of magnetic separation, the higher the grade of iron after magnetic separation, but the metal recovery rate will be reduced to varying degrees. (3) Temperature has an obvious influence on the experimental results. The optimum calcination temperature is 1175 °C. After magnetic separation, TFe = 74.7% and the recovery rate is 73.90%.

2.2 Direct Reduction-Melting Experiment of Sulfuric Acid Slag in Rotary …

21

Table 2.4 Scanning electron microscope-energy spectrum analysis results No.

1#

2#

3#

4#

5#

6#

Na

0.57

Mg

1.69

1.12

2.32

1.37

Al

4.91

0.72

2.16

4.95

Si

26.53

1.86

17.11

28.71

K

1.45

0.79

1.59

Ca

14.81

9.26

14.81

Ti

0.52

1.25

C

2.12

56.71

37.52

27.29

60.36

16

FeS

Matte

100

S Mn

0.55 100

O Mineral name

8#

0.69

Cu

Fe

7#

0.49

Metallic iron MFe

4.92

70.78

32.31

1.95

44.04

24.27

36.05

45.44

Glassy phase

Floating body Fex O

CaFe(SiO4 )

Residual carbon

Glassy phase

Note 1# point 1 in Fig. 2.15a; 2# point 2 in Fig. 2.15a; 3# point 2 in Fig. 2.15b; 4# point 3 in Fig. 2.15b; 5# point 2 in Fig. 2.15c; 6# point 3 in Fig. 2.15c; 7# point 2 in Fig. 2.15d; 8# point 3 in Fig. 2.15d

(4) The change trend of Cu in magnetic materials is consistent with iron grade, and only when the crushing particle size is small enough (42 µm) can Cu and iron phase be separated obviously. (5) Mineral analysis of the best group of reduction products shows that metallic iron is dispersed in the glass phase without obvious aggregation, while copper is embedded in the metallic iron phase in the form of pyrite and matte.

2.2 Direct Reduction-Melting Experiment of Sulfuric Acid Slag in Rotary Hearth Furnace The iron grade of sulfuric acid slag can be improved by beneficiation technology, so it can be used as raw material in iron and steel industry. Therefore, using sulfuric acid slag to extract iron has a good prospect, which not only brings huge economic benefits to acid making industry and iron and steel industry, but also contributes greatly to environmental protection in China. At present, although there are many research achievements on iron extraction from sulfuric acid slag, it is difficult to put it into production on a large scale, and a lot of research and experiments are needed to use the latest technology and equipment to treat sulfuric acid slag.

22

2 Direct Reduction Iron Process

The properties of sulfuric acid slag are similar to that of iron and steel plant dust in many aspects, but because sulfuric acid slag also has its particularity, such as high sulfur content, it will have a good development prospect to choose rotary hearth furnace coal-based direct reduction method to treat sulfuric acid slag [13, 14]. However, there is almost no report on the research results of using rotary hearth furnace to treat sulfuric acid slag at present, which makes a series of conditions and influencing factors in the direct reduction process of sulfuric acid slag need to be studied in detail before the industrial test. This experiment is based on 500 kg rotary hearth furnace. According to the theoretical research and the requirement of direct reduction pellet, the influence of roasting temperature, roasting time, C/O, basicity and other factors on the treatment of sulfuric acid slag by direct reduction process was mainly investigated. Through the study of direct reduction theory, it is not difficult to find that the above factors will have great influence on the product quality of direct reduction pellets. The feasibility of iron extraction from sulfuric acid slag by direct reduction process was studied by single factor experiment, and the influence of process parameters on experimental results was studied, which can be used as theoretical basis to guide industrial production.

2.2.1 Experimental Materials and Methods There are three main raw materials used in the experiment: sulfuric acid slag, pulverized coal and calcium carbonate. High-grade pyrite is produced by high-efficiency enrichment of sulphur and iron technology developed by Institute of Process, Chinese Academy of Sciences. The iron grade of sulphuric acid slag discharged from this pyrite is more than 60%. This kind of sulfuric acid slag is used in this experiment. The analysis results of its main chemical components are shown in Table 2.5, and the particle size analysis and XRD analysis of sulfuric acid slag are shown in Figs. 2.16 and 2.17. The pulverized coal used in the experiment was analyzed by Beijing Coal Science Research Institute, and the analysis results are shown in Table 2.6. The limestone used in the experiment is white powder with a purity of more than 99.0%. Its composition is shown in Table 2.6. Table 2.5 Chemical composition analysis of sulfuric acid slag (mass fraction) Composition TFe FeO SiO2 Al2 O3 CaO MgO Na2 O K2 O P Content, %

63.6 11

3.98

0.94

0.29 0.18

0.19

S

Co

Ni

0.08 0.03 0.55 0.03 0.01

2.2 Direct Reduction-Melting Experiment of Sulfuric Acid Slag in Rotary …

23

Fig. 2.16 Particle size distribution of sulfuric acid slag

Fig. 2.17 XRD analysis of sulfuric acid slag

Table 2.6 Analysis of chemical composition of pulverized coal (mass fraction) Σ Composition Fixed C Volatile matter Ash content Analytical water Content, %

50.65

39.18

6.26

3.91

100

Total sulfur 0.48

2.2.2 Temperature and Calcination Time Parameters The calcination time determines whether the reaction can reach equilibrium. If the calcination time is too short, the reaction may not be sufficient. If the calcination time is too long, it will not only waste heat source, but also reduce the production efficiency. Moreover, the metallized pellets may be oxidized again due to the weakening of reducing atmosphere, which is not conducive to improving the quality of metallized pellets.

24

2 Direct Reduction Iron Process

Fig. 2.18 Effect of time on metallized pellets: C/O = 1.0, temperature is 1200 °C, alkalinity is 0.1

Figure 2.18 shows the curves of TFe, mFe, and mFe in metallized pellets with calcination time. The total iron content increased with the increase of time and remained unchanged after 60 min, which indicated that the reaction had reached equilibrium at this time. The content of metallic iron increased first and then decreased, and reached a stable level in 60–90 min, which was due to the incomplete reaction in the early stage, resulting in low metallization rate, while the reduction atmosphere disappeared due to the exhaustion of carbon elements in the pellets in the later stage of the reaction, resulting in the oxidation of metallic iron again. The change of total iron and metallic iron determines the change trend of metallization rate as shown in Fig. 2.18. Therefore, the roasting reduction time is more suitable at 60 min, at which time the metallization rate and iron grade are maintained at a higher water product. Carbon reduction of iron oxide is a process of capturing oxygen atoms in iron oxide. It is considered that only CO is produced in the reduction process, so the amount of reducing agent can be measured by the atomic ratio of C to O, that is, C/O. When the carbon content C/O = 1, it means that the carbon content can theoretically reduce all iron oxides, and the products are metallic iron and gas CO. The carbon content of carbon-containing pellets has a great influence on metallized pellets, which affects the total iron content, metallization rate and reaction rate. Under the conditions of reduction temperature 1200 °C, time 60 min and alkalinity 0.1, the influence of carbon content on iron content and metallization rate was investigated experimentally. The results are shown in Fig. 2.19. With the increase of carbon content, the metallization rate increases continuously, but when C/O increases after 1.2, the metallization rate remains basically unchanged, which is due to the reduction of iron elements in pellets after the increase of carbon content; With the increase of carbon content, the total iron content gradually decreases, which is due to the fact that more ash in the corresponding pulverized coal enters the pellet after the increase

2.2 Direct Reduction-Melting Experiment of Sulfuric Acid Slag in Rotary …

25

Fig. 2.19 Curve of TFE% and metallization rate versus C/O, temperature is 1200 °C, time is 36 min, alkalinity is 0.1

of carbon content, which leads to the decrease of iron grade of the pellet. Therefore, the higher the carbon content, the better. Excessive carbon content not only causes serious waste of resources, but also reduces product quality; From the change curve of metallic iron content, it can be seen that it first increases and then decreases with the increase of carbon content, Because the reaction is incomplete when the amount of carbon is low, the increase of carbon will make iron more reduced, so the content of metallic iron will increase. However, when it increases to a certain extent, the amount of metallic iron in pellets will not increase, but the total mass of pellets will still increase, so the percentage content of metallic iron will decrease as well as total iron. According to the above analysis, it can be seen that when the carbon content is about 1.2, the metallization rate basically reaches the maximum, and the metallic iron content is also the highest. Temperature also has a great influence on direct reduction. According to thermodynamic theory analysis, it can be seen that the reaction mechanism is different with different temperature. In this experiment, the influence of temperature on the related technological indexes of metallized pellets was investigated when the carbon content was 1.2 and 1.4, the calcination time was 60 min, the alkalinity was 0.1, and the temperature ranged from 1175 to 1250 °C. As shown in Figs. 2.20 and 2.21, it can be seen that the change trend of the two figures is basically the same, and the metallization rate is basically the same in each stage. The group with total iron and metallic iron percentage of 1.2 carbon content is obviously higher than that with 1.4 carbon content, which is consistent with the conclusion drawn in the process of investigating carbon content. With the change of temperature, the metallization rates of all iron and metal iron boxes first increase and then decrease with the temperature. When the

26

2 Direct Reduction Iron Process

temperature is lower, the reaction kinetics condition is insufficient, which is the main factor leading to the low metallization rate. However, there are two possible reasons for the slight decrease of metallization rate when the temperature is too high. One is that the carbon in the pellets is rapidly consumed because the temperature is too high and the reaction is very intense. Continuing heating in the furnace will change the composition of the furnace gas from reducibility to oxidation, thus oxidizing the reduced metallic iron and reducing the metallization rate; Another possibility is due to the production of low melting point compound ferroolivine (2FeO · SiO2 ). It can be seen from Fig. 2.22 that when the temperature is higher, the produced 2FeO · SiO2 will melt first, and the generated liquid phase will wrap the iron oxides participating in the reaction, which will hinder the reaction, resulting in low metallization rate. In order to verify the existence of the above two possibilities, XRD analysis and comparison were carried out on metallized pellets at different temperatures. As shown in Fig. 2.23, it can be seen that the metallic iron content is very low at 1175 °C, while the FeO content is relatively high, indicating that the reaction is very incomplete. In XRD analysis, iron olivine was found but no detectable C was found, that is, when the carbon content was 1.2 in the temperature range from 1175 to 1250 °C, the carbon element was basically exhausted, so it could not be explained that the final pellet was oxidized due to the accelerated carbon consumption due to the increase of temperature. Through the above analysis and discussion, it can be seen that the best reduction temperature is 1200 °C.

Fig. 2.20 Effect of temperature on metallized pellets C/O = 1.2

2.2 Direct Reduction-Melting Experiment of Sulfuric Acid Slag in Rotary …

27

Fig. 2.21 Effect of temperature on metallized pellets C/O = 1.4

Fig. 2.22 Phase diagram of SiO2 –FeO system

2.2.3 Basicity Parameter Control The effect of alkalinity on metallized pellets was investigated when the carbon content was 1.2, the temperature was 1200 °C, the calcination time was 60 min and the alkalinity was R = 0.1, 0.3, 0.5, 0.7 and 0.9, as shown in Fig. 2.24. With the increase of alkalinity, metallization first increased and then decreased, TFe% and MFe% first decreased and then increased, and finally decreased. This is due to the reduction reaction of iron [15, 16]. Fe2 SiO4 + 2C = 2Fe + SiO2 + 2CO

(2.1)

Fe2 SiO4 + 2C + 2CaO = 2Fe + Ca2 SiO4 + 2CO

(2.2)

28

2 Direct Reduction Iron Process

Fig. 2.23 XRD analysis of metallized pellets at different temperatures

Fig. 2.24 Effect of alkalinity on metallized pellets

It can be seen from Table 2.1 that the reaction starting temperature of formula (2.1) is much higher than that of formula (2.2). After adding limestone, the reaction (2.2) can occur, which is beneficial to iron reduction. However, the addition of limestone will increase the total weight of pellets and bring in CO2 , which is not conducive to iron reduction. The increase of total weight of pellets will reduce the grade of iron. To sum up, the optimum experimental parameters of coal-based direct reduction of sulfuric acid slag are studied in this experiment: carbon content is 1.2, roasting temperature is 1200 °C, roasting time is 60 min. Under these conditions, metallized pellets with TFe% = 73.2%, MFe% = 61.8% and metallization rate up to 91.5% can

2.2 Direct Reduction-Melting Experiment of Sulfuric Acid Slag in Rotary …

29

be obtained, which is suitable for the high temperature and high efficiency of rotary hearth furnace reduction process [17].

2.2.4 Transformation and Micromorphology of Iron Oxides During Reduction In the experiment, the XRD analysis was carried out on the samples with a group of time, and the changes of main components were observed and compared. We can see the gradual transformation process of iron oxides in carbon-containing pellets with the change of reduction time. As shown in Fig. 2.25, the main component of the initial sulfuric acid slag is Fe2 O3 . After 3 min reduction, Fe2 O3 almost disappears, resulting in more Fe3 O4 and some FeO. After 30 min reduction, there is still a small amount of FeO, and metallic iron appears, and the component is basically stable after 9 min. By analyzing and comparing the XRD results, it is verified that the transformation process of iron oxide in the whole reduction process is Fe2 O3 → Fe3 O4 → FeO → Fe. The anti-oxidation ability of metallized pellets is closely related to its microstructure, so its internal structure plays a key role in the anti-oxidation ability of metallized pellets. Figure 2.26 shows the sample preparation of a group of metallized pellets with carbon content C/O = 1.2, calcination temperature 1200 °C and calcination time 60 min, and the SEM observation and energy spectrum analysis at three selected points. Through the results of energy spectrum analysis, it can be seen that the white part in the picture is metallic iron; The gray part contains a lot of silicon, oxygen,

Fig. 2.25 Comparison of XRD analysis results of reduction products at different times

30

2 Direct Reduction Iron Process

Fig. 2.26 Electron microscopic observation and energy spectrum analysis of metallized pellets

aluminum and iron, etc. The components should be SiO2 , Al2 O3 and FeO, which are slag phase; The black part contains a large amount of element C, which is the residual carbon that is not completely reacted. It can be seen that the metallic iron has appeared continuous crystal phenomenon, the slag phase is dispersed between the metallic iron phases, and the carbon content is concentrated and the amount is relatively small. The appearance of metallic iron is helpful to the densification of metallized pellets and has a positive effect on preventing secondary oxidation. However, the ash in coal will hinder the polymerization of metallic iron, so the ash content of pulverized coal should be as low as possible. However, increasing the temperature or prolonging the calcination time is beneficial to the consumption of carbon and the crystallization of metallic iron.

2.2.5 Reduction Melting Theory and Experiment of Carbon-Containing Pellets of Sulfuric Acid Slag In the rotary hearth furnace process, the slag and iron are partially melted when the carbon-containing pellets are at 1300–1400 °C, and the preliminary separation

2.2 Direct Reduction-Melting Experiment of Sulfuric Acid Slag in Rotary …

31

Fig22..2.27 Flow chart of reduction melting experiment

begins [18, 19]. According to the theoretical analysis of melting reaction, it can be seen that the influence of alkalinity on melting reaction will be much greater than that of direct reduction except carbon content, temperature and time. Because it involves higher temperature, partial melting of slag and iron will occur, and the melting of slag phase plays an important role in melting reaction. After the rotary hearth furnace test, the pellets were collected for laboratory melting study. So as to ensure that the metal yield can be achieved. In the melting separation experiment, the influence of various factors on it is studied by single factor variable experiment. The experimental variables include carbon content, temperature, time and alkalinity. Separable slag and pearl iron can be obtained by the experiment. The research objects are the behavior characteristics of sulfur in slag and iron, iron content in slag and metal recovery rate, etc. The melting experiment of dried carbon-containing pellets was carried out in a fast heating electric furnace. The experimental steps are as follows (Fig. 2.27): (1) Raise that temperature of the 1600 °C fast heating electric furnace to the require temperature; (2) Weighing the pellets and putting them into a graphite crucible, putting the graphite crucible containing the pellets into a furnace chamber after the temperature of the electric furnace is constant, and closing the furnace door; (3) When the pellet is first put into the furnace, the temperature in the furnace will drop rapidly, when the temperature reaches stability again, timing will be started, the crucible will be taken out after reaching the set time, and the air cooling will be covered with other idle graphite crucibles in time; (4) After the sample is cooled, separating the slag and iron, respectively loading them into sample bags and numbering them; (5) Carrying out the next group of experiments according to the above experimental steps; (6) After the experiment is completed, the obtained slag and iron are tested and analyzed according to the analysis requirements.

32

2 Direct Reduction Iron Process

2.2.6 Melting Test Results When the temperature is 1450 °C, the carbon content is 1.2 and the melting time is 20 min, the iron recovery rate changes with alkalinity as shown in Fig. 2.28. It can be seen from the ternary phase diagram of CaO–SiO2 –Al2 O3 system that the melting point of slag system first decreases and then increases with the increase of basicity. Comparing the slag phase XRD analysis of the two groups with basicity of 1.2 and 0.9, it can be seen from Fig. 2.29 that the group with high basicity has more calcite feldspar and less anorthite. The low melting point of slag is beneficial to the separation of slag and iron in melting stage, while the high melting point of slag leads to the poor separation of slag and iron, which leads to a lot of iron being mixed into slag phase and reduces the recovery rate of iron. To sum up, when the alkalinity is 1.2, the iron recovery rate can reach the maximum. Fig. 2.28 Variation of iron recovery rate with alkalinity

Fig. 2.29 XRD analysis of slag phase with different basicity

2.2 Direct Reduction-Melting Experiment of Sulfuric Acid Slag in Rotary …

33

When the temperature is 1450 °C, the alkalinity is 1.2 and the melting time is 20 min, the curve of iron recovery rate with carbon content is shown in Fig. 2.30. When the carbon content is low, the reduction process is insufficient, which leads to less pearl iron collected and lower iron recovery rate. When the carbon content is high, more ash is brought in by pulverized coal, and the ash in pulverized coal hinders the polymerization of metallic iron in the melting process, which leads to fine particles of metallic iron, difficult separation of slag and iron, and low iron recovery rate. As shown in Fig. 2.31, a group of slag with high carbon content has higher iron content. Therefore, the carbon content of 1.2 is more appropriate. When the alkalinity is 1.2, the carbon content is 1.2 and the melting time is 20 min, the iron recovery rate changes with temperature as shown in Fig. 2.32. The increase of temperature is beneficial to the increase of iron recovery in both direct reduction process and melting process. Because of the higher temperature, the direct reduction Fig. 2.30 Variation of iron recovery rate with carbon content

Fig. 2.31 XRD analysis of slag with different carbon content

34

2 Direct Reduction Iron Process

process is accelerated and the melting process time is increased, which is beneficial to the separation of slag and iron; With the increase of temperature, the fluidity of slag and iron is enhanced, and the slag and iron can be separated well, so that the iron recovery rate is improved [20, 21]. As shown in Fig. 2.33, the content of metallic iron in the slag group with low temperature is slightly more than that in the slag group with high temperature, and the content of iron olivine which is difficult to react is also more in the slag group with low temperature. Increasing the temperature is beneficial to the reduction of iron in iron olivine. Therefore, under these conditions, the melting temperature is more suitable at 1450 °C. When the alkalinity is 1.2, the carbon content is 1.2 and the temperature is 1450 °C, the iron recovery rate changes with time as shown in Fig. 2.34. It can be seen that the metallization rate gradually increases with the extension of melting time, and finally tends to be stable. When the melting time is too short, the reduction reaction and melting process are not good. With the extension of time, the reduction is fully Fig. 2.32 Variation of iron recovery rate with temperature

Fig. 2.33 XRD analysis of slag phase at different temperatures

2.2 Direct Reduction-Melting Experiment of Sulfuric Acid Slag in Rotary …

35

carried out, and the elongation of melting process can make slag and iron better separated [22, 23]. As shown in Fig. 2.35, comparing the XRD analysis of slag under different melting time, it can be seen that the content of a group of ferropyroxene (FeSiO3 ) and ferroolivine (Fe2 SiO4 ) with short time is very high, which indicates that these difficult reduction reactions are not fully carried out; The content of iron oxide is similar, but the content of metal iron in the group with short melting time is higher, which shows that there is more iron in the slag with short melting time. Therefore, under these conditions, the melting reaction time is reasonable at 20 min. To sum up, the optimum technological parameters of melting experiment are basicity 1.2, carbon content 1.2, temperature 1450 °C and melting time 20 min. Under these conditions, the recovery rate of iron can reach 92.6%. Fig. 2.34 Variation of iron recovery rate with time

Fig. 2.35 XRD analysis of slag under different melting time

36

2 Direct Reduction Iron Process

2.2.7 Behavior Analysis of Sulfur Element in Melting Process Sulfuric acid slag is the slag after pyrite roasting to produce sulfuric acid, which contains a lot of sulfur in the form of ferrous sulfide. Because there is no separation of slag and iron in the direct reduction process, the removal of sulfur can only rely on gasification desulfurization. In the melting reaction, sulfur can not only be gasified and desulfurized in the direct reduction stage, but also rely on slag for desulfurization. The alkalinity of slag has great influence on the distribution of sulfur in slag and iron. When the carbon content is 1.2, the temperature is 1450 °C and the melting time is 20 min, the sulfur content in slag and iron changes with alkalinity as shown in Fig. 2.36. It can be seen from the figure that the desulfurization rate increases and then decreases with the increase of alkalinity. When the alkalinity increases, there is enough CaO in the pellets to carry out desulfurization reaction [24]: FeS + CaO + C = Fe + CaS + CO Therefore, the sulfur content in the slag will increase and the desulfurization rate will be improved. When the slag alkali is too high, although the desulfurization reaction is sufficient, the slag amount increases, which is not conducive to iron aggregation and easy to make slag inclusion in pearl iron. However, the melting process is in a semi-melting state, and the diffusion conditions are extremely poor. The product CaS is mainly concentrated at the interface between slag and iron, and the sulfur content of pearl iron will increase after being sandwiched with slag with higher sulfur content.

Fig. 2.36 Variation of S content and desulfurization rate in slag and iron with alkalinity

2.2 Direct Reduction-Melting Experiment of Sulfuric Acid Slag in Rotary …

37

2.2.8 Summary In this chapter, according to the rotary hearth furnace ironmaking process, the resources of sulfuric acid slag are recycled and the iron is extracted. In this chapter, the direct reduction of carbon-containing pellets from sulfuric acid slag to produce metal pellets and reduction melting to produce pearl iron were studied, and the optimum process parameters of direct reduction and reduction melting under corresponding conditions were found out. The following conclusions are drawn: (1) In the direct reduction process, the optimum process parameters for metallized pellets are: carbon content 1.2, calcination temperature 1200 °C, calcination time 60 min. Under these conditions, metallized pellets with TFe% = 73.2%, MFe% = 61.8% and metallization rate up to 91.5% can be obtained. (2) When the carbon content is 1.2 and the calcination temperature is 1200 °C, the reduction process of iron oxide in the direct reduction process is determined to be Fe2 O3 → Fe3 O4 → FeO → Fe by XRD analysis of the composition of metallized pellets in different time periods. (3) The microstructure of metallized pellets was observed by scanning electron microscope and energy spectrum analysis of the metallized pellets produced under the optimum technological parameters. The metallic iron was in the state of continuous crystal, with slag phase and a small amount of C dispersed in it. (4) The most favorable technological parameters of melting experiment for metal recovery are alkalinity 1.2, temperature 1450 °C, carbon content 1.2 and melting time 20 min. Under these conditions, the metal recovery can reach 92.6%. (5) The desulfurization ability is poor in the production of metallized pellets by direct reduction experiment of sulfuric acid slag and pearl iron by melting experiment, and the sulfur content of the products is about 1%, so the products are suitable as raw materials for producing sulfur-based free-cutting steel.

Chapter 3

Digestion of Solid Waste During Sintering-Pelletizing Process

Abstract Municipal waste incineration fly ash (referred to as fly ash) is a significant solid by-product that requires safe treatment and resource utilization. We propose using fly ash instead of lime as part of the flux in the sinter process, owing to its alkaline composition. Through thermodynamic calculations and experimental verification, we found that CaO from fly ash can form complex calcium ferrite with Fe2 O3 , which can effectively replace lime to form the binder phase. The Al2 O3 impurities present in fly ash promote the formation of acicular calcium ferrite, which improves the reducibility and mechanical strength of the sintered ore. However, an Al2 O3 content greater than 2% can deteriorate the metallurgical properties of the sintered ore. Alkaline metal chlorides in fly ash can improve the dissolution of Al2 O3 and SiO2 in silicate, but most are volatilized, with little effect on the binder phase. The best metallurgical properties were obtained when fly ash replaced 25% lime as the sintering alkaline flux. Replacing lime with fly ash during sintering has promising application prospects. Keywords Municipal waste incineration fly ash · Sintering co-processing · Composite calcium ferrite · Thermodynamic calculation

3.1 Return Sintering of Steel Slag as Raw Material 3.1.1 Research Status of Steel Slag Utilization Meng Fei, Liu Qingcai et al. [25] used the tailings after magnetic separation of steel slag to replace some sintering solvent to carry out sintering pot experiment, and studied the metallurgical properties of sinter under different tailings blending ratios. The study showed that calcium ferrite phase, tumbler strength, reducibility and finished product rate increased with the increase of tailings blending ratio, and when the tailings addition ratio was 5%, they reached the maximum. Liu Shouping, Wen Guangyuan et al. [26] carried out sintering experiments with converter steel slag with a particle size of 2–6 mm after process quenched by air.

© Metallurgical Industry Press 2023 K. Dong et al., Resource Utilization of Solid Waste by Thermometallurgy in Steel Processes, https://doi.org/10.1007/978-981-99-5655-5_3

39

40

3 Digestion of Solid Waste During Sintering-Pelletizing Process

The experimental results showed that addition of steel slag improved the air permeability of sintered material layer, thus improving the production efficiency of sintering machine. However, with the increase of steel slag blending ratio, the reduction degree of sinter decreases. Li Shuzhuang [27] obtained the following conclusion through converter steel slag sintering experiment: When the proportion of steel slag is 5.6%, the metallurgical properties of sinter are the best; when the granularity of sinter is 0–4 mm, it is beneficial to improve the crystallization and melting conditions of sinter, and the strength of sinter is the best at this time; because steel slag is clinker, with the increase of sinter ratio, the fuel consumption in sintering process is smaller. Wu He and Yue Long studied the basic sintering characteristics of gas quenching steel slag and other iron-containing raw materials through TSJ-3 infrared rapid high-temperature micro-sintering system. Through exploration, the conclusion is as follows: gas quenching steel slag can produce high-strength bonding phase at high temperature, with good fluidity and reasonable assimilation temperature; steel slag contains a large number of oxides similar to quicklime, which makes it possible to be used as sintering solvent. However, the iron grade is low, and if it is added as sintering solvent, the sinter grade will be reduced [28]. Shen Delin [29] carried out sintering experiments with steel slag of Baotou Steel. Combined with various detection methods, the following conclusion is obtained: P existing in steel slag will react with C2 S in solid solution during sintering to produce β-C2 S with stable physicochemical properties, thus improving the pulverization resistance of sinter; when the proportion of steel slag is 4%, 36.5 kg quicklime can be saved for every 1 t sinter produced. Yu Mingxing and Sun Ying [30] summed up through production practice that the grain size of steel slag should be less than 10 mm, the content of acidic oxide should not be too high and the chemical composition should be stable. Anyang Iron and Steel Group Co., Ltd. further expanded the recycling utilization scale of steel slag by improving the production process.

3.1.2 Properties of Raw Materials The neutralization powder used in this experiment was taken from a steel plant in Tangshan, Hebei Province, and detected by X-ray fluorescence spectroscopy (XRF). The results shown in Table 3.1 are obtained. Based on the characteristics of different mineral powders, combined with granulation performance, mineral phase composition, alkalinity, different particle content and other factors, the neutralization powder components were prepared to achieve the best sintering performance (Table 3.2). The steel slag used in this experiment is the converter steel slag produced in the converter workshop of this steel plant, and the chemical composition of the steel slag is detected, shown in Table 3.3.

3.1 Return Sintering of Steel Slag as Raw Material

41

Table 3.1 Analysis on chemical composition of neutralization powder (mass fraction) Element

TFe

MgO

Al2 O3

S

Zn

Hamersley

61.32

4.23

0.03

0.13

2.42

0.41

0.02

0.02

5.08

Yandi Fines

57.63

6.23

0.05

0.06

1.59

0.46

0.01

0.01

9.60

BA1

65.33

1.58

0.11

0.10

1.30

0.77

0.01

0.01

0.85

BA2

60.98

10.65

0.40

0.26

0.86

6.20

0.01

0.01

2.07

CArajas

65.18

7.64

0.15

0.06

0.15

27.25

0.04

0.01

0.23

0.00

4.54

70.05

3.94

0.43

0.23

0.03

0.00

20.82

Lime

SiO2

CaO

FeO

Other

Table 3.2 XRF composition analysis of iron concentrate powder (mass fraction) Composition

TFe

Fe2 O3

FeO

SiO2

CaO

MgO

Al2 O3

P

S

Other

Content, %

70.19

60.50

10.05

3.98

6.35

2.55

5.34

0.05

0.01

1.24

Table 3.3 XRF composition analysis of steel slag (mass fraction) Composition

TFe

Fe2 O3

FeO

SiO2

CaO

MgO

Al2 O3

P

S

Other

Content, %

30.69

15.65

16.70

16.40

36.60

3.70

2.27

0.46

0.45

0.85

3.1.3 Sintering Pot Experiment The main raw materials used in this sintering experiment are neutralization powder, return mine, steel slag, quicklime, limestone, etc. According to the characteristics of sintering experiment, it is necessary to weigh the raw materials according to a certain proportion and mix them evenly before adding them into the sintering pot. According to the general sinter properties in the sintering workshop, the sintering alkalinity (R = 2) is selected, and the composition of sintering mixture shown in Table 3.4 can be obtained, in which the variable is the blending ratio of steel slag. After weighing the sintering raw materials, they are mixed on the mixing table. After manual mixing, they are loaded into the granulator for second mixing, and Table 3.4 Composition content of sintering raw materials with different proportions (mass fraction) Composition (%)

TFe

Fe2 O3

0

42.04

14.66

1

54.06

29.84

2

52.85

41.67

3

51.60

4

50.55

5

49.31

FeO

SiO2

CaO

MgO

Al2 O3

P

S

Other

4.17

8.34

2.65

3.62

0.45

0.21

42.04

9.86

14.11

4.16

8.32

2.59

3.56

0.44

0.20

11.07

14.11

4.16

8.32

2.59

3.56

0.44

0.20

13.02

40.68

13.87

4.14

8.28

2.42

3.44

0.42

0.18

14.82

39.66

13.62

4.12

8.24

2.40

3.37

0.40

0.16

16.70

38.85

13.35

4.11

8.22

2.36

3.33

0.37

0.14

18.16

42

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Fig. 3.1 Experimental device

then the sintering mixture can be loaded into the sintering pot for ignition. The experimental device is shown in Fig. 3.1.

3.1.4 Experimental Results 3.1.4.1

The Effect of Steel Slag Dosage on Sintering Process

As can be seen from Fig. 3.2, with the increasing amount of steel slag, the sintering temperature of the sintering pot rises continuously. The main reason is that converter steel slag belongs to clinker itself, which is easy to melt and low in energy consumption during sintering, and steel slag contains high CaO. When the alkalinity remains constant, the amount of limestone (CaCO3 ) is reduced correspondingly, and the energy consumption of CaCO3 decomposition in sintering process is reduced correspondingly, resulting in the increase of sintering temperature. It can be seen from Fig. 3.3 that with the increasing of steel slag dosage, the vertical sintering rate of sintering pot increases continuously. Steel slag is a particle with high strength, and the particle size of sintered material is improved after being mixed with sintered material. Steel slag is the core of raw material pelletizing in the mixing process, which improves the pelletizing effect of sintered material and increases the air permeability of sintered material, thus shortening sintering time and increasing vertical sintering speed. It can be seen from the Fig. 3.4 that when a small amount of steel slag is added to the sintering raw material (1–2%), the sinter product ratio is improved compared with that without steel slag. However, when the amount of steel slag increases gradually, the sinter product ratio begins to decline. It is found that the content of CaO in converter steel slag is extremely high, generally above 30%, and the content of FeO is generally above 10%. CaO appears in the mineral form of dicalcium silicate and tricalcium silicate, which is easy to react with iron minerals to form calcium-iron olivine with low melting point, which

3.1 Return Sintering of Steel Slag as Raw Material

43

Fig. 3.2 The effect of steel slag on sintering temperature

Fig. 3.3 The effect of steel slag on vertical sintering rate

leads to the appearance of liquid phase in advance and the increase of liquid phase in sintering process. The quantity and quality of liquid phase are the decisive factors of sinter, so the increase of liquid phase is an important reason for the improvement of sinter quality. The increase of liquid phase and the good crystallization degree of iron-bearing phase and cementing phase change the structure of ore phase and further improve the tumbler strength of sinter. When the steel slag dosage is 3%, the yield and tumbler strength are close to the reference period. With the increase of steel slag addition, both of them decrease gradually, which results from the fact that the amount of steel slag is too large, the sintering temperature increases sharply, the holding time is too short, and the phenomenon of local over incineration is formed. The liquid phase is difficult to reach equilibrium in a short time, which leads to the incomplete melting and crystallization of most sintering materials, and the strength

44

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Fig. 3.4 The effect of steel slag addition on sinter product ratio

Fig. 3.5 The effect of steel slag addition on strength of sintering drum

and yield of sinter gradually decline (Fig. 3.5). Therefore, too large steel slag ratio has adverse effects on the strength and product ratio of sinter.

3.1.4.2

The Effect of Steel Slag Dosage on Metallurgical Properties of Sinter

The metallurgical properties of sinter with particle size larger than 25 mm are tested, and the test results are shown in Table 3.5. Among them, No. 1–6 indicate that the steel slag dosage is 0%, 1%, 2%, 3%, 4% and 5%, respectively. It can be seen from Table 3.5 that the reduction rate of sinter shows a downward trend with the increase of steel slag ratio. Compared with the reference period, when

3.2 Application of Waste Incineration Fly Ash in Sintering Process

45

Table 3.5 The metallurgical properties of sinter with particle size larger than 25 mm Experiment No.

Reduction rate, %

Pulverization index RDI+6.3

RDI+3.15

RDI−0.5

1

89.5

83.25

92.63

2.87

2

89.1

83.87

92.87

2.31

3

89.2

84.05

93.32

1.21

4

87.6

83.11

91.36

1.87

5

86.2

82.33

90.76

2.42

6

85.4

81.64

90.41

2.67

the steel slag ratio is 5%, the reduction rate decreases from 89.5 to 85.4%. There are two reasons: on the one hand, addition of steel slag will increase the liquid phase in the sintering process, which will result in the compact fiber structure and the decrease of porosity of sinter, thus reducing the reduction rate of sinter; At the same time, because the ratio of coke powder is unchanged, the reducing atmosphere is enhanced after adding steel slag, the sintering temperature is increased, there are more floating bodies in sinter, and the grains are fully grown, resulting in the reduction degree of sinter being reduced. On the other hand, when the ratio of steel slag increases, the iron grade of sinter will decrease, which will lead to the decline of reduction performance. It can be seen from Table 3.5 that the pulverization performance of sinter has been obviously improved. When the ratio of RDI+6.3 , RDI+3.15 and RDI−0.5 steel slag is 2%, the pulverization performance is improved most, and the pulverization index at low temperature is increased by 0.69% compared with the reference period. The reasons for the improvement of sinter pulverization performance are as follows: on the one hand, the liquid phase in sinter increases, the strength increases, and the amount of Fe2 O3 decreases, which inhibits the volume expansion caused by the lattice transformation of Fe2 O3 + Fe3 O4 during cooling, and reduces the low temperature pulverization rate of sinter; on the other hand, steel slag contains a certain amount of MgO, which improves the bonding performance and liquid crystal state of sinter, and it is beneficial to reduce the pulverization rate of sinter.

3.2 Application of Waste Incineration Fly Ash in Sintering Process Incineration fly ash (hereinafter referred to as fly ash) generated by waste incineration refers to the residues collected in flue gas purification system (APC) and heat recovery and utilization system, which are generally gray or dark gray, with particle size usually less than 300 μm and diversified particle morphology [31]. The main elements in fly ash are Si, Ca and Al. In addition, it also contains a small amount

46

3 Digestion of Solid Waste During Sintering-Pelletizing Process

of K, Na, Cl, Fe, Ti, Mg and trace elements such as Pb, Cr, Cd, Zn, Hg, Cu, Ni, As and others, as well as certain concentrations of dioxin, furan, polycyclic aromatic hydrocarbons and other organic pollutants [32]. Dioxins and furans are classified as extremely toxic substances. Because of its great adverse effects on human and animal health [33, 34]. Therefore, fly ash is also classified as hazardous waste disposal by the Technical Policy for Prevention and Control of Hazardous Waste Pollution. Because of different raw materials and incineration methods, the composition and properties of fly ash are quite different, and the color is mainly from gray to black. There are many kinds of fly ash particles, such as rod, polyhorny, cotton wool and spherical irregular polymer shapes [35]. Dioxin decomposes at 705 °C, and can be completely decomposed at 2 s at high temperature (higher than 800 °C). Dioxin can be synthesized again at 250–500 °C. The sinter bed can be divided into sinter bed, combustion bed, preheating bed, drying bed and over wetting bed according to temperature change and physical and chemical reaction. Figure 3.6 shows the temperature distribution of sintered material layer. As can be seen from Fig. 3.6, the combustion zone is the area with the highest temperature in the sintered material layer, and the temperature can reach above 1300 °C. Under this condition, dioxin in fly ash can be completely decomposed; The temperature of sintering drying preheating zone is 100–700 °C, which can provide conditions for the re-synthesis of dioxin. Generally speaking, the content of dioxin in flue gas after sintering with fly ash is uncertain, but the following aspects can be considered to control the dioxin in sintering flue gas. (1) Add inhibitors. Inhibitors such as ammonia, urea and carbohydrazide can remove chlorine and destroy metal catalysts, thus inhibiting the formation of dioxin. The main principle is that these inhibitors will form ammonia compounds during heating, which will reduce the concentration of HCl and inhibit the catalysis of copper and other metals, and hinder the synthesis of dioxin at medium and low temperatures, thus reducing the formation of dioxin. (2) Optimize process parameters. By controlling sintering process parameters, such as sintering atmosphere and Fig. 3.6 Temperature distribution of sinter bed

3.2 Application of Waste Incineration Fly Ash in Sintering Process

47

sintering time, the amount of dioxin generated can be reduced, such as controlling the rapid cooling of bellows flue gas at the back of sintering and accurately controlling the sintering endpoint at the tail of sintering machine. Oxygen is one of the essential factors of dioxin synthesis. Choosing suitable oxygen content in sintering process can not only ensure the combustion effect of fuel, but also reduce the generation of dioxin. (3) Sintering flue gas circulation. Studies shows that some or all of the sintering flue gas is circulated into the sintering material layer, which can not only reduce the consumption of solid fuel and the emission of pollutants from sintering flue gas, but also make dioxin in flue gas decomposed at high temperature on the combustion belt, thus reducing the emission of dioxin. A foreign sintering machine adopts this technology, and the emission reduction of dioxin reaches 70%, and the emission reduction of particulate matter and nitrogen oxides is nearly 45%.

3.2.1 Physicochemical Parameters of Fly Ash (1) Chemical composition of waste incineration fly ash are shown in Table 3.6. (2) Main phase and particle size distribution are shown in Fig. 3.7. The main particle size of fly ash is between 5–10 and 30–50 μm. In the granulation process, this particle size will be bonded and pulverized. The uniformity and granularity of granulation are the key points worthy of attention. Table 3.6 Composition of waste fly ash (wt%) Element

Ca

Cl

Na

K

S

Si

Fe

Mg

Al

Zn

Content

34.13

33.63

12.86

8.40

3.89

2.23

1.08

0.95

0.79

0.77

Element

Ti

P

Pb

Br

Cu

Sr

Mn

Cr

Sn

Cd

Content

0.44

0.35

0.14

0.10

0.07

0.05

0.04

0.03

0.03

0.02

Fig. 3.7 Fly ash particle size distribution

48

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Table 3.7 Composition of fly ash after incineration (wt%) Element

CaO

Cl

Na

K

S

SiO2

Fe2 O3

MgO

Al2 O3

ZnO

Content

71.33

5.50

3.80

1.99

2.57

4.66

2.26

1.99

1.65

1.61

Element

TiO2

P2 O5

PbO

Br

Cu2 O

Sr

MnO

Cr

Sn

Cd

Content

0.920

0.731

0.293

0.209

0.146

0.104

0.084

0.063

0.063

0.042

(3) Composition of fly ash after incineration are shown in Table 3.7. Calcium oxide is the main component of fly ash after incineration. On the premise of not affecting the quality of sinter, fly ash can replace CaO in a large amount during sintering. From the incineration loss results, a large amount of alkali metal volatilization will dissolve in the wet layer and dry layer, which may lead to the blockage of the material layer and affect the smooth sintering process. Volatilization of a large amount of alkali metals in the sintering process will also have a little negative impact on the combustion of coke, reducing the combustion capacity of coke. From the analysis of incineration loss of raw materials, fly ash will have a certain degree of negative impact on sintering process.

3.2.2 Sintering Experimental Procedure Homogenizing ore, sintering powder, flux (limestone, dolomite, quicklime), coke powder and returned ore are weighed and added one by one according to this sequence. The total amount of dry materials in each cup is about 30 kg, and each charge is between 26 and 27 kg. Then, they are mixed evenly on the artificial iron sheet, and then a proper amount of water is added after mixing evenly. After the first mixing, the mixed materials are put into a cylinder mixer for mixing. If the moisture is insufficient, the second mixing can be properly added with water spray to keep the moisture at 6.5% ± 0.5% (considering the large particle size of our uniform ore, the moisture should not be too high), and the second mixing time is 3 min. Subsequently, the mixed raw materials are added into a sintering cup with a thickness of about 470 mm (excluding the height of the bedding material). Before adding, about 3 kg of sinter is laid at the bottom of the sintering cup as the bedding material. After loading, the exhaust fan is controlled to make the ignition negative pressure 8 kPa, the air valve and natural gas valve are adjusted, the ignition time is 1.5 min, the ignition temperature is 1050 °C, and the sintering negative pressure is controlled to 10 kPa after ignition. Sintering parameters are shown in Table 3.8. After sintering, the economic and technical indexes of sintering process and the strength, chemical composition and metallurgical properties of sinter were measured, respectively. After the sintering experiment, 10, 20 and 10 g sinter samples are taken from the upper, middle and lower parts of the sinter column after the sinter is cooled respectively, and crushed in an electromagnetic crusher for 1 min. After screening through a 140-mesh sieve, appropriate samples were taken and sent to the Chemical Analysis

3.2 Application of Waste Incineration Fly Ash in Sintering Process

49

Table 3.8 Experimental parameter control of sintering pot Item

Material bed, mm

Ignition temperature, °C

Ignition time, min

Ignition negative pressure, kPa

Sintering negative pressure, kPa

Moisture of mixture, %

Mixing time, min

Parameter

500

1050

1.5

8

10

6.5 ± 0.5

3

Center of University of Science and Technology Beijing for total iron and ferrous oxide content analysis (sample mass is greater than 1 g), XRD phase analysis (high temperature X-ray diffractometer (model SmartLab9)), and XRF element analysis (X-ray fluorescence spectrometer (model Axios max)). After artificial crushing, all sinters were tested for falling strength. After the test was completed, 1.6 kg sinter with particle size greater than 25 mm and 1.4 kg sinter with particle size greater than 10 mm and less than 25 mm were screened for drum index test. The reduction degree refers to the national standard of China, the reduction determination method of iron ore. Total iron content A, ferrous iron content B, sample weight M 0 (500 g), sample weight before reduction M 1 , experimental weight M 2 after 180 min, particle size 10–12.5 mm, reduction atmosphere 30% CO + 70% N2 , reduction volume flow rate 15 L/min. The reduction degree is calculated as follows:  RI =

 M1 − M2 0.11B + × 100 × 100% 0.43A 0.43A × M0

(3.1)

The equipment and process parameters used in the test of low temperature reduction pulverization method in China National Standard (GB/T 13242—91) are basically the same as those used in the test of iron ore reducibility. The reduction temperature is 500 °C ± 10 °C, the reduction time is 60 min, the reduction gas composition is CO 20%, CO2 20%, N2 60%; The concentration of H2 is less than 0.2% or 2.0%. After reduction, the samples were cooled by N2 , then all the samples were loaded into a small drum (φ130 mm × 200 mm), rotated at a speed of 30 r/min for 10 min, and the samples after the drum were screened. The reduction pulverization index was taken as the percentage of the ratio of the mass of + 6.3, + 3.15 and − 0.5 mm particles to the total mass of the samples after reduction. RDI+3.15 was used as the assessment index, while RDI+6.3 and RDI−0.5 were only used as reference indexes.

3.2.3 Proportioning Design of Fly Ash Addition in Sinter In the experiment, single garbage fly ash was added. Among them, it is added according to two formulations, such as 2.5, 5, 7.5 and 10%. Table 3.9 shows the experimental group numbers.

50

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Table 3.9 Experimental group number Experimental group

1

2

3

4

5

6

7

Dust and sludge, %

2.5

2.5

2.5

2.5

2.5

5

5

Experimental group

8

9

10

11

12

13

14

Dust and sludge, %

5

5

5

7.5

7.5

7.5

7.5

Experimental group

15

16

17

18

19

20

Dust and sludge, %

7.5

10

10

10

10

10

Mixed addition considers the influence of different components, and fly ash is mainly used to replace part of quicklime and limestone in sintering raw materials. Fe, Si and Al in the raw material composition of fly ash will also have a certain degree of influence. Tables 3.10, 3.11 and 3.12 show the batching with different fly ash proportions and preset sinter composition. We simplified the composition system after batching, and regarded alkaline oxides as CaO and acidic oxides as SiO2 . This is a ternary phase diagram composed of Fe2 O3 –CaO–SiO2 , as shown in Fig. 3.8. The mass fractions of Fe2 O3 , SiO2 and CaO were 80%, 4.5–5% and 8–10% respectively. By adding some fly ash instead of quicklime, the content of total iron SiO2 will increase, and the composition point will shift away from Fe2 O3 to the apex of SiO2 . If the addition amount is very low, Table 3.10 Composition and proportion of neutralization powder (wt%) No.

Name

Neutralization powder proportion

TFe

CaO

MgO

Al2 O3

1#

Hamersley

30

61.32

2#

Yandi Fines 18.75

57.63

4.23

0.03

0.13

2.42

0.41

6.23

0.05

0.06

1.59

0.46

3#

Fine powder

0

66.53

6.34

0.03

0.2

0.28

4#

BA1

37.5

65.33

1.58

0.11

0.1

1.3

5# 6#

BA2

1.25

60.98

10.65

0.4

0.26

CArajas

12.5

65.18

7.64

0.15

0.064

No.

Name

Ig

S

P

As

Zn

Pb

K2 O

Na2 O

1#

Hamersley

5.08

0.024

0.106

0

0.015

0

0.028

0.027

2#

Yandi Fines

9.6

0.013

0.045

0

0.01

0

0.023

0.008

3#

Fine powder

0.85

0.035

0.013

0

0.005

0

0.017

0.012

4#

BA1

2.07

0.005

0.078

0

0.004

0

0.023

0.011

5#

BA2

0.23

0.013

0.084

0

0.005

0.052

0.03

6#

CArajas

0

0.038

0.032

0

0.009

0.025

0.011

Note Ig—Ignition loss

SiO2

FeO

MnO

22.1 0.77

0

0.86

6.2

0

0.15

27.25

0

0

3.2 Application of Waste Incineration Fly Ash in Sintering Process

51

Table 3.11 Fly ash addition ratio (wt%) Component

Proportioning

Incineration loss

Component

Proportioning

Incineration loss

Neutralization powder

63.2

4.103125

Neutralization powder

63.2

4.103125

Return ore

20

0

Return ore

20

0

Quicklime

7.4

20.81

Quicklime

5.9

20.81

Limestone

0.7

44.08

Limestone

0.7

44.08

Magnesium powder

1.7

49.17

Magnesium powder

1.7

49.17

Coke powder

4.5

87.12

Coke powder

4.5

87.12

Fly ash

2.5

45.13

Fly ash

5

45.13

Neutralization powder

63.2

4.103125

Neutralization powder

63.2

4.103125

Return ore

20

0

Return ore

20

0

Quicklime

3.4

20.81

Quicklime

1.1

20.81

Limestone

0.7

44.08

Limestone

0.7

44.08

Magnesium powder

1.7

49.17

Magnesium powder

1.7

49.17

Coke powder

4.5

87.12

Coke powder

4.5

87.12

Fly ash

7.5

45.13

Fly ash

10

45.13

Table 3.12 Preset sinter composition (wt%) Fly ash ratio

R

MgO

TFe

Al2 O3

SiO2

CaO

2.5

2.160

2.006399

48.0

1.681797

4.512561

9.748673

5

2.043

1.958179

47.5

1.737667

4.630951

9.459536

7.5

2.070

1.952031

47.1

1.756731

4.676081

9.67876

10

2.086

1.942625

46.7

1.778646

5.02

8.6344

there are not many sintered materials with offset of composition points, and they all exist in spinel phase. It can’t cause phase change. In terms of composition, it is completely feasible to add fly ash. However, the influence on sinter quality index needs experimental demonstration.

52

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Fig. 3.8 Fe2 O3 –CaO–SiO2 ternary phase diagram

3.2.4 Sintering Results 3.2.4.1

The Effect of Fly Ash Addition on Blending Granulation

In the process of sintering granulation, a large amount of fly ash has caused some troubles to granulation. Because of the low particle size of fly ash itself, the phenomenon of uneven granulation is caused in the granulation process. When the amount of fly ash is 2.5%, there is almost no difference between the granulation group and the reference group without fly ash. Granulation is more uniform. When the amount of fly ash exceeds 5%, the granulation begins to appear moderate block adhesion in different degrees. This is also acceptable for sintering process. When the proportion of fly ash reaches 10%, a large number of block adhesion and unformed particles begin to appear (Figs. 3.9 and 3.10). The granulation distribution is very uneven, and a large number of particles within 1 mm are formed, which is fatal to the sintering process. Because of the relatively low granulation rate, large particles must be discarded and returned to granulation, which has a great impact on the smelting process. In the actual production process, the amount of fly ash should be controlled within 10% as much as possible.

3.2 Application of Waste Incineration Fly Ash in Sintering Process

53

Fig. 3.9 Picture of 5% fly ash granulation

Fig. 3.10 Picture of 10% fly ash granulation

3.2.4.2

The Effect of Fly Ash Addition on Sintering Process

Sintering experiment of 5% fly ash is shown in Fig. 3.11. During the experiment, the blending of fly ash caused some troubles to the sintering process. When the fly ash addition exceeds 5%, the descending rate of combustion layer becomes slow and the longitudinal combustion rate is uneven. When the fly ash addition exceeds 10%, the longitudinal combustion rate deviates greatly. Because of the difference of granulation and the accumulation of volatile alkali metals in the fly ash, the negative pressure of suction air is not smooth, which has a negative impact on the smoothness of the whole sintering. From the perspective of actual production, less than 5% fly ash is allowed. Most sintering machines lack real-time temperature measurement and control equipment. The instability of sintering process is fatal to sintering process. The accumulation of alkali metals in the material layer and its influence on coke combustion can not be ignored. When the fly ash addition is less than 5%, it has almost no effect on the combustion progress, and there is no obvious gap with the traditional sinter. When the fly ash addition is 2.5%, it is completely acceptable.

54

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Fig. 3.11 Sintering experiment of 5% fly ash

3.2.4.3

Basic Element Composition

Basic Element Composition of Sinter are showen in Fig. 3.12. The addition of fly ash has a weak influence on the main components of sinter. The main components TFe, CaO, SiO2 and Al2 O3 can keep stable. The main components can meet the charging requirements of blast furnace and maintain a good alkalinity balance in blast furnace.

3.2.4.4

Alkali Metal and Heavy Metal Composition

The content of alkali heavy metals is a key index to measure whether sinter can enter the furnace. When the addition of fly ash is 2.5 and 5%, the alkali metal content of sinter is controlled at a low value. However, when the addition of fly ash is 5%, the experimental results of several furnaces are unstable, and the alkali metal content rises and falls due to the instability of sintering process. This situation should be avoided in actual production. When the amount of fly ash is more than 7.5%, a large number of alkali metal residues lead to the sinter can not meet the standard of entering the furnace, which will lead to poor operation of the blast furnace and deterioration of the environment in the furnace. From the perspective of alkali metals, the fly ash addition is about 2.5%, which is completely feasible and alkali metals in fly ash can enter the dust in large quantities. When the dosage reaches 5%, although some heats have achieved good results, the quality control of sinter is unstable, so it is not recommended to dosage. When the proportion of fly ash is more than 5%, it will have a negative impact on blast furnace to a certain extent (Fig. 3.13).

3.2 Application of Waste Incineration Fly Ash in Sintering Process

55

Fig. 3.12 Basic element composition of sinter

Fig. 3.13 Alkali heavy metal content in sinter

3.2.4.5

The Effect of Sinter Product Ratio

Finished product ratio is the percentage of finished sinter (particle size greater than 6.8 mm) in sinter. It is closely related to the strength of sinter, which has high strength, strong crushing and wear resistance and high yield. The mineralization rate of sinter in the reference group is above 79%, and the yield of sinter is basically unaffected when fly ash is added less than 5%. However, when the amount of fly ash is more than 7.5, the mineralization rate decreases greatly and fluctuates obviously, without

56

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Fig. 3.14 Sinter product ratio

Fig. 3.15 Incineration-out phenomenon in sintering process

the production stability. Pulverization and insufficient combustion of sinter appear in local areas, which causes great damage to the sinter (Fig. 3.14). Because fly ash will bring some negative effects in the process of granulation and sintering, a large proportion of fly ash will burn out in the middle of sinter. Uneven heat transfer leads to excessive heat concentration in some areas, resulting in excessive liquefaction (Fig. 3.15). This has a bad influence on sinter mineralization.

3.2.4.6

The Effect of Tumbler Strength Index

Tumbler strength is an important index to evaluate the quality of sinter, which reflects the mechanical strength of sinter and has great influence on the technical and economic indexes of blast furnace smelting. Thick bed sintering, reasonable ore

3.2 Application of Waste Incineration Fly Ash in Sintering Process

57

Fig. 3.16 Tumbler strength index

blending and suitable process operation parameters have great influence on tumbler strength. The addition of fly ash has little effect on the tumbler strength of sinter (Fig. 3.16). However, because the tumbler strength is determined by mineralization, although there is no obvious difference in tumbler strength, the high proportion of fly ash has a certain negative effect on the overall sinter quality.

3.2.4.7

Removal and Distribution of Alkali Metals

In order to quantitatively calculate the migration and distribution behavior of harmful elements, we define two physical quantities, where η is the removal rate of harmful elements, which is the ratio of the difference between the total amount of harmful elements in raw materials minus the total amount of harmful elements in sinter and the total amount of harmful elements in raw materials. η=

MA × wtA % − MB × wtB % × 100% MA × wtA %

where M A is the amount of charge per experiment; M B is the weight of sinter; wtA % is the percentage of harmful elements in raw materials; wtB % is the percentage of harmful elements in sinter. Although the low proportion of fly ash has a good smooth furnace condition and the average content of sodium and potassium salts has a reasonable removal rate range, the unstable removal rate is still a factor worth considering. When the fly ash addition exceeds 5%, this unstable trend increases obviously. Although a higher removal rate occurs when the fly ash addition exceeds 10%, it does not affect the overall lack of removal rate (Fig. 3.17). This phenomenon is probably caused by the unsatisfactory sintering process.

58

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Fig. 3.17 Na, K removal rate

3.2.5 Summary (1) When a small proportion of fly ash into the sintering process, there is no negative impact on the sintering process and the ore-forming quality of sinter, can be used in the actual production process. (2) The sinter quality and alkali metal enrichment remain at a high level when the fly ash addition reaches 5%, but the fly ash addition will produce certain unstable factors in sintering production, which is not recommended in the actual production process. (3) When the proportion of fly ash is high (> 7.5%), the sintering process will be greatly affected, the sintering index will be deteriorated sharply, and the sintering process will be harmful. (4) In practical production, adding a small amount of fly ash is beneficial to the enrichment of alkali metals in fly ash, and has no obvious influence on sintering process.

3.3 Formation Mechanism of Binder Phase in Sintering Process with Fly Ash Instead of Lime Because in the enlarged experimental scale, because of uneven mixture, there may be differences in mineral phase distribution. Through laboratory experiments, the mechanism of fly ash on the formation of binder phase in sinter is discussed emphatically, and then the possible promotion and suppression effect of fly ash mixed with sinter on subsequent processes is clarified.

3.3 Formation Mechanism of Binder Phase in Sintering Process with Fly …

59

3.3.1 Experimental Design As shown in Fig. 3.18, the overall test plan needs to be formulated first: the first part is mainly to understand the basic physical and chemical properties of fly ash raw materials, and obtain relevant experimental parameters through the basic characteristics test; The second part is mainly about the thermodynamic calculation and the verification process of hot state experiment; The third part, through the discussion and analysis of the experimental results, we finally get the experimental conclusions. In order to thermodynamically explore the behavior and properties of different components in fly ash during the formation of the sintering binder phase, we made mixtures (labeled 1–5) of varying fly-ash-to-lime mass ratios using sintering raw materials commonly found in the sintering workshop of an iron and steel plant. Table 3.13 lists the recipes, and Table 3.14 summarizes the chemical composition of each mixture. An alkalinity of R = 2 [11] was selected for the high-temperature test at 1250 °C. We used a muffle furnace and the system was open to the atmosphere. The mineral powders were first ground to less than 0.074 mm, and placed in a drying box at 105 °C for 120 min. The sintered mixture was prepared according to the proportions given in Table 3.14. The lime and fly ash were also ground to less than 0.074 mm, and all the raw materials were made into 200 g samples [12]. As shown in Fig. 3.19, a mixture was placed into a 200 mL corundum crucible and heated to 1250 °C for 30 min [13, 14]. The mixtures produce large amounts of calcium ferrite liquid phase during the high-temperature process that reacts with Al2 O3 of the crucible. Therefore, it is necessary to crush the crucible after cooling to room temperature to remove the sintered product. The sintered product was removed and broken into a sample of less than 25 mm, and the sample was made into a cylindrical shape using resin to facilitate subsequent characterization experiments.

3.3.2 Thermal Dynamic Calculation 3.3.2.1

Phase Diagram Calculation

Calculations are performed using the Phase Diagram module in FactSage. When calculating the ternary phase diagram, it is necessary to set the oxygen partial pressure and the variable temperature range under the air condition. In the binary phase diagram calculation, the temperature is taken as the ordinate, and the element content is the variable to control the oxygen partial pressure under the air condition.

60

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Fig. 3.18 The source of fly ash and the overall scheme of the experiment

3.3 Formation Mechanism of Binder Phase in Sintering Process with Fly …

61

Table 3.13 Sintering mixtures comprising different proportions of fly ash (wt%) Label

Fly ash

Lime

Fly ash:lime

Mineral powder

1

0.00

11.8

0:100

88.20

2

3.32

9.96

25:75

86.72

3

7.58

7.58

50:50

84.84

4

13.26

4.42

75:25

82.32

5

21.2

0.00

100:0

78.80

Table 3.14 Chemical composition of sintering mixtures (wt%) Label

LOI

TFe

Fe2 O3

SiO2

CaO

MgO

Al2 O3

FeO

Zn

K2 O

Na2 O

1

6.07

55.22

74.95

4.17

8.34

0.55

1.41

3.54

0.01

0.05

0.02

2

6.66

54.33

73.74

4.16

8.32

0.52

1.69

3.48

0.03

0.28

0.53

3

6.64

53.20

72.22

4.14

8.29

0.49

2.04

3.40

0.07

0.59

1.19

4

8.39

51.69

70.17

4.13

8.25

0.44

2.52

3.29

0.11

1.00

2.08

5

9.78

49.57

67.31

4.10

8.20

0.37

3.18

3.14

0.17

1.57

3.31

Notes LOI—burning loss, TFe—the ratio of iron to all elements

Fig. 3.19 Muffle furnace experimental device

3.3.2.2

Gibbs Free Energy Calculation

Calculations are performed using the React module in FactSage. Input the reactants and products of various chemical reactions, adjust the oxygen partial pressure under air conditions, and explore the value of Gibbs free energy with temperature.

62

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Table 3.15 Chemical composition of fly ash and its burnt samples (wt%) Component

CaO

SiO2

NaCl + KCl

MgO

Al2 O3

Fe2 O3

ZnO

PbO

Fly ash

42.36

3.34

40.51

1.49

1.11

1.00

0.85

0.1

900 °C

72.46

6.54

12.71

4.55

3.14

1.84

1.23

0.02

1200 °C

79.41

5.34

5.42

3.81

3.47

2.37

1.30

0.01

3.3.2.3

Liquid Composition Calculation

Calculations are performed using the Equilib module in FactSage. The reactants and reaction conditions were input, the oxygen partial pressure under air conditions was adjusted, and Slag-liq was used as the target phase. The reaction system was set up to gradually heat up from normal temperature to 1350 °C to form a pure liquid phase. Using the non-equilibrium cooling calculation module, the pure liquid phase was cooled to explore the composition and proportion of the binder phase and other phases after liquid phase cooling under the influence of different Al2 O3 contents.

3.3.2.4

Viscosity Calculation

Calculations were performed using the Viscosity module in FactSage. The liquid component content at a certain temperature is calculated by the Equilib module, and the solid–liquid mixing ratio is calculated at the same time, and the viscosity of the solid–liquid mixture is calculated by the formula.

3.3.3 Results and Discussion 3.3.3.1

Chemical Composition of Fly Ash

Elemental and compound analysis of fly ash and its high-temperature burnt samples was carried out by X-ray fluorescence analysis (XRF). It can be seen from Table 3.15 that fly ash is mainly composed of CaO and alkali metal chlorides, with a small amount of metal oxides. Redetermination after heating fly ash at 900 °C for 3 h in a muffle furnace, revealed that mainly alkali metal chlorides were volatilized; similar results were obtained following heat treatment at 1200 °C [14].

3.3.3.2

Particle Size Distribution of Fly Ash

The fly ash particle size was measured by Malvern Laser Particle Size Tester (Malvern Instruments 2000), and the results are shown in Fig. 3.20. The main particle size is between 20 and 50 μm, which accounts for 50% of the total volume. The sintering

3.3 Formation Mechanism of Binder Phase in Sintering Process with Fly …

63

Fig. 3.20 Particle size distribution of fly ash

process requires that less than 5% of the raw materials is made up of particles less than 5 mm in size. If the raw material is too fine, it severely affects the sinter’s permeability. Therefore, the amount of fly ash added during the sintering process is limited [15, 16].

3.3.3.3

Specific Surface Area and Porosity of Fly Ash

The surface area of a 3.22 g sample of fly ash was analyzed using a porosity analyzer (MicroActive for ASAP 2460 Version 2.02). The medium was N2 at a temperature of − 195.8 °C. The resulting specific surface area and porosity analysis reports are given in Tables 3.16, 3.17 and 3.18. Table 3.16 Fly ash specific surface area analysis report Parameter

Explanation

Value, m2 /g

Single point surface area

P/P0 = 0.249

5.5209

BET surface area

Micropores, mesopores, and macropores

5.5931

Langmuir surface area

Monolayer adsorption model

7.5752

t-Plot micropore area

1.1392

t-Plot external surface area

4.4539

64

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Table 3.17 Fly ash pore volume analysis report Parameter

Explanation

Value, cm3 /g

Single point adsorption total pore volume of pores

P/P0 = 0.99 Diameter < 200.0820 nm

0.016203

Single point desorption Total pore volume of pores

P/P0 = 0.15 Diameter < 1.8126 nm

0.002551

t-Plot micropore volume

0.000524

BJH Adsorption cumulative volume of pores

Diameter 1.70–300.00 nm

0.015503

BJH Desorption cumulative volume of pores

Diameter 1.70–300.00 nm

0.015382

Table 3.18 Fly ash pore size analysis report Parameter

Explanation

Value, nm

Adsorption average pore diameter

4 V/A by BET

11.5877

Desorption average pore diameter

4 V/A by BET

BJH Adsorption average pore diameter

4 V/A

14.1554

BJH Desorption average pore diameter

4 V/A

13.2675

1.82410

Tables 3.16, 3.17 and 3.18 show the specific surface area, pore size and pore volume of the fly ash. It can be seen that fly ash has certain microporous and mesoporous characteristics, and the pore structure is dominated by flat slits and cracks. These slits are important structures for absorbing heavy metal elements and dioxins.

3.3.3.4

Phase Composition of Fly Ash

X-ray diffraction (XRD) was used to carry out phase analysis of the three fly ash samples. It can be seen from Fig. 3.21 that peaks for CaO and alkali metal chlorides (NaCl, KCl) predominate the diffractogram of unburnt fly ash. A large amount of alkali metal chlorides was volatilized after heat treatment at 900 °C. However, a large amount of oxides still remained in the system. During heat treatment at 1200 °C, the heavy metal oxides are lost, and CaO becomes the main residual component. A small amount of CaO reacts with SiO2 present in the system to form CaSiO3 at high temperatures [17]. The fly ash samples were observed under field emission scanning electron microscopy (FE-SEM). As shown in Fig. 3.22, use 20.00 kV scanning voltage, WD = 12.2 mm, where the magnification of the left image is 10.10KX, and the magnification of the right image is 3.15KX, most of the impurities in the fly ash exist in the form of molecular crystals that are relatively concentrated. As the product is collected in the system, each substance solidifies at a certain temperature, rendering it more concentrated. The larger ore phase molecules (CaO and Ca2 SiO4 , and Al2 O3 and MgO) are attached to the surface. Alkali metals exist as chlorides exhibiting regular cubic structure, and are concentrated together to form complex structures [18, 19].

3.3 Formation Mechanism of Binder Phase in Sintering Process with Fly …

65

Fig. 3.21 XRD diffractograms of fly ash and its burnt samples

Fig. 3.22 SEM and EDS images of fly ash with the following impurities identified: (1) PbS, (2) KCl, (3) Al2 O3 , (4) NaCl, and (5) Ca2 SiO4

3.3.3.5

Effect of Al2 O3 Content in Fly Ash on Sintered Binder Phase

Fly ash contains a small amount of Al2 O3 , which is one of the main components of SFCA, and can significantly affect the sintered binder phase. Table 3.19 summarizes possible reactions of Al2 O3 . In the Fe2 O3 –Al2 O3 –CaO–SiO2 system (Fig. 3.23), CaO and Fe2 O3 formed at a lower temperature. After the temperature reached 1000 °C, Al2 O3 had better solid solubility in CaO · Fe2 O3 , which enabled their blending to form aluminum-containing

66

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Table 3.19 Chemical reactions involving Al2 O3 in the sintering system Label

Reaction

1

CaO + Al2 O3 → CaO · Al2 O3

2

12CaO + 7Al2 O3 → 12CaO · 7Al2 O3

3

Al2 O3 + CaO + SiO2 → CaO · Al2 O3 · SiO2

4

7Al2 O3 + 12CaO · Fe2 O3 → 12CaO · 7Al2 O3 + 6Fe2 O3

5

Al2 O3 + CaO + 2SiO2 → CaO · Al2 O3 · 2SiO2

6

4CaO + Al2 O3 + Fe2 O3 → 4CaO · Al2 O3 · Fe2 O3

calcium ferrite, which is an important reaction intermediate for the formation of SFCA [20]. From the phase diagram of the Fe2 O3 –CaO–Al2 O3 –SiO2 quaternary system in an O2 atmosphere generated using FactSage software (Fig. 3.23), the liquid phase is mainly concentrated in the area with high Fe2 O3 content and low Al2 O3 content. However, the formation temperature of composite aluminosilicate is higher. At point 8, the liquid phase formation temperature is approximately 1250 °C, and the content of Al2 O3 is less than 1%. At point 7, after the Al2 O3 content exceeds 10%, the liquid phase generation temperature becomes 1350 °C.

Fig. 3.23 Equilibrium phase diagram of Fe2 O3 –Al2 O3 –CaO–SiO2 quaternary system

3.3 Formation Mechanism of Binder Phase in Sintering Process with Fly …

67

To explore the effect of Al2 O3 content on the formation of the sintered liquid phase and the phase of the precipitates, the liquid-phase precipitates were studied by a Scheil-Gulliver Cooling Calculation in the Equilib module of FactSage, as shown in Fig. 3.24. Figure 3.24a shows the relative composition of the sintered system when the liquid phase precipitation starts at the sintering temperature of 1350 °C, and Fig. 3.24b shows that of the completely cooled system. When no Al2 O3 is added to the system, a large amount of iron oxide does not dissolve into the liquid phase, and 55% of the system is in the liquid phase. The Scheil cooling module (non-equilibrium cooling module) in the Equilib module in FactSage was used for calculation from the point where the entire sintering reaction system begins to cool down from the highest sintering temperature until the liquid phase completely disappears. Fe2 O3 continued to precipitate and increased to 56% of the system, while the content of iron spinel did not change. The composition of the liquid phase was mainly silicate (Ca2 SiO4 ), hematite (Fe2 O3 ) and calcium ferrite (CaFe2 O4 , CaFe4 O7 ). Upon an increase in the content of Al2 O3 , the portion of the liquid phase gradually increased, and more iron oxide could be dissolved into the sintering liquid phase. When the system was completely cooled the content of silicate and ordinary calcium ferrite increased slightly, while the content of SFCA increased (by 5% and 19% for 1% and 2% Al2 O3 content, respectively), and the Fe2 O3 content dropped below 50%. This is ascribed to the facile reaction of Al2 O3 with silicate in the system, resulting in a large increase in the content of SFCA with increasing Al2 O3 content [21, 22]. When the Al2 O3 content in the system increased to 3%, the liquid phase content also increased, but more Fe2 O3 was dissolved in the enlarged liquid phase. At an Al2 O3 content of 4%, Fe2 O3 was completely dissolved in the liquid phase. At the same time, the content of complex iron oxide (spinel) increased. After this system was cooled, the silicate content in the system was higher, and a small amount of anorthite (Ca2 Al2 SiO7 ) began to appear in the system, while the content of calcium ferrite and SFCA decreased greatly. This is due to the similar sizes of Al3+ and Fe3+ (ionic radii of 51 and 64 pm, respectively); with an increase in Al2 O3 content, more Al3+ replace Fe3+ to form calcium aluminate (CaO · 2Al2 O3 ), resulting in the presence of more free Fe3+ , which in turn increases the amount of spinel [23].

3.3.3.6

Influence of Alkali Metals in Fly Ash on Sintered Binder Phase

The alkali metals are present in fly ash in the form of chlorides. As the temperature increases during the sintering process, most of the alkali metal chlorides are volatilized and separated from the system below 1200 °C, with a small amount being oxidized under the atmosphere and remains in the system [24]. The main reactions are shown in Table 3.20, and their Gibbs free energy curves are shown in Fig. 3.25. Alkali metal oxides (K2 O + Na2 O) are easily solid-dissolved into silicates at high temperatures. This enrichment promotes the dissolution of weakly acidic Al2 O3 and SiO2 , which can support the formation of SFCA [25]. The dissolution of (K2 O + Na2 O) also reduces the viscosity of the glass phase and improves gas permeability.

68

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Fig. 3.24 The effect of Al2 O3 content on the composition of the system at 1350 °C (a) and after cooling to room temperature (b) Table 3.20 Reaction of alkali metal oxides in the system

Label

Reaction

7

2K2 O + 2Al2 O3 + 8SiO2 → 4KAlSi2 O6

8

2Na2 O + 2Al2 O3 + 8SiO2 → 4NaAlSi2 O6

9

K2 O + CaSiO3 → CaO + K2 SiO3

10

Na2 O + CaSiO3 → CaO + Na2 SiO3

3.3 Formation Mechanism of Binder Phase in Sintering Process with Fly …

69

Fig. 3.25 Gibbs free energy curve for the reaction of alkali metal oxides

Therefore, the increase of alkali metal content will lead to the deterioration of the quality of the sintered binder phase. The conditions in Fig. 3.26 consisted of Fe2 O3 , CaO, and SiO2 contents of 83%, 10%, and 5%, respectively, (basicity R = 2) in an air atmosphere (PO2 = 0.21 atm). Under these conditions, the sintering liquid phase increases with the phase diagram with changes in alkali metal content and temperature. It can be seen from Fig. 3.25 that with the increase of alkali metal content, the temperature of the sintered liquid phase decreases, and simultaneously, the aluminosilicate material in the system begins to increase, and some alkali metals are solid-dissolved into the aluminosilicate or composite paste. The formation of alkali metal compounds in calcium acid affects the precipitation of calcium ferrite. The production of more alkali metal compounds will change the properties of the sinter. Figure 3.27 plots the viscosity of the system as a function of alkali metal oxide content. Equation (3.2) is the viscosity calculation formula of the sintered solid–liquid mixture. μ=

ν (1.35 × L − 0.35)2.5

(3.2)

where μ is the viscosity of the sintered solid–liquid mixture; ν is the pure liquid phase viscosity; L is the liquid phase in the mixture. The viscosity of the pure liquid phase is calculated by the Viscosity module in FactSage, which requires the Equilib module to calculate the content of each component in the pure liquid phase. The viscosity of the system gradually decreases with increasing alkali metal oxide content. This promotes the precipitation of the glass phase in the liquid phase, and

70

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Fig. 3.26 The effect of K2 O content on the sinter system

Fig. 3.27 Effect of alkali metal oxide content on system viscosity

3.3 Formation Mechanism of Binder Phase in Sintering Process with Fly …

71

Table 3.21 Chemical composition of sintered products with five different ratios (wt%) Label

Fe2 O3

CaO

SiO2

Al2 O3

MgO

P2 O5

ZnO

K2 O

Na2 O

Fly ash:lime

1

78.54

10.82

5.44

1.76

0.97

0.32

0.02

0.05

0.2

0:100

2

78.56

11.15

5.62

2.05

1.00

0.42

0.07

0.15

0.86

25:75

3

75.06

11.58

5.76

2.37

0.95

0.39

0.05

0.22

1.23

50:50

4

72.82

11.43

6.11

2.76

0.95

0.31

0.03

0.29

1.56

75:25

5

69.48

11.32

6.60

2.89

0.90

0.61

0.09

0.36

3.41

100:0

the alkali metal oxide acts as the fixation of the glass phase. This stabilizes aluminosilicate in the system and inhibits the dissolution of Fe2 O3 in the aluminosilicate, resulting in the unfavorable decrease of Fe2 O3 content in SFCA. Therefore, the alkali metal oxide has an inhibitory effect on the formation of SFCA, and its content in the mixture needs to be lowered [26].

3.3.3.7

XRF and XRD Results of Sintered Products

The chemical composition and mineral phases of each mixture were determined using XRF and XRD, respectively. It is evident from Table 3.21 that the content of iron oxides (Fe2 O3 ) in the sintered product gradually decreased as the proportion of fly ash increased. When all the lime in the mixture is replaced by fly ash (Label 5), the iron oxide content is the lowest. In addition, the admixture of fly ash resulted in more Fe2 O3 participating in the reaction. This is ascribed to fly ash increasing the reaction of Fe2 O3 with SiO2 and Al2 O3 to form SFCA, which increases the quality of calcium ferrite. Furthermore, it promotes the formation of the calcium ferrite interwoven structure, and improves the quality and metallurgical properties of the sinter, similar to the results found by Zeng [38]. As shown in Fig. 3.28, the mixture with a fly-ash-to-lime mass ratio of 25:75 (Label 2) shows highest the diffraction peak intensity for c-Ca2 Fe2 O5 . The small amount of Al2 O3 can effectively promote the formation of SFCA, which promotes the transformation of acicular calcium ferrite into the interwoven structure. As the proportion of fly ash increases, the intensity of the Ca2 Fe2 O5 diffraction peak decreases while the diffraction peak of Ca2 Fe9 O13 gradually increases. This indicates a rise in the Fe/Ca ratio, similar to the results found by Decker [39].

3.3.3.8

Optical Microscopy of Sintered Product

Epoxy resin was used to shape cylindrical samples that were inspected under an optical microscope to observe the morphology of calcium ferrite. As shown in Fig. 3.29, different compounds were observed, as indicated by the numbers: (1) hematite, (2) plate calcium ferrite, (3) acicular calcium ferrite, and (4) silicate. The

72

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Fig. 3.28 XRD diffractograms of sintered mixtures with different proportions of fly ash

calcium ferrite in the sintered products is mainly plate calcium ferrite, and acicular calcium ferrite begins to appear with increasing fly ash content. When a small amount of fly ash replaces lime, the calcium ferrite in the system showed an interwoven structure with the growth of plate-like calcium ferrite, and the content of hematite decreased. However, when the proportion of fly ash increased further, acicular calcium ferrite began to disappear, and the contents of plate calcium ferrite, hematite, and SiO2 increased, similar to the results reported by Yi [40]. The larger number of impurities in fly ash compared to lime causes the sintered product to contain more impurities. The amount of silicate glass phase gradually increased with increasing fly ash content, and the silicate formation was mainly localized in the spaces between calcium ferrite and hematite, similar to the results reported by Bethanis and Zeng [38, 41].

3.3.3.9

SEM Images of Sintered Products

To further investigate the effect of different fly ash addition ratios on the morphology and composition of calcium ferrite in the sintered products, the distribution of compounds was analyzed using FE-SEM combined with energy dispersive spectrometry (EDS). Results for the different mixtures are shown in Figs. 3.30, 3.31, 3.32, 3.33 and 3.34.

3.3 Formation Mechanism of Binder Phase in Sintering Process with Fly …

73

Fig. 3.29 Microstructures of high-temperature calcined samples of mixtures with fly-ash-to-lime mass ratios. Indicated are (1) hematite, (2) plate calcium ferrite, (3) acicular calcium ferrite, and (4) silicate a 0:100, b 25:75, c 50:50, d 75:25, e 100:0

As shown in Table 3.22, when the sintering flux contains no fly ash (Label 1) the molecular formula of plate calcium ferrite is Ca1.3 Si0.5 Fe3.4 O4.4 , and that of silicate is Ca4.6 Si2.1 O3.3 . For a 50:50 mixture of fly ash and lime (Label 3) the molecular formulas of calcium ferrite and silicate are Ca1.1 Al0.6 Fe4.1 O3.9 and Ca3.3 Si1.4 O5.2 , respectively. When only fly ash is added to the mixture (Label 5), the molecular formula of calcium ferrite is Ca1.1 Si0.3 Al0.5 Fe4.2 O3.8 , and that of silicate is Ca4 Si1.7 O4.3 [36]. When lime is used as the alkaline sintering flux, the calcium ferrite in the sintering product is mainly plate-shaped columnar calcium ferrite, which has good mechanical properties. The proportion of acicular calcium ferrite in the product increases when increasing amounts of fly ash replace lime. The overall calcium ferrite with interwoven structure has excellent strength and reducibility. However, when only fly ash is used as the sintering flux, the proportion of acicular calcium ferrite in the product decreases, resulting in poorer strength and reducibility of the sintered product, similar to the results reported by Wang [43]. The initial increase when a small amount of incineration fly ash is added is ascribed to the presence of Si and Al that promotes the formation of acicular calcium ferrite. However, when an excessive amount of fly ash is added, the structure of calcium ferrite changes. In addition, a large number of impurities in fly ash reduces the Fe content, resulting in a decreased amount of calcium ferrite in the product, which has an adverse effect on the sintered product, similar to the results reported by Decker, Kasper and Yin [39, 40].

74

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Fig. 3.30 SEM and XRF images (a), and EDS scan energy spectra (b) at points 1 and 2 of sintered products with a fly-ash-to-lime mass ratio of 0:100

3.3.4 Conclusions The salient findings of this study can be summarized as follows: (1) The main effective component of fly ash is calcium oxide, which acts as a basic flux in the sintering process, and reacts with iron oxides and silicon-alumina oxides in iron ore to form the main binder phase of sintered ore. calcium ferrite.

3.3 Formation Mechanism of Binder Phase in Sintering Process with Fly …

75

Fig. 3.31 SEM and XRF images (a), and EDS scan energy spectra (b) at points 1 and 2 of sintered products with a fly-ash-to-lime mass ratio of 25:75

(2) The content of Al2 O3 in the sintering mixture must be below 3% to maintain good reducibility of the sinter and avoid deterioration of its metallurgical properties. (3) Alkali metal chlorides in the fly ash are volatilized under high-temperature oxidative conditions, and a small amount remains in the system in the form of oxides that are easily combined with SiO2 and Al2 O3 to form aluminosilicates. This reduces the viscosity of the liquid phase, prohibiting dissolution of Fe2 O3

76

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Fig. 3.32 SEM and XRF images (a), and EDS scan energy spectra (b) at points 1 and 2 of sintered products with a fly-ash-to-lime mass ratio of 50:50

into the silicate, which negatively impacts the sintered binder phase. It is thus necessary to minimize the content of alkali metal oxides in the mixture. (4) When the Fe/Ca ratio in the system is low, and the resulting SFCA has an interweaved structure with excellent metallurgical properties. When the proportion of fly ash increases, the amount of acicular calcium ferrite in the sinter decreases, and the strength and reducibility of sintered ore decrease.

3.3 Formation Mechanism of Binder Phase in Sintering Process with Fly …

77

Fig. 3.33 SEM and XRF images (a), and EDS scan energy spectra (b) at points 1 and 2 of sintered products with a fly-ash-to-lime mass ratio of 75:25

(5) The defect of this study is that it fails to comprehensively solve the problem of removal and enrichment of alkali heavy metals during sintering, which is the main factor affecting environmental protection, but has little effect on the binder phase of products during sintering.

78

3 Digestion of Solid Waste During Sintering-Pelletizing Process

Fig. 3.34 SEM and XRF images (a), and EDS scan energy spectra (b) at points 1 and 2 EDS analysis of sintered products with a fly-ash-to-lime mass ratio of 100:0

3.3 Formation Mechanism of Binder Phase in Sintering Process with Fly … Table 3.22 Molecular formulas of main products of sintering with different fly ash addition ratios

79

Label Fly ash:lime Calcium ferrite

Silicate

1

0:100

Ca1.3 Si0.5 Fe3.4 O4.4

Ca4.6 Si2.1 O3.3

2

25:75

Ca1.5 Si0.5 Fe3.2 O4.2

Ca4.7 Si2.3 O3.5

3

50:50

Ca1.1 Al0.6 Fe4.1 O3.9

Ca3.3 Si1.4 O5.2

4

75:25

Ca8.5 Al5.6 Fe5.2 O3.4

Ca3.4 Si2.1 O4.1

5

100:0

Ca1.1 Si0.3 Al0.5 Fe4.2 O3.8 Ca4 Si1.7 O4.3

(6) In summary, it is found that using a small amount of fly ash as an alkaline flux for sintering can not only dispose of solid waste, but also effectively improve the metallurgical properties of sintered ore, but when the amount of fly ash is too large, it will have a negative impact. Therefore, in the actual production process, it is necessary to properly proportion according to the sintering raw materials. (7) The substantial processing capacity of steel production with existing environmental protection measures offers a resource utilization opportunity. Using the sintering process to consume urban waste incineration fly ash not only effectively disposes of bulk solid waste, but also helps realize its harmless treatment at much reduced production cost.

Chapter 4

Case Analysis of Blast Furnace Consumption Technology and Process

Abstract Copper slag and steel slag can be combined for iron extraction, with raw materials including Cu slag, steel slag, and lime. Sintering ore with a suitable ratio of these materials can produce iron concentrate powder with an iron grade of over 30 and alkalinity of 1.5. This powder can then be used to prepare oxidized pellets. However, softening and dropping temperatures of artificial minerals are low, resulting in poor air permeability and liquid flooding in the furnace. Separating metals from the large volume of slag requires high temperatures and temperature control above 1550 °C. After joint reforming and iron extraction, the remaining tailings can be consumed in cement and material production. Small blast furnaces can be used as solid waste disposal equipment, with technical assistance required for the heatintensive process of extracting iron from solid waste. The smelting process produces high calorific value gas that can be utilized through waste heat recovery and gas utilization facilities. Keywords Copper slag · Iron extraction · Artificial minerals · Blast furnace · Waste heat recovery

4.1 Production Case of Metallurgical Solid Waste Disposal in Tunnel Kiln-Blast Furnace Process Based on the research results of University of Science and Technology Beijing for many years, this project successfully solves the problem related to difficult treatment of steel slag as well as iron and steel dust sludge, and realize the cycle with the phase I project of the company, so that energy and waste can be fully utilized, thus truly achieving the industrial chain with “cyclicity”. Currently, the project has been put into production. The regenerated slag treated by tunnel kiln-melting gas-making furnace meets the use requirements of construction materials. Because the content of silicon (aluminum) oxygen tetrahedron with potential hydration activity in reconstituted regenerated slag is three times that of cement clinker. Therefore, water-quenched slag and steel slag from gas-making furnace are taken as building materials to © Metallurgical Industry Press 2023 K. Dong et al., Resource Utilization of Solid Waste by Thermometallurgy in Steel Processes, https://doi.org/10.1007/978-981-99-5655-5_4

81

82

4 Case Analysis of Blast Furnace Consumption Technology and Process

prepare the high-performance all-solid waste cementitious material, which can be used to prepare ultra-high-performance steel fiber reinforced concrete. Its compressive strength reaches 160 MPa, which is more than 5 times that of ordinary building concrete. The calorific value and purity of gas produced by the gas-making furnace meet the industrial energy demand. Because there are reducing substances in gasmaking furnace, excess reaction between C and O2 generates a large amount of CO. Under the condition of controlling the coal injection rate, the gas with stable composition can be produced. The production of solid waste carbothermal detoxification kiln needs the support of hot air. However, if the hot air is directly discharged through the flue, a large amount of residual energy will be wasted. At present, the residual energy recovery is carried out for high temperature hot air, and a large amount of hot air can be used for pellet drying and heat preservation. Iron by-products will be produced in the production process of solid waste reforming regeneration furnace, and its iron content is over 95%. Therefore, the complete reuse and utilization of all products can be realized. Among them, gas from gas-making furnace, molten regenerated slag and hightemperature hot air from tunnel kiln will be supplied to the production project of tailings ultrafine powder for internal digestion, regenerated slag will be used for the manufacture of construction material cementitious materials, and gas will be used for heat and energy supply. In this project, steel slag, dust and sludge produced by iron and steel industry are used as raw materials for comprehensive utilization of solid waste. It is planned to treat 200,000 tons of solid waste annually, including 100,000 tons of dust and sludge, 60,000 tons of regenerated iron-rich powder, 40,000 tons of steel slag, 70,000 tons of regenerated slag, 120,000 tons of melted iron by-products, 3 × 108 m3 of gas and 6.5 × 108 m3 of waste hot air.

4.1.1 Process Introduction The main raw materials used in this project are iron and steel solid wastes produced by metallurgical enterprises, including iron-containing solid wastes such as iron and steel dust sludge as well as a small amount of steel slag. The basic idea is as follows: Carbothermal detoxification is carried out by pre-reduction to reduce the difficulty of post-treatment, and then solid waste is reformed and regenerated by gas-making furnace to make slag and gas (Fig. 4.1). The solid waste treatment process of tunnel kiln-gas-making furnace has the advantages such as low investment, recycling, low cost, good environmental protection and good economic and social benefits. This process conforms to the current industrial policy of solid waste treatment in China, adapts to the national conditions of resources and energy in China, and opens up a new field for iron and steel industry in China to embark on the road of sustainable development. Through the process of tunnel kiln reduction—melting gas—making furnace melting, the whole process aims to realize mineral phase recombination and

4.1 Production Case of Metallurgical Solid Waste Disposal in Tunnel …

83

Fig. 4.1 Process roadmap

iron element separation of iron and steel solid waste in the way of mixing pelletmaking-reduction-melting. High-quality regenerated slag is used in the production of construction materials cementitious materials in phase I project, and the gas generated by gas-making furnace is used for drying and energy supply in phase I project. This project realizes the recycling of energy and effectively treats bulk solid waste. In terms of reduction equipment, small blast furnace is adopted, and the mixed materials are loaded into the melting furnace through conveyor belt for deep reduction of melting. The volume of melting furnace is 150 m3 , and the annual gas output is 3.0 × 108 m3 . Meanwhile, a supporting bag filter is set on the top of the furnace to separate metal-rich dust from high calorific value gas. Gas enters the gas storage and transportation equipment. A bag filter is added at the appropriate position of the melting furnace to remove dust, and the dust is further treated and utilized. The metallized pellets produced by pre-reduction tunnel kiln are used as raw materials in the melting gas-making furnace, and the metallization rate MFe of pellets is 70–80%. About 270 kg briquette coal and 200 kg small-size coke with particle size of 10–25 mm are used for every ton of melted iron by-products produced, which is difficult to be used in ordinary blast furnaces. Raw materials and fuel are loaded by the inner and outer layers of the shaft furnace, and lowered, respectively. Therefore, the fuel will not be squeezed and broken by mineral materials, which is a doublelayer charging structure. Metallized pellets gradually descend, heat and even melt in the inner layer, that is, the central furnace. The fuel descends from the outer layer to the lower exhaust outlet for combustion, and CO produced by direct reduction and

84

4 Case Analysis of Blast Furnace Consumption Technology and Process

incomplete combustion generates secondary combustion at the upper exhaust outlet when rising, so as to further heat up and promote the melting of burden. Metallized pellets are further reduced and melted in a gas-making furnace, and slag and iron are separated. By adjusting the basicity of slag, most sulfur can enter the slag, which has a strong desulfurization effect, ensures the quality of melted iron by-products, reduces the emission of SO2 into the atmosphere. Therefore, it is beneficial to environmental protection. The calorific value of gas produced by melting gas-making furnace is higher than that of ordinary blast furnace gas, which can be used as fuel for pre-reduction rotary hearth furnace, or for power generation and other external gas supply besides its own hot-blast stove.

4.1.2 Major Equipment Technology and Production Capacity 4.1.2.1

Production Capacity of Tunnel Kiln

The basic parameters of tunnel kiln are shown in Table 4.1. The length of furnace body is 135 m, and 350,000 tons of metallized pellets are produced every year. Hot pellets are prepared by hot pressing process for melting in blast furnace. In order to avoid deformation and pulverization caused by gravity extrusion in the production process of tunnel kiln, crucible is used to protect pellets during reduction. The pellet loading platform adopts fully automatic assembly, with 3–4 mechanical arms installed, each crucible is integrally connected and fixed with the arms, and then loaded in layers. Automatic loading and feeding are realized in the whole process, which reduces labor and is accurate in loading. In the pre-reduction stage of raw materials, the tunnel kiln roasting process is prepared. The length, inner width, inner height and cross-sectional area of tunnel kiln are 135 m, 2.5 m, 2 m and 5 m2 , respectively. The whole pre-reduction process Table 4.1 Basic parameters of tunnel kiln Item

Parameter

Height of furnace body

4m

Inner wall height of furnace body

2m

Inner wall width of furnace body

2.5 m

Effective area of furnace body inner bore

5 m3

Furnace body length

135 m

Effective area of pre-reduction tunnel kiln

270 m2

Unit productivity of pre-reduction tunnel kiln

110 kg/m2 h

Annual effective working time of pre-reduction tunnel kiln

7000 h (about 292 days)

Annual pellet production capacity of pre-reduction tunnel kiln

350,000 t

4.1 Production Case of Metallurgical Solid Waste Disposal in Tunnel …

85

Fig. 4.2 Main part of tunnel kiln (overhaul period)

is realized, and the process is completely automated. Through the track, the circular operation of the skip car is realized. During the whole reduction process in the kiln, the skip car is sent into the tunnel kiln by hydraulic press, and the time for the skip car to pass through the tunnel kiln is determined according to the raw materials and loading conditions. According to the actual production situation, the traveling time of the skip car can be changed. In addition, the reduction tail gas is discharged at the appropriate position of the tunnel kiln, some of which is used for drying in the pellet-making stage, and the remaining waste heat is used for the phase I project. Dust removal equipment is added at the appropriate position of tunnel kiln, and bag filter is used for dry dust removal. If other products need to be produced, smelting equipment can be added to carry out subsequent treatment on rare metal dust and metal salt (Fig. 4.2). The discharging workshop adopts the same fully automated equipment as the loading workshop. In the same loading workshop, discharge in layers, and directly invert the whole crucible group to ensure clean discharge. The track conveyor is installed to convey the used crucible group to the loading workshop. The conveying device can realize positioning and cooperate with the handle of the loading workshop. At this point, the pre-restore stage is completed, and the whole unmanned operation can be realized. After reduction, the pellets are partially pulverized, and the hot pellets are sifted, and the powder raw materials are continuously returned to the batching stage for recycling. In actual production, the pulverization ratio is less than 5%. 3 large-scale mixing silos are adopted, and 2 conveyor belts are adopted to feed the pre-reduction products, pulverized coal and coke into the blending silos at the same time. After being fully mixed, they fall on the conveyor belt and are transported to the feed inlet of blast furnace by the conveyor belt.

86

4.1.2.2

4 Case Analysis of Blast Furnace Consumption Technology and Process

Blast Furnace Part

Gas-making furnace includes main body of gas-making furnace, circulating fan, hot blast stove, bag filter and other main equipment, and its main parameters are shown in Table 4.2. The taphole of blast furnace is shown in Fig. 4.3. The discharged molten iron is directly poured and cooled, which is used to produce block pig iron for subsequent sales. Table 4.2 Basic parameters of blast furnace Item

Parameter

Furnace hearth diameter

2700 mm

Furnace bosh diameter

3200 mm

Furnace throat diameter

2200 mm

Effective volume

150 m3

Utilization coefficient

2.5–3 t/(m3 days)

Effective volume of gas-making furnace

150m3

Utilization coefficient of gas-making furnace

2.7 t/(m3 days)

Annual effective operation time of gas-making furnace

7200 h (about 300 days)

Annual production capacity of molten iron by-products of gas-making 120,000 t furnace Coke particles entering the furnace

200 kg/tHM

Briquette entering furnace

270 kg/tHM

Grade of pellet entering furnace

TFe ≥ 58%

Metallization rate in furnace

≥ 60%

Gas production

3 × 108 m3

Slag-iron ratio

585 kg/tHM

Fig. 4.3 Tapping part of small blast furnace (overhaul period)

4.1 Production Case of Metallurgical Solid Waste Disposal in Tunnel …

87

4.1.3 Energy Consumption of Production Revenue and Expenditure The main raw materials and components are shown in Table 4.3. Steel slag comes from a surrounding iron and steel enterprise, and converter steel slag is used to adjust the basicity of smelting slag. Iron-rich powder is obtained by deep crushing and magnetic separation of steel slag, the total iron content is more than 75%, and the iron in the iron-rich powder is mainly metallic iron. Iron and steel dust sludge is composed of iron-containing solid wastes such as iron oxide scale, steelmaking dust and sintering dust. Reducing agent and mixed raw materials are proportioned, and metallized and reduced in tunnel kiln, and the composition of metallized pellets is shown in Table 4.4. The metallization rate of pellets is 71.04% when the gas supplied by the melting furnace is fully reused to the tunnel kiln. When a higher metallization rate (> 80%) is achieved, additional natural gas supplement is needed to maintain the energy consumption in the tunnel kiln. Under the existing matching rhythm, too high premetallization rate will lead to a substantial increase in cost and energy consumption. A large amount of aluminum and magnesium oxides are brought in by solid waste materials. The alumina content of slag is higher than that of conventional blast furnace. Table 4.5 shows the main chemical composition of slag. During the production process, the viscosity and fluidity of slag will increase slightly. The theoretical smelting temperature is between 1500 and 1550 °C to ensure the separation of slag and metal. Higher Al2 O3 content is beneficial to improve the silicate strength of water slag. The strength of recovered water slag products is higher. As shown in Table 4.6, due to the use of a large amount of steel slag and recovered iron-rich powder, the P content in raw materials is higher. The content of P in molten Table 4.3 Main raw materials Composition

TFe

MFe

Fe2 O3

FeO

CaO

MgO

SiO2

Al2 O3

Steel slag

22.6

4.2

6.7

17.5

34.5

9.5

17.7

2.5

Iron-rich powder

75.3

47.0

13.6

19.9

3.4

3.6

6.6

2.4

Iron and steel dust sludge

61.7

16.9

50.5

10.9

10.3

3.9

3.6

3.8

Table 4.4 Composition of metallized pellets (wt%) Metallization rate, %

TFe

MFe

Fe2 O3

FeO

CaO

MgO

SiO2

Al2 O3

71.04

55.2

39.2

0.0

20.6

12.7

4.7

8.1

3.3

Table 4.5 Composition of molten slag (wt%) Composition

CaO

SiO2

Al2 O3

MgO

MnO

FeO

1/2S

TiO2

Content

39

28

15

14

0.7

0.6

0.7

1.2

88

4 Case Analysis of Blast Furnace Consumption Technology and Process

Table 4.6 Composition of molten iron obtained (wt%)

Σ

Composition

Fe

C

Si

Mn

P

S

Content

95.0137

3.6156

0.3600

0.6134

0.3693

0.0280

100

Table 4.7 Composition of gas received (vol.%) Composition

CO

CO2

H2

CH4

H2 O

N2

Content

39.1%

1.8%

4.7%

0.5%

2.9%

51.0%

iron is higher than 0.3%. Because the annual output of the production line is limited, pig iron can be mixed into the conventional converter in a small proportion, which will not cause additional dephosphorization burden. As shown in Table 4.7, the CO content of the obtained gas is 39.1%, which can fully meet the needs of hot blast stoves. And a large amount of gas can be reused in the pre-reduction tunnel kiln. Small proportion of gas can be reused to material processing production line. The annual output and energy consumption index are shown in Fig. 4.4. The annual output of molten iron is 120,000 t, the production of regenerated slag is 70,000 t, and the consumption of solid waste is over 200,000 t. Among them, the average pulverized coal used for pre-reduction is about 47,000 t, and the coal injection system of blast furnace consumes 40,000 t of pulverized coal. The coke consumption is 30,000 t. The annual production of metallized pellets in tunnel kiln is 260,000 t, all of which are used for further smelting in blast furnace, and there is no surplus for export. Under the optimal smelting index, the consumption of solid waste per ton of iron is 718.6 kg/tHM, the consumption of coal ratio is 551.9 kg/tHM, and the consumption of coke ratio is 300 kg/tHM. Compared with the reported DK process, the fuel consumption is slightly higher. At present, the existing problem is that the large tunnel kiln has limited treatment capacity and high energy consumption, which can not meet the treatment requirements of bulk solid waste. At present, it can only adapt to serve one or two surrounding iron and steel enterprises with an annual output of about 5 million tons.

4.2 Disposal of All Solid Waste in Copper Slag-Steel Slag Blast Furnace Process In order to realize the disposal of bulk solid waste, the pre-iron process needs to have the characteristics of rapid production and large output. Because the pre-reduction process needs a long holding time, the output is low under the same scale equipment. Referring to the DK small blast furnace technology which has been put into use in Europe, sintered and oxidized pellets are proposed as the preparation process of artificial mineral, which can realize large-scale consumption of solid waste with the

4.2 Disposal of All Solid Waste in Copper Slag-Steel Slag Blast Furnace …

89

Fig. 4.4 Annual production and consumption indicators

use of blast furnace. This chapter mainly puts forward a future process assumption and scheme, which has not been implemented in engineering at present. The research scheme refers to the blast furnace smelting process design and related experimental research methods. The ideal process flow is shown in Fig. 4.5. Copper slag, steel slag and lime are mixed and granulated in proportion, and then alkaline sinter is produced in a sintering machine. The flotation tailings are further crushed and magnetic separated to prepare iron-containing concentrate powder. After iron concentrate is pelletized, oxidized pellets are prepared by belt roaster or shaft furnace. The sinter, oxidized pellets and coke are distributed in a small blast furnace according to a fixed proportion. The molten iron is produced by referring to the blast furnace process, and the collected molten iron is used to prepare copper-containing iron-based materials such as weathering steel, and the tailings obtained from the production are further recycled. In view of the above anticipated smelting process, this chapter explores the smelting process in an all-round way through various laboratory experiments, and explores the feasibility of the process and possible problems. In order to

90

4 Case Analysis of Blast Furnace Consumption Technology and Process

Fig. 4.5 Schematic diagram of process

understand more easily, the copper slag referred to in this chapter is the tailings after flotation.

4.2.1 Solid Waste Raw Materials After flotation of smelting slag produced in the smelting process, tailings are obtained. The raw material composition is shown in Tables 4.8, 4.9, 4.10 and 4.11. After deep crushing, the flotation tailings are further magnetically separated. After magnetic separation, iron concentrate is obtained and used to prepare oxidized pellets. Steel slag is taken from a steel production enterprise around copper smelting enterprise, which is produced by electric arc furnace process. After metallic iron is screened out by the steel slag produced by electric arc furnace based on crushing—magnetic separation process, the residual tailings are used as raw materials for preparing sinter. In order to control the alkalinity, the additional lime components are shown in Table 4.11. LOI (Loss of Ignition) represents the weight of lime burned at high temperature. TFe (Total Fe) in the table refers to the total iron content. Table 4.8 Chemical composition of flotation tailings (FT) (wt%) Composition

TFe

CaO

SiO2

Al2 O3

FeO

Fe2 O3

Content

42.46

3.16

26.61

5.38

37.157

16.514

Composition

P

MgO

S

R

Fe3+ /Fe2+

Fe/SiO2

Content

0.094

2.5

0.698

0.12

0.4

1.52

4.2 Disposal of All Solid Waste in Copper Slag-Steel Slag Blast Furnace …

91

Table 4.9 Chemical composition of magnetic separation refined powder (IC) (wt%) Composition

TFe

CaO

SiO2

Al2 O3

FeO

Fe2 O3

Content

44.98

2.77

19.48

5.87

32.657

27.971 Fe/SiO2 2.31

Composition

P

MgO

S

R

Fe3+ /Fe2+

Content

0.0966

1.54

0.898

0.14

0.77

Table 4.10 Chemical composition of steel slag (SS) (wt%) Composition

TFe

CaO

SiO2

Al2 O3

FeO

Fe2 O3

Content

21.38

44.002

18.9857

3.28667

22.487

5.557

Composition

P

MgO

S

R

Fe3+ /Fe2+

Fe/SiO2

Content

0.322

16.2

1.41

2.317

0.222

1.126

Table 4.11 Lime composition (wt%) Composition

SiO2

CaO

MgO

Al2 O3

LOI

FeO

K

Na

Content

4.54

70.05

3.94

0.43

20.81

0.23

0.2

0.022

4.2.2 Design and Preparation of Artificial Mineral Under this goal, 2 ratios can be considered for preparing sinter. One is to use copper slag as the leading factor, and use steel slag and lime to mix and adjust alkalinity for sintering. The other is dominated by steel slag, and a small amount of copper slag and lime are used. These two methods can control alkalinity and total iron content. There is a small difference in MgO and Al2 O3 in sinter, as shown in Table 4.12. Before the sinter is formally prepared, the return mine is prepared according to a fixed proportion. In the sintering process, the return mine also uses all solid wastes as raw materials. The sinter was broken to below 3.5 mm after return mine is sintered in furnace, which was used as return mine for subsequent sintering experiment. The difference between the two formulations is reflected in MgO. Because dolomite needs to be added in the steelmaking process for slag making, steel slag brings a large amount of MgO into the sintering process. Table 4.12 Experimental formula table (%) No.

Proportioning

Water

Steel slag

Copper slag

Caustic lime

Return mine

Coal char

Return mine 1

19.43

56.13

19.43

0

5

5

Return mine 2

48.58

35.62

10.79

0

5

5

Sample 1–3

18

52

18

7

5

5

Sample 4–6

45

33

10

7

5

5

92

4 Case Analysis of Blast Furnace Consumption Technology and Process

Using copper slag FT, steel slag SS and lime as raw materials, sinter was prepared in a 40 kg sintering pot. After mixing, granulation, ignition and sintering, sinter for iron extraction was prepared. The ignition temperature is 1150 °C, the ignition negative pressure is 9 kPa, and the sintering negative pressure is 11 kPa. In the sintering process, the plug-in thermocouple is used to record the temperature of the material layer. The experimental flow is shown in Fig. 4.7. After the sintering pot experiment, the sintering quality of sinter was evaluated by means of dropping screening test and tumble strength test. Refer to Chinese national standards GB/T 8209—1987 and YB/T 4606—2017 for specific implementation methods. The test results of sinter dropping screening and tumble strength are shown in Figs. 4.8 and 4.9. From the perspective of dropping screening and tumble index, the sinter quality of each furnace reaches the ideal index. The sintering mineralization rate and tumble index are higher than the required values [43, 44]. The consolidation mechanism of pellets can be divided into crystal bridge consolidation, solid phase bonding and liquid phase bonding. There are also differences

Fig. 4.7 Sinter preparation process

Fig. 4.8 Dropping screening index

4.2 Disposal of All Solid Waste in Copper Slag-Steel Slag Blast Furnace …

93

Fig. 4.9 Indicators of tumble strength

in the selection of calcination temperature [43, 44]. According to the composition of pellets, the phase diagram of CaO–SiO2 –FeO–O2 –Al2 O3 –MgO was drawn by FactSage 7.1, and the temperature selection design was carried out. The relationship between oxygen partial pressure and liquidus is shown in Fig. 4.10. During the oxidation roasting process, the oxygen potential corresponding to the pellet body gradually increased. Magnetite and Fayalite in raw materials will produce Fe2 O3 in the oxidation process, which will produce new crystal bridges and then consolidate. With the increase of temperature, adhesion between solids will also occur.

Fig. 4.10 Phase diagram of CaO–SiO2 –FeO–O2 –Al2 O3 –MgO

94

4 Case Analysis of Blast Furnace Consumption Technology and Process

The choice of experimental temperature depends on whether liquid phase consolidation occurs. 1200 and 1150 °C are selected as calcination temperatures, and the calcination system is shown in Table 4.13. The mixed binder was prepared by using 95% Liufangzi bentonite [45] and 5% CMC fiber as pellets. 5 g binder was added to every 100 gIC, and then pelletized by disc pelletizer. The diameter of pellets is about 10–12 mm. The prepared pellets were put into a corundum crucible with holes, heated from room temperature to 900 °C at a rate of 10 °C/min, and kept for 0.5 h to preheat and roast. Then, the temperature was raised to the roasting temperature at a rate of 10 °C/min, and the roasting was carried out at high temperature after holding for 1 h. After roasting, the temperature is lowered to 900 °C and soaked for 0.5 h, and then cooled to room temperature with the furnace. During the experiment, a large amount of air is blown at the bottom of the tubular furnace to ensure that the roasting atmosphere has high oxidizability. After the pellets are prepared, the compressive strength of the pellets is tested. When the compressive strength is greater than 2000 N, the pellets are considered to be successfully prepared. The experimental flow is shown in Fig. 4.11. In this chapter, fully oxidized pellets are taken as the research object in the preparation of pellets. When the traditional three-stage furnace is used as the preparation equipment, the oxidation of iron-containing silicates such as is unstable, and the Table 4.13 Pellet roasting system (°C) No.

Preheat temperature

Calcination temperature

Soaking temperature

1

900

1200

900

2

900

1150

900

Fig. 4.11 Pellet preparation process

4.2 Disposal of All Solid Waste in Copper Slag-Steel Slag Blast Furnace …

95

Fig. 4.12 Compressive strength test results

short oxidation time leads to the fluctuation of iron oxide content in pellets. In this chapter, excessive oxidation roasting is carried out to ensure its repeatability. The test results of compressive strength of oxidized pellets are shown in Fig. 4.12. The preparation temperature of No. 1 pellet is 1200 °C, which is higher than the liquidus temperature of copper slag. Liquid consolidation can effectively improve the strength of pellets. Compared with pure iron ore roasting, the higher silicon content of copper slag raw material leads to a large number of liquid phase formation, and its compressive strength is between 2500 and 3100 N, with an average value of about 2800 N. The roasting temperature of No. 2 pellet is 1150 °C. Although there is no liquid phase formation, the compressive strength of No. 2 pellet can be stabilized at about 2000 N on average by the formation of new crystal bridge and solid phase sintering. Copper slag can be prepared into pellets meeting the charging conditions by oxidation roasting.

4.2.3 Reducibility of Artificial Mineral Reducibility is an important index to evaluate metallurgical properties of artificial mineral under simulated reduction conditions in the upper high temperature zone of blast furnace [45]. The sinter and pellet were reduced at 900 °C with CO/N2 = 30/70 mixed gas. The value RI obtained from reducibility test is used as a reference index of indirect reduction degree in furnace. The RI is shown in formula (4.1). ( RI =

) M1 − Mε 0.11B + × 100 × 100% 0.43A 0.43A × M0

(4.1)

where A is the total iron content in artificial mineral; B is the FeO content in artificial mineral; M 1 is the sample weight before reduction; M e is the sample weight after

96

4 Case Analysis of Blast Furnace Consumption Technology and Process

Fig. 4.13 Reducibility test index of sinter. a Weight loss curve; b RI curve

reduction; M 0 is the sample weight, taking 500 g. The reducibility test method follows the Chinese national standard GB/T 13241—2017. In this chapter, RI index is only used to evaluate the deoxidation characteristics of artificial minerals, and can not fully reflect the reduction reaction of iron oxides. The weight loss during indirect reduction is caused by the reduction of iron oxides and the removal of volatile elements from copper slag. During the reduction process, the weight loss curve of sinter fluctuates greatly (Fig. 4.13), and the reducibility of sinter is unstable. Different proportions lead to the difference of reducibility of sinter. The RI of sinter with copper slag as the main raw material is between 55 and 65%. Under the same test conditions, RI index is lower than that of conventional ironmaking sinter. The RI of sinter dominated by steel slag is between 40 and 55%. The instability of sinter quality will cause great difficulties to the subsequent melting process. No. 4 sinter began to lose weight after introducing reducing gas for 10 min. The encapsulation of silicate can not make iron oxide contact with gas in time, which leads to the lag of weightlessness start-up time. The reducibility test of pellets shows great differences (Fig. 4.14). When reducing gas was introduced for 50 min, the weight loss of No. 1 pellet began, and the start time of weight loss lagged obviously compared with that of No. 2 pellet. With the increase of time, the weight loss of pellets is basically linear. In the previous characterization of pellets, it is found that a large number of silicates are encapsulated, and the iron oxides in the pellets can react only after the silicates in the periphery are decomposed. Liquid phase consolidation is extremely unfavorable for reduction of pellets prepared from solid waste. 2 roasting methods lead to obvious difference in RI. The RI of No. 2 pellet is more than 45%, while that of No. 1 pellet is around 25%. Further improvement of RI requires longer reduction time. Although the iron content of pellets is high, it can effectively reduce the amount of slag in the subsequent melting process. Artificial mineral has poor reducibility, which will increase the heat consumption of direct reduction. At present, in traditional blast furnace ironmaking, high pellet ratio is one

4.2 Disposal of All Solid Waste in Copper Slag-Steel Slag Blast Furnace …

97

Fig. 4.14 Reducibility test index of oxidized pellets. a Weight loss curve; b RI curve

of the means to reduce energy consumption, but this method may not be applicable under the condition of taking solid waste as raw material. The reduced sinter is mainly composed of Ca–Al–Mg silicate, Fe (Ca, Al) silicate, Wuestite and metallic iron. In the reduction process, Magnetite and Hematite react with CO and change to FeO form. The transformation of ferrous oxide to metallic iron is the key to indirect reduction. Because iron silicate is difficult to reduce, a large amount of iron oxides remain in the ore phase. Compared with traditional sinter, the soft melting drip temperature may drop significantly [46, 47, 48].

4.2.4 Charge Melting Dropping Characteristics The softening-melting and dropping behavior of iron-bearing furnace burden directly affects the fuel consumption and smelting efficiency of blast furnace. The dropping performance of blast furnace comprehensive burden softening mainly includes softening performance melting performance dropping performance and gas permeability of charge column and so on which can be tested by dropping experiment. The experimental samples are sinter and pellets after reduction test, and 180 g samples are weighed according to the ratio for mixed distribution. The upper and lower coke quantities are 20 g respectively. The machined graphite crucible is used as the experimental vessel, and the specific operation mode is fully introduced in the relevant literature. The load is set to 1 kg/cm2 . The temperature was raised to 1000 °C at the rate of 10 °C/min. Because the artificial mineral has been reduced by gas in the reduction test, when the temperature is below 1000 °C, the gas introduced into the furnace body is N2 . After 1000 °C, the temperature rises slowly at 3 °C/min, and the inlet gas is 70% N2 and 30% CO. From the dropping characteristic curve, the characteristic indexes which characterize the dropping process of reduction softening of comprehensive burden can be

98

4 Case Analysis of Blast Furnace Consumption Technology and Process

obtained, including softening start temperature, softening end temperature, melting start temperature, dropping temperature and so on. The total characteristic value S is shown in Eq. (4.2). (Td S=

(Δ Pm − Δ Ps )dT

(4.2)

Ts

where T s is the starting melting temperature, that is, the steep rise temperature of pressure difference; T d is the temperature at which dripping begins; Δ Pm is the maximum pressure difference in molten state; Δ P is a fixed value, which is 490 Pa in this chapter. The softening-melting and dropping data is shown in Fig. 4.15, and the specific parameters are shown in Table 4.14. The combination of No. 2 sinter and No. 2 pellet shows better reducibility in reducibility test. With the increase of reduction degree and metallization rate, the decrease of FeO content in gangue will lead to the increase of melting temperature. Generally, the increase of indirect reduction degree will lead to the widening of softening interval TB and the narrowing of softening temperature interval T, which is beneficial to the air permeability of burden. From the experimental results, the artificial mineral with copper slag and steel slag as raw materials has the opposite result. The combination of No. 4 sinter and No. 1 pellet with the worst reducibility has the phenomenon of widening softening interval and narrowing softening interval. Under the condition of high metal and low gangue in conventional ironmaking, the higher indirect reduction degree increases the melting point of gangue and metal iron content, increases the deformation resistance of burden, and strengthens the air permeability of burden. The increase of melting point of slag and metal is beneficial to the narrowing of soft melting interval. The increase of indirect reduction degree improves the softening-melting and dropping performance of burden. When solid waste is used as smelting raw material, the iron content in artificial mineral is low, the gangue ratio is high and the reducibility is poor. Due to the low content of metallic iron, the soft melting characteristics and fluidity of artificial gangue determine its dropping characteristics. The worse the indirect reduction degree, the higher the FeO content, the lower the viscosity and the better the fluidity of the slag with basicity equal to 1. Under the two experimental conditions, the temperature difference of T s is small, and the melting temperature of slag is consistent. Because of the high FeO content, the slag viscosity of No. 2 sample is lower than that of No. 1 sample with the increase of temperature, which can effectively explain the narrowing of soft melting interval. During the experiment, the TB of samples with low indirect reduction degree widened. Metals are a few phases in reduced artificial minerals, and the deformation resistance of minerals mainly depends on gangue. The shrinkage interval of No. 2 sample widens, which may have little relationship with indirect reduction degree. The difference of sinter composition may also be the factor affecting the dropping performance. The proportion of steel slag and MgO content

4.2 Disposal of All Solid Waste in Copper Slag-Steel Slag Blast Furnace …

99

Fig. 4.15 Dropping test data. a 75% 2# sinter + 25% 2# pellet; b 75% 4# sinter + 25% 1# pellet

in No. 4 sinter are higher. The promotion of MgO to smelting properties has been reported in many literatures. Because of the low dropping temperature, C and FeO cannot react quickly due to temperature limitation. This phenomenon leads to the coke retention zone and the middle and lower part of the furnace body taking on the main task of direct reduction. The large amount of heat absorbed by direct reduction in the belly means that higher tuyere combustion temperature is needed to maintain sufficient heat supply. When copper slag is reduced by blast furnace process, there are differences between blast furnace design, charge layer structure and temperature distribution and traditional blast furnace.

100

4 Case Analysis of Blast Furnace Consumption Technology and Process

Table 4.14 Droping experimental parameters No.

1

Ingredients

75% 2# sinter + 25% 2# pellet 75% 4# sinter + 25% 1# pellet

2

Temperature TBS at shrinkage of 10%, °C

1184

1180

Temperature TBE at shrinkage of 40%, °C

1217

1222

Δ TB = TBE − TBS, °C

33

42

Shrinkage value at T s , mm

30

29

Shrinkage value at T d , mm

58

54

Temperature T s when the pressure difference begins to rise sharply, °C

1221

1224

Initial dripping temperature T d , °C

1285

1269

Δ T = T d − T s , °C

64

45

Maximum pressure difference Δ Pm , Pa

610 × 9.8

609 × 9.8

Total characteristic value S, kPa °C

351.23

246.52

4.2.5 Reduction Melting Results The melting experiment was carried out in a high temperature tubular furnace. The coke was crushed to a block with a particle size of 5 mm and laid on the bottom of a graphite crucible. The 6–10 mm size of massive sinter and pellets are mixed according to the preset ratio, and then placed on the upper layer of coke. The total mass of sinter and pellet is 500 g, and excess coke is 200 g. After heating up to 900 °C at a rate of 10 °C/min, a straight tube was inserted into the bottom of the material layer, and the mixed gas of 70% N2 and 30% CO was injected into the material layer at a rate of 10 L/min, and the temperature was kept for 1 h. After the heat preservation is finished, keep the gas flowing in and raise the temperature to 1100 °C at a rate of 3 °C/min, and pull out the straight bobbin and fix it above the material layer. Gas injection before 1100 °C simulates indirect reduction period, and injection after 1100 °C ensures that the reducing atmosphere of the material layer is always maintained. A round baffle is customized above the injection pipe to prevent solid materials from splashing during gas injection. After 1100 °C, the slag and metal can be separated fully by heating at a rate of 3 °C/min to the target temperature of melting separation. After heating, the slag and metal can be fully separated by holding for 3 h. After heat preservation, the sample is cooled to room temperature with the furnace. After separation, the graphite crucible was broken, and metal and slag samples were taken for subsequent characterization

4.2 Disposal of All Solid Waste in Copper Slag-Steel Slag Blast Furnace …

101

Fig. 4.16 Relation between metal yield mass and melting temperature

and analysis. During the melting experiment, the spiral condenser tube was used to collect dust for subsequent characterization. Temperature selection is an important parameter in melting experiment. When the iron oxide in the slag is completely reduced by coal, because a small amount of elements such as C and Si are dissolved in the molten iron, the mass of iron samples collected in the experiment should be greater than that of the total iron in the ingredients. When considering the temperature of slag and iron in conventional ironmaking, 1505 °C is usually used as the reference temperature of molten iron and slag. It is found that when the temperature is less than 1550 °C, the large slag amount and the limit of slag viscosity lead to the difficulty of molten iron separation (Fig. 4.16). A large number of unseparated metal droppings are embedded in the inside and surface of tailings, which directly affects the quality of metal recovery. In the actual production process, higher slag gold temperature is needed to meet the demand of metal recovery, which also means that slag needs to take away more heat. Because of the higher reduction temperature, the dissolved amount of C and the reduction amount of Si in metal are larger than those in conventional ironmaking process. In the subsequent steelmaking process, the control of decarburization and temperature rise needs to be adjusted. Because of the introduction of a large amount of P and S in raw materials, the difficulty and production cost of subsequent dephosphorization and desulfurization are greatly increased. According to the previous design, producing some weathering steel products containing Cu and P can effectively avoid this problem, but residual elements such as As, Pb and Zn in molten iron will have a negative impact on the quality of subsequent steel. The components of molten iron and slag obtained are used for energy consumption evaluation and calculation, as shown in Tables 4.15 and 4.16.

102

4 Case Analysis of Blast Furnace Consumption Technology and Process

Table 4.15 Components of pig iron obtained (1550 °C) (wt%) Component

C

S

P

Si

Mn

As

Cu

Content

4.5

0.15

0.12

0.62

0.12

0.47

0.593

Component

Pb

Zn

Sb

Cr

Cd

Mo

Content

0.003

0.010

0.742

0.2345

0.00084

0.11949

Table 4.16 Obtained slag composition (1550 °C) (wt%) Component

SiO2

CaO

Al2 O3

MgO

FeO

MnO

Content

37.84

39.01

11.72

5.83

1.07

0.637

Component

Cu

Pb

Zn

As

Cr

Sb

Content

0.02

0.0016

0.0112

0.0139

0.034

0.01683

4.2.6 Material Energy Balance in Blast Furnace Melting Process Fuel substitution can be realized by injecting pulverized coal into tuyere, and coke consumption can be reduced. Because the reasonable design between pulverized coal injection rate and blast is not clear, the whole coke is used as the smelting carbon source in the calculation process of this chapter. Figure 4.17 shows the relationship between direct reduction degree and coke consumption per ton of iron. According to the above analysis, the reducibility of artificial mineral is extremely poor, and the estimated direct reduction degree in furnace is between 0.55 and 0.65 or above. The worse the quality of artificial mineral, the higher the direct reduction degree in furnace. With the increase of direct reduction degree, the coke consumption per ton of iron is about 640–680 kg. Also directly related to direct reduction degree are hot air volume and gas composition. With the increase of direct reduction degree, the demand for burning coke and heat release increases, and the hot air volume per ton of iron also increases greatly, which makes it difficult to control the air volume due to the fluctuation of artificial mineral’s own quality. With the increase of direct reduction degree, the endothermic ratio of reduction reaction increases slightly, as shown in Fig. 4.18. The physical heat carried away by slag accounts for the main part of heat expenditure. Although direct reductionmagnetic separation process can reduce coke consumption, it also needs a lot of heat supply for slag heating because of the high proportion of gangue in solid waste raw materials. High temperature characteristics are beneficial to the volatilization and removal of harmful elements in slag, so it is necessary to consider the environmental hazards of slag and the consumption of smelting energy.

4.3 Prospect of Solid Waste Treatment by Blast Furnace Process

103

Fig. 4.17 Material consumption of blast furnace relationship between coke consumption per ton of iron and r d (a), and relationship between gas composition, theoretical air volume and r d (b)

Fig. 4.18 Ratio of balanced heat income items to expenditure items in boiler

4.3 Prospect of Solid Waste Treatment by Blast Furnace Process Blast furnace ironmaking is the most mainstream method in China, and years of technical accumulation can ensure that there is no technical difficulty in the implementation of this process. Compared with direct reduction method, blast furnace smelting has higher energy utilization efficiency and stronger reduction environment. With the residual energy recovery device, the energy consumption can be controlled within an acceptable range. On the premise of iron oxide modification, some low coke ironmaking processes such as COREX can also be used for reference. The primary purpose of solid waste reforming is not to produce high-quality molten iron, but to realize large-scale consumption of refractory solid waste and

104

4 Case Analysis of Blast Furnace Consumption Technology and Process

hazardous waste. Its raw material adaptability can be extended to coal gangue, sulfuric acid residue, red sludge, fly ash and so on. At present, 1000 m3 small blast furnace is facing the demand of upgrading from small to large, and the existing small blast furnace can reduce the cost of rebuilding enterprises with waste disposal needs. In the actual production process, it is necessary to consider relevant specific technical problems. Because the reaction of harmful elements in the furnace greatly hinders the smooth operation of the blast furnace, it may lead to many problems such as nodulation and blockage of the blast furnace. When implemented concretely, the life of furnace body and actual smelting effect are still unclear. At present, it is still to be discussed whether it is necessary to reduce the shaft of blast furnace. Because of the large amount of slag in the furnace, there is a need for continuous slag discharge in the production process. In the case of difficult separation of slag and iron, whether the cooled slag needs further crushing and magnetic separation needs to be verified. From the experimental results, the molten iron still contains As and other harmful elements. At present, As residue is a key problem in iron and steel smelting, although some literatures put forward the method of removing arsenic from molten steel. Due to the influence of harmful elements, solid wastes such as copper slag can not produce high-quality steel materials at present. This process is limited to copper slag reforming and molten iron production, and how to further develop it is still a problem to be solved. It is necessary to control the contents of As and Sb in raw materials before reforming copper slag and extracting iron. Using blast furnace as the main smelting process, there are two main ways to choose the pre-iron process. One is to prepare metallized pellets by pre-reduction, including tunnel kiln and rotary hearth furnace. The advantage of this process is that the raw materials containing Pb and Zn heavy metals and Na and K alkali metals can be enriched and recovered in the pre-reduction stage without causing harm to the smooth operation of blast furnace. The main defect is that the production capacity of pre-reduction is limited, so it is necessary to design the blast furnace body according to the treatment capacity of pre-reduction. Due to the fluctuation of pre-reduction itself, it is difficult to realize the coordination and optimization between them. When sinter and pellet are used as iron pre-process, it has the advantages of large output and rapid production. However, the heavy metal treatment burden of alkali is borne by the blast furnace alone, and the smooth operation of the furnace body may be greatly hindered. Secondly, the indirect reduction in blast furnace is extremely limited, and pre-metallization means such as tunnel kiln can reduce the coke consumption in direct reduction in blast furnace. A large amount of heat consumption of slag can be supplemented by means of pulverized coal blending and blowing. With the increase of coke price, the economic advantage of blast furnace disposal is lower.

Chapter 5

Disposal of Solid Waste in Steelmaking Process

Abstract The reaction mechanism models of dechlorination and [Zn] reclaiming in the roasting steelmaking zincrich dust process are studied. The dust collected from a steelwork contains 63.8% zinc and 3.18% chlorine (mass percent), of which, almost all zinc elements exist in ZnO and ZnCl2 forms, and all the chlorine elements are stored in ZnCl2 . When the dust is roasted at above 732 °C in an air atmosphere, the ZnCl2 in the steelmaking zinc-rich dust is volatilized into steam and then oxidized into ZnO. Finding the position where the chemical reaction occurs is the key to determining the reaction mechanisms of dechlorination and [Zn] reclaiming. In this study, two groups of thermal experiments are designed and executed for roasting in different atmosphere environments and at different roasting temperatures. Based on the experiment results, the mechanism model is discussed and built, and the reaction of dechlorination and [Zn] reclaiming is shown to be a multi-step process. Because O2 from the air cannot transmit into the dust particle interior or dust bed effectively, the chemical reaction of [Zn] reclaiming occurs in the external gas environment outside of the dust, where the [Zn] recalcining reaction should be limited by the dynamics of new nucleation of ZnO solids. Keywords Mechanism model · Dichlorination · Thermal experiment

5.1 Reuse and Zinc Enrichment Technology of Zinc-Containing Dust in Electric Arc Furnace 5.1.1 Brief Introduction of Process In an electric arc furnace steelmaking enterprise, all scrap steel is used as raw material, so a large amount of zinc-containing dust is produced. Zn in dust mainly exists in the form of zinc ferrite. The original small rotary hearth furnace was rebuilt to recover zinc-containing dust, but it was abandoned because of high energy consumption and poor economy. Meanwhile, it is difficult to carry out the extraction of zinc ferrite. The project team proposed that zinc-containing dust and coal are mixed to form briquetting, which will be used in electric arc furnace to realize carbon pick-up and © Metallurgical Industry Press 2023 K. Dong et al., Resource Utilization of Solid Waste by Thermometallurgy in Steel Processes, https://doi.org/10.1007/978-981-99-5655-5_5

105

106

5 Disposal of Solid Waste in Steelmaking Process

Fig. 5.1 Schematic diagram of electric arc furnace cycle

iron increase, and Zn in dust can be collected after multi-furnace cycle. When the ZnO content in dust is more than 90%, centralized recovery is carried out. The process improvement scheme is shown in the Fig. 5.1. Carbon powder is mixed with steelmaking dust with low zinc content, and then briquetting is carried out. The cold briquetting is reused for smelting in electric arc furnace, and the high temperature environment of smelting in electric arc furnace is used to realize carbon pick-up and iron recovery. Because Zn has high volatilization, Zn is effectively enriched after multiple furnaces (about 20 furnaces). When the zinc content in zinc-rich dust is higher than 60%, ZnO oxide can be enriched by high temperature roasting. Because there may be various organic compounds containing Cl in the raw materials of electric arc furnace scrap steel, and ZnCl2 exists in the dust. Due to the high-temperature volatilization of zinc chloride, zinc oxide and chloride can be separated efficiently at 700–900 °C. The separated ZnCl2 and ZnO are collected independently, which has a very high sales prospect.

5.1.2 Zn Recycling and Enrichment After smelting for 4–5 days in electric arc furnace, the dust collector is cleaned, and after all the dust is collected, the briquetting operation is carried out. The curing period of dust briquette and dust collection time should be synchronized as much as possible. The curing of cold briquette is completed in 4–5 days. The cold briquette is reused to electric arc furnace for smelting, and the consumption per ton of steel is 15–20 kg, so as to ensure that the generated dust can be completely reused. With every 10 days as a small cycle, the Zn content in dust is measured regularly.

5.1 Reuse and Zinc Enrichment Technology of Zinc-Containing Dust …

107

After 10–20 cycles (about 100–200 days), when the zinc content in the dust collected by the dust collector is more than 60%, the circulation is stopped. Rotary hearth furnace is used for recovery.

5.1.3 Separation of ZnCl2 by High Temperature Roasting The shortage of raw materials is becoming an increasingly prominent issue for the zinc smelting industry in China. Secondary resources, such as steelmaking dust, galvanizing dross, and blast furnace dust, are important raw material supplies. Because of circulation and accumulation effects, EAF steelmaking by galvanized steel and scrap parts produces a large amount of zinc-rich dust. If the steelmaking dust can be fully utilized, it will effectively alleviate the domestic shortage of raw materials. Steelmaking zinc-rich dust cannot be directly used for zinc smelting. The contained [Cl] element enters the electrolyte in ion form (Cl− ) and corrodes the system’s equipment, such as the positive plate. Eventually, the zinc hydrometallurgy process and the working environment are affected and they deteriorate. Dechlorination from the steelmaking zinc-rich dust is necessary. The roasting method is commonly used for defluorination and dechlorination, which gives obvious advantages of being a simple process and having good effects. In the process of roasting steelmaking zinc-rich dust, the synchronous loss of zinc is large. It is necessary to extensively study the internal reaction mechanism to reduce the actual [Zn] loss and achieve complete dechlorination. The reactions of dechlorination and [Zn] reclaiming in the roasting steelmaking zinc-rich dust process occur through a typical multi-step process, which results in morphology changes and the mass transport of reactants and products. In this chapter, the physical and chemical properties of steelmaking zinc-rich dust are confirmed firstly. Then, thermal experiments are designed to validate the key reaction characteristics. Finally, the reaction mechanism model of dechlorination and [Zn] reclaiming in the roasting steelmaking zinc-rich dust process is discussed and built.

5.1.3.1

Discussion About the Reaction Mechanism

The Physical and Chemical Properties of the Steelmaking Zinc-Rich Dust In steelmaking using scraps in an electric arc furnace or induction furnace, zinc and other elements are evaporated into the furnace gas at a high temperature, and then they condense with the oxygen in the furnace gas gradually. Finally, the combination product aggregates into fine particles, which is steelmaking dust. The steelmaking dust is commonly processed into auxiliary materials and reused in steelmaking production to recover [Fe]. [Zn] gradually circulates and accumulates, and

108

5 Disposal of Solid Waste in Steelmaking Process

the heavy use of galvanized scrap steel and scrap parts aggravates the enrichment. Finally, steelmaking zinc-rich dust forms. In this study, steelmaking zinc-rich dust was collected from a steelwork. A chemical composition analysis was carried out firstly, and the dust was shown to contain 63.8% zinc and 3.18% chlorine, as shown in Fig. 5.2a. Then, the crystal structure of the dust was determined by XRD, and the results (Fig. 5.2b) show that almost all zinc exists in the form of ZnO or ZnCl2 ; The content of ZnO is extremely high in the dust, and all chlorine is stored in ZnCl2 . Finally, the particle size was measured by a laser particle size distribution analyzer, and the steelmaking zinc-rich dust particles were all found to be very small, mainly in the range of 1–10 µm and mostly 2–5 µm. The statistics of the particles size are shown in Fig. 5.2c. Thermodynamics Analysis of Dechlorination and [Zn] Reclaiming ZnCl2 is volatile matter, whose melting point, boiling point and saturated vapor pressure at different high temperatures are shown in Table 5.1. The saturated vapor pressure of ZnCl2 changes by order of magnitude from temperatures of 550 to 650 °C, which means a greater gasification rate, and ZnCl2 boils at a temperature of 732 °C. In the roasting process, the ZnCl2 in the dust volatilizes by following the reaction equation shown below:

Fig. 5.2 Physical and chemical properties of the steelmaking zinc-rich dust

5.1 Reuse and Zinc Enrichment Technology of Zinc-Containing Dust …

109

Table 5.1 Physical property parameters of ZnCl2 material Phase transition temperature

Saturated vapor pressure

Melting Boiling 550 °C 318 °C

650 °C

750 °C

850 °C

950 °C

732 °C 4.38 × 103 Pa 3.15 × 104 Pa 1.46 × 105 Pa 4.93 × 105 Pa 1.32 × 106 Pa

ZnCl2 (s/l) = ZnCl2 (g) When the steelmaking zinc-rich dust is roasted under atmospheric pressure or negative pressure at temperatures above 650 °C, ZnCl2 will evaporate into the gas phase and be dechlorinated directly. As reported in reference, the method of roasting at the temperature of 700 °C for 3 h is commonly used for dechlorinating some [Zn] or [Pb] rich raw materials. In the process of ZnCl2 transformation into the gas phase, synchronous loss of zinc is unavoidable. ZnCl2 may react with O2 following the equation below in an oxygen-containing atmosphere such as air. In view of the phase change of ZnCl2 , the relationship between the Gibbs free energy of the chemical reaction and the temperature is shown in Fig. 5.3. 2ZnCl2 + O2 = 2ZnO + 2Cl2 According to the reaction thermodynamics data, when ZnCl2 is in the solid or liquid phase at temperatures below 732 °C, the Gibbs free energy is always above 0 kJ/mol, and no reaction occurs. At temperatures above the boiling point, the Gibbs free energy falls to below 0 kJ/mol. There is opportunity for ZnCl2 in the steelmaking zinc-rich dust to be volatilized into steam and then be oxidized into ZnO when the dust is roasted at above 732 °C in air atmosphere. If the two-step reaction occurs in an ideal position, efficient dechlorination and no loss of zinc elements will be achieved at the same time. Fig. 5.3 Gibbs free energy of the [Zn] reclaiming reaction under different standard conditions with a changing temperature

110

5 Disposal of Solid Waste in Steelmaking Process

Discussion on the Reaction Mechanism and Hypotheses The dust particles can be approximated as spheres. The unidirectional mass transfer process of ZnCl2 from the inner of the dust particle to the outside conforms to the shrinking unreacted core model with a constant particle size. ZnCl2 diffuses into the gas phase on the surface of the unreacted core, and nonvolatile matter forms the layer of residual porous media, which is almost all ZnO. ZnCl2 steam transmits through the layer of residual porous media first, and then diffuses into the gas phase. As the reaction progresses, the diameter of the unreacted core decreases, the thickness of the residual porous media increases, and the overall size of the dust particles remains the same, as shown in Fig. 5.4a. For the overall dechlorinating process of the accumulation layer with a large number of dust particles, the reaction time for dechlorination of a single dust particle can be ignored, and the dechlorinating process of the accumulation layer conforms to the shrinking unreacted core model too, as shown in Fig. 5.4b. Roasting steelmaking zinc-rich dust in the air atmosphere at a temperature above 732 °C causes mass transportation of ZnCl2 steam from the inner and oxygen from the gas environment in opposite directions. Affected by the two mass transportations and the chemical reaction, the reaction process of dechlorination and [Zn] reclaiming may be complicatedly changed and constrained by four influential factors, which are shown below: (1) The diffusion rate of O2 transmission through the boundary layer and the layer of residual porous media; (2) The diffusion rate of ZnCl2 steam transmission through the layer of residual porous media and the boundary layer; (3) The chemical reaction rate of ZnCl2 and O2 reacting with each other; (4) The diffusion rate of generated Cl2 transmission into the gas atmosphere. Sufficient reactants are required for the chemical reaction. In the roasting process, if the internal diffusion rate of O2 is higher and the external diffusion rate of ZnCl2 steam is lower, adequate reactant concentrations will be available in the internal slot of the residual porous media. On the contrary, if the internal diffusion rate of O2 is lower and the external diffusion rate of ZnCl2 steam is higher, the necessary reactant concentration may only be provided in the gas environment outside of the particles. In a general situation, the speed of the chemical reaction between gaseous reactants is especially fast, but the nucleation of the ZnO generated from the gas phase may be a restricted link, which is in the solid-phase at the reaction temperature. If the chemical reaction occurs in the internal slot of the residual porous media, because the ZnO existing in the dust particles is already the best nucleation agent, the chemical reaction will complete quickly, and ZnO will be reclaimed in the dust particles internally. If the chemical reaction occurs in the gas environment, the chemical reaction process may last for a long time, and the generated ZnO will leave away the original dust particles, which makes it difficult to reclaim zinc. In essence, the position at which the chemical reaction occurs is the key characteristic for the reaction mechanism of dechlorination and [Zn] reclaiming in the roasting

5.1 Reuse and Zinc Enrichment Technology of Zinc-Containing Dust …

111

Fig. 5.4 Shrinking unreacted core model of the dust dechlorination process with a constant particle size. a Unreacted nuclear model; b ZnCl2 stacked layer reaction model

steelmaking zinc-rich dust process. There are two possibilities in our assumptions: In Hypothesis A, O2 from atmospheric enters the residual porous media and reacts with ZnCl2 , and the generated ZnO is attached to the gaps in the layer of the residual porous media (as shown in Fig. 5.5a); in Hypothesis B, the ZnCl2 steam crosses the layer of residual porous media and diffuses into the gas phase, where it reacts with O2 to generate ZnO in the solid phase. Then, the solid ZnO settles onto the surface of the dust or dissipates into the environment (as shown in Fig. 5.5b). All the generated Cl2 gas discharges into the exhaust gas. To verify which hypothesis is consistent with reality, some thermal experiments are necessary.

112

5 Disposal of Solid Waste in Steelmaking Process

Fig. 5.5 Hypothetical models about the dechlorination and [Zn] reclaiming reactive processes. a Gas–solid reaction model; b Gas–gas reaction model

5.1.3.2

Experimental Schemes

According to the characteristics of the mechanism models, the atmospheric environment and the roasting temperature have different impacts on the reactive results. In this chapter, two groups of thermal experiments were designed to discuss and identify the reaction mechanisms of dechlorination and [Zn] reclaiming in the roasting steelmaking zinc-rich dust process. Experiments (a) Involving Roasting in Different Atmosphere Environments In experiment (a), dust in crucibles was heated and roasted under different gas atmospheres, and the results of dechlorination and [Zn] reclaiming were investigated after the roasting treatment. To prepare the experiment objects, every 100 g of dust was weighed and charged into each crucible, and the dust was flatly paved on the bottom of the crucible. The crucibles were made of alumina material and were cylindrical in shape with a diameter of 60 mm and a height of 75 mm. Under different atmospheres, argon or air was sprayed by a quartz tube or no air was sprayed for 60 min. The quartz tube was vertically inserted 2 cm below the upper edge of the crucible along the central axis of the crucible. During blowing, a forced gas flow formed in the free space of the crucible. The forced gas flow should not disturb the tiled dust, and the flow rate was confirmed as 100 mL/min. In terms of the experiment equipment, a fast heating muffle furnace of 1600 °C was prepared, and its semi-sealing and big free space guaranteed the long-term stability of the atmosphere. The experimental gas was supplied by high pressure vessels, and the flow rate was controlled by a pressure-reducing valve and gas flowmeter. The gas was carried through gas pipes, and a heat exchange pipe was equipped in the furnace hearth to ensure that the temperature of the blowing gas was consistent with the furnace. The equipment system for experiment (a) is shown in Fig. 5.6. The experiment operation was determined as described below: The muffle furnace was first heated up while empty, and the blowing gas started up too. When the

5.1 Reuse and Zinc Enrichment Technology of Zinc-Containing Dust …

113

Fig. 5.6 Experiment system for experiment (a) involving roasting in different atmosphere environments

temperature reached the set point, and the gas blowing flow was stabilized, the prepared dust with the crucible was put into the furnace, and after adjusting the position of the blowing pipe and closing the furnace door, the roasting experiment was timed. When the experiment was finished, the dust with the crucible was removed and covered, and the argon gas was blown for dust cooling during sealing. After the experiment, the mass of the dust was weighed, and the compositions of [Zn] and [Cl] in the remaining powder were measured too. To reduce the error, each experiment scheme was carried out three times on average. The specific experimental schemes are shown in Table 5.2.

114

5 Disposal of Solid Waste in Steelmaking Process

Table 5.2 Schemes for experiment (a) involving roasting in different atmosphere environments Items

a-1

a-2

a-3

Temperature, °C

900

900

900

Dust Weight, g

100

100

100

Roasting time, min

60

60

60

Atmosphere

Flowing argon

Stationary air

Flowing air

Gas composition

Ar

–

Air

Gas flow rate, mL/min

100

0

100

Gas consumes, mL

6000

–

6000

Repeat times

3

3

3

Experiment (b) Involving Roasting at Different Roasting Temperatures In experiment (b), the prepared dust was heated and roasted in air atmosphere at different temperatures, and the weight and compositions of the remaining powders were investigated after the roasting treatment too. The dust with crucibles were prepared in the same way and roasted at stationary air atmosphere for 60 min at temperatures of 750, 800, 850, 900, and 950 °C. The operation process was as follows: After the furnace temperature reached the experimental temperature, the prepared dust in crucibles were put into the furnace, and the timer was started. When the roasting time was over, the dust with the crucible was removed and covered, and the argon gas was blown for cooling while sealing. The dust was weighed and chemically examined. Every scheme was carried out three times on average. The equipment system and the experimental schemes for experiment (b) are shown in Fig. 5.7 and Table 5.3.

5.1.3.3

Experimental Results and Discussion

In accordance with the schemes of experiment (a), nine experimental dusts were roasted under three different gas environments, and all the residual powders were weighed and chemically examined. The experimental results are shown in Table 5.4, and the average calculation is shown for each scheme. Under the three different gas environments, the roasted dusts all dechlorinated completely, and the average dechlorination rate reached 98.51%. The gas environment had little influence on the dechlorination of the steelmaking zinc-rich dust. Based on the changes in weight and composition, the dechlorinated mass, the synchronous dezincification, and the actual [Zn] loss were calculated, and the difference between the synchronous dezincification and the actual [Zn] loss represented the mass of [Zn] reclaiming, and the proportion of the [Zn] reclaiming to the synchronous dezincification represented the reclaiming rate of [Zn]. All data are shown in Table 5.5.

5.1 Reuse and Zinc Enrichment Technology of Zinc-Containing Dust …

115

Fig. 5.7 Experiment system for experiment (b) involving at different roasting temperatures Table 5.3 Schemes for experiment (b) involving at different roasting temperatures Items

b-1

b-2

b-3

b-4

b-5

Temperature, °C

750

800

850

900

950

Dust mass, g

100

100

100

100

100

Period, min

60

60

60

60

60

Atmosphere

Stationary air

Repeat times

3

3

3

3

3

116

5 Disposal of Solid Waste in Steelmaking Process

Table 5.4 Results from experiments (a) involving roasting in different atmosphere environments Schemes Atmosphere Dust Residual Content Average Content Average Average weight, weight, of Zn, weight of Cl, weight dechlorination g g % of Zn, g % of Cl, g rate, % a-1

a-2

a-3

Flowing argon

Stationary air

Flowing air

100

88.1

69.19

100

88.0

69.27

60.958

0.108

100

87.8

69.42

0.108

100

88.5

69.04

100

88.2

69.29

100

87.9

69.39

100

88.0

69.18

100

88.2

69.02

0.011

100

87.8

69.34

0.011

61.069

0.107

0.043

0.095

97.03

0.038

98.8

0.010

99.7

0.043 0.043 60.880

0.011

Table 5.5 Calculation of dechlorination and [Zn] reclaiming in experiment (a) involving in different atmosphere environments Schemes

Atmosphere

Dechlorination, g

Synchronous dezincification, g

Actual [Zn] loss, g

[Zn] reclaiming, g

Rate of [Zn] reclaiming, %

a-1

Flowing argon

3.086

2.842

2.842

0.000

0

a-2

Stationary air

3.142

2.894

2.699

0.195

6.74

a-3

Flowing air

3.170

2.920

2.903

0.017

0.58

In the experiment of scheme (a-1), argon gas was used to isolate oxygen from the air and accelerate the external diffusion of ZnCl2 , and the chemical reaction of ZnCl2 steam in the crucible was eliminated. As a result, all [Zn] in the ZnCl2 steam was lost (2.843 g), and the mass of [Zn] reclaiming was null. In the experiment of scheme (a-2), the experimental condition was consistent with general roasting in an air environment. With the volatilization of ZnCl2 steam, the dechlorination mass and synchronous dezincification are 3.142 g and 2.894 g respectively, and the mass of [Zn] reclaiming is 0.195 g, which represented 6.74%. In the experiment of scheme [a-3], flowing air accelerates the external diffusion of ZnCl2 , and gives little impact on the internal diffusion. The dechlorination and synchronous dezincification masses were 3.170 g and 2.920 g respectively, and the [Zn] reclaiming mass was only 0.017 g, which represented 0.58%. The experiment (a) results show that in the three different atmosphere environments, the actual measured masses of [Zn] reclaiming were all very low. The ZnO generated in the gas phase was affected by the gas flow in the crucible, which accelerates the external diffusion of ZnCl2 steam and is hard to deposit in the crucible. Comparing the experiment results of scheme (a-2) and scheme (a-3), which represent deposition or none, respectively, the mass of [Zn] reclaiming decreased from

5.1 Reuse and Zinc Enrichment Technology of Zinc-Containing Dust …

117

0.195 to 0.017 g. According to the most conservative estimation, for the confirmed [Zn] reclaiming reaction, 91.28% of the chemical reaction occurs in the gas environment, compared with 8.72% in the internal gaps of the residual porous media layer. Considering the total mass of zinc that needs to be reclaimed, only 0.58% occurs in the internal gaps of the residual porous media layer. The reaction location of ZnCl2 reclaiming to generate ZnO mainly occurs in the gas phase, and the reaction characteristics of roasting dechlorination and the [Zn] reclaiming process are consistent with the Hypothesis B. As verification, the dust was roasted at different temperatures in experiment (b). A total of 15 roasting experiments were conducted under five schemes, and the roasted dust was weighed and assayed. All experimental results are shown in Table 5.6. With an increase in roasting temperature, the dechlorination rate rose gradually, and dechlorination was almost complete at temperatures over 800 °C. The results of dechlorination and [Zn] reclaiming at different roasting temperatures are shown in Table 5.7. As the results show, the mass of [Zn] reclaiming slightly increased first and then gradually decreased with an increase of the roasting temperature. The [Zn] reclaiming rate was more representative of the [Zn] reclaiming reaction degree, which decreased with an increase in the reaction temperature, and the [Zn] loss significantly increased (see Fig. 5.8). According to thermodynamic data, the Gibbs free energy of the [Zn] reclaiming reaction decreases as the reaction temperature increases, which means that the trend of [Zn] reclaiming and the results of the thermal state experiment are contrary to the Table 5.6 Results from experiments (b) involving at different roasting temperatures Schemes Temperature, Dust Residual Content Average Content Average Average °C weight, weight, of Zn, weight of Cl, weight dechlorination g g % of Zn, g % of Cl, g rate, % b-1

b-2

b-3

b-4

b-5

750

800

850

900

950

100

91.2

68.54

100

91.0

68.71

62.51

0.61

100

91.4

68.36

100

89.0

69.89

100

89.1

68.89

100

89.0

69.89

100

90.0

68.12

100

89.0

68.89

100

89.0

68.89

100

88.5

69.04

100

88.2

69.29

100

87.9

69.39

100

87.2

68.86

100

86.6

70.32

0.057

100

87.1

69.91

0.057

0.556

82.5

0.098

96.9

0.067

97.9

0.038

98.8

0.050

98.4

0.61 0.61 62.204

0.11 0.11 0.11

61.312

0.074 0.075 0.075

61.069

0.043 0.043 0.043

60.923

0.057

118

5 Disposal of Solid Waste in Steelmaking Process

Table 5.7 Calculation of dechlorination and [Zn] reclaiming in experiment (b) involving at different roasting temperatures Schemes

Temperature, °C

Dechlorination, g

Synchronous dezincification, g

Actual [Zn] loss, g

[Zn] reclaiming, g

Rate of [Zn] reclaiming, %

b-1

750

2.624

2.416

1.295

1.121

46.40

b-2

800

3.081

2.838

1.569

1.269

44.71

b-3

850

3.113

2.867

2.488

0.379

13.22

b-4

900

3.142

2.894

2.699

0.195

6.74

b-5

950

3.129

2.882

2.877

0.005

0.17

Fig. 5.8 Results of dechlorination and [Zn] reclaiming with a changing temperature

thermodynamic results. The chemical reaction process is suppressed by the reaction kinetic conditions. According to the Hypothesis A, the ZnCl2 reclaiming reaction occurs in the layer of residual porous media in which the main constituent is solid ZnO; thus, there is no nucleation problem in the formation of ZnO. The results of experiment (b) contradict with the Hypothesis A. According to the Hypothesis B that ZnCl2 steam reacts with O2 in the gas phase outside of the dust particles, the difficulty of producing a new nucleation ZnO solid phase from gas phase increases with the environmental temperature, and the chemical reaction speed is suppressed. Finally, the ZnO generated and dropped in the crucible decreases. The results of experiment (b) verify the Hypothesis B.

5.1 Reuse and Zinc Enrichment Technology of Zinc-Containing Dust …

5.1.3.4

119

Mechanism Model of Dechlorination and [Zn] Reclaiming

Based on the results of the thermal experiments, the main reaction processes of dechlorination and [Zn] reclaiming are clear: ZnCl2 is transformed into the gas phase inside the dust particles and spreads out and then diffuses into the surrounding gas phase, where it reacts with O2 and ZnO is generated. The reaction mechanism mode is shown in Fig. 5.9. The reaction of dechlorination and [Zn] reclaiming is a multi-step process, and the reaction steps are as follows: (1) When the dust is heated, ZnCl2 boils and becomes steam on the surface of unreacted nuclear matter, and the nonvolatile components form a residual porous media layer at the same time. (2) ZnCl2 steam spreads from the inner part of the dust to the outside. It transmits through the residual porous media layer and the gas boundary layer and, in turn, diffuses into the gas phase. (3) O2 from the air is transmitted to the vicinity of the dust particles. (4) ZnCl2 and O2 encounter and react with each other, and ZnO and Cl2 are generated. (5) The generated ZnO nucleates and grows into solid particles. Under the action of gravity, some ZnO solid particles settle on the surface of dust particles, while other ZnO solid particles remain loose when they are far away from the dust and the crucible. (6) The generated Cl2 diffuses into the gas phase and becomes loose.

Fig. 5.9 Reaction mechanism model of dechlorination and [Zn] reclaiming in the roasting steelmaking zinc-rich dust process

120

5 Disposal of Solid Waste in Steelmaking Process

According to the reaction mechanism model, the concentration of ZnCl2 steam in the dust particles decreases as the distance from the ZnCl2 boiling interface increases, and it is low in a gas environment, and the reaction may be limited by the dynamics of the new nucleation of ZnO solid. ZnCl2 steam spreads a long distance without reacting, and the generated ZnO from the slow chemical reaction appears in the area, which is far away from the dust and the crucible. It cannot be collected efficiently, resulting in little [Zn] being reclaimed. If the concentration of O2 is higher in the residual porous media layer, the chemical reaction of [Zn] reclaiming must be unstoppable, and the process of ZnCl2 spreading is blocked, which means more [Zn] is reclaimed. The key reason for the chemical reaction occurring in the external gas environment outside of the dust is that O2 from the air environment cannot be transmitted into the dust particle interior or dust bed efficiently. Accelerated transmission of O2 can change the reaction area and cause more [Zn] reclaiming to occur in the roasting steelmaking zinc-rich dust process.

5.1.3.5

Conclusions

(1) The reaction mechanism model of dechlorination and [Zn] reclaiming from steelmaking zinc-rich dust is shown in Fig. 5.9, and the reaction is a multi-step process, as follows: ➀ When the dust is heated, ZnCl2 boils and becomes steam on the surface of unreacted nuclear matter, and the nonvolatile components form a residual porous media layer at the same time. ➁ ZnCl2 steam spreads from the inner part of the dust to the outside. It transmits through the residual porous media layer and the gas boundary layer and, in turn, diffuses into the gas phase. ➂ O2 from the air transmits to the vicinity of the dust particles. ➃ ZnCl2 and O2 encounter and react with each other, and ZnO and Cl2 are generated. ➄ The generated ZnO nucleates and grows into solid particles. Under the action of gravity, some ZnO solid particles settle on the surfaces of dust particles, while other ZnO solid particles remain loose when they are far away from the dust and the crucible. ➅ The generated Cl2 diffuses into the gas phase and becomes loose. (2) The key reason that the chemical reaction occurs in the external gas environment outside of the dust is that O2 from the air environment cannot be transmitted into the dust particle interior or dust bed effectively. Accelerated transmission of O2 can lead to the solution undergoing efficient [Zn] reclaiming while dechlorinating from the steelmaking zinc-rich dust. (3) Restricted by the reaction mechanism, the reaction rate of [Zn] recalcining is limited, and lots of zinc is lost synchronously without being recovered, so the actual measured mass of [Zn] reclaiming is very low. The reaction of [Zn] recalcining should be limited by the dynamics of new nucleation of ZnO solid, which is more difficult at higher temperatures, although the Gibbs free energy of the [Zn] recalcining reaction is lower.

5.2 Direct Reduction of Red Mud for Refining Desulfurization

121

5.2 Direct Reduction of Red Mud for Refining Desulfurization Red mud in aluminium industry belongs to the bulk high pollution solid waste. More than 30 million tons of high-iron red mud are produced annually in China, which has huge environmental and safety hazards. In this study, the high-iron red mud is regarded as a resource rich in Fe, Al and Na, and the basic research on the recycling process of high-iron red mud is carried out, so as to realize the full resource utilization of high-iron red mud under the conditions of no pollution, high efficiency, low cost and high added value. In this section, sodium removal and deep reduction of red mud are realized on a pilot scale, and the possibility of further use of red mud in refining desulfurization slag is discussed through laboratory studies.

5.2.1 Simple Method for Na Removal by Red Mud Pretreatment Under normal temperature, the red mud with high iron was treated by CO2 , which can react with caustic soda, sodium aluminate and sodium silicate soluble in water to form NaHCO3 , and most alkaline substances can be removed. In the process of sodium removal from red mud, the reaction between sodiumalkali substance and CO2 gas is a complex reaction process involving solid–liquid-gas three phases. Figure 5.10 is a schematic diagram of sodium removal kinetics of red mud. The process can be described as follows: First of all, CO2 gas diffuses from bubbles to the gas–liquid reaction interface and reacts with water to form carbonic acid, and bicarbonate ions and hydrogen ions are dissociated from carbonic acid in aqueous solution; sodium-alkali substances in red mud diffuse from red mud particles to solid–liquid reaction interface and dissociate sodium ions and hydroxide ions by water reaction; sodium ions and carbonate ions combine to form sodium bicarbonate in red mud, and hydrogen ions and hydroxide ions combine to form water. There is a dynamic equilibrium among all chemical substances.

Fig. 5.10 Schematic diagram of sodium removal kinetics of red mud

122

5 Disposal of Solid Waste in Steelmaking Process

Chemical reactions are often considered to be completed in a moment, and the limiting factor of sodium removal kinetics is mainly the diffusion time of sodiumalkali substances and CO2 gas. In the experiment, CO2 gas flow far exceeds the reaction consumption, the dissolution of CO2 in red mud is saturated. Further increase of CO2 gas flow rate can not significantly improve the sodium removal rate, so the influence of CO2 dissolution on kinetics can be ignored. It can be seen from the experiment that the red mud particle size has a significant effect on the sodium removal rate, and the red mud particle size directly affects the diffusion of sodiumalkali substances, which is the main limiting factor affecting the reaction. Red mud has small particle size and large specific surface area, which is beneficial to the diffusion of sodium-alkali substances and promotes the reaction. Previous laboratory studies have shown that the only significant factor affecting the sodium removal rate of red mud is the particle size of red mud, while other factors have little influence and can be almost ignored. Therefore, the simplest treatment method can be considered for CO2 sodium removal of high-iron red mud. The deep sodium removal of red mud was carried out by introducing CO2 gas into a hydrothermal reactor, and the solid-to-liquid ratio could be 1/4 in the process of sodium removal. Under the condition of large-scale preparation, the sodium removal time was controlled within 40 min. According to the experimental scheme, the sodium removal experiment of red mud was carried out, and the red mud after sodium removal was obtained. The red mud after sodium removal was sampled, and the samples were detected and analyzed by XRF fluorescence analysis, and the chemical components of red mud were obtained. 4 samples of red mud after sodium removal were obtained in the experiment, and the main components and sodium removal rate of each red mud after sodium removal were listed in Table 5.8 by XRF detection. It can be seen from Table 5.8 that the sodium removal rate of red mud can reach more than 72.5% by adding 4 times of water and introducing sufficient CO2 gas for 40 min at normal temperature. Considering that the process of adding stirring equipment in the actual production process is complicated, and the improvement of sodium removal rate is very limited, stirring treatment is not added in the actual production process. XRD energy spectrum analysis and SEM scanning electron microscope analysis were carried out on red mud after sodium removal from Shandong Aluminum Plant. The XRD energy spectrum analysis results of fine-grained red mud after sodium Table 5.8 Experimental results (wt%) Experiments Particle Components of red mud after sodium removal Sodium size of Fe2 O3 Al2 O3 SiO2 TiO2 Na2 O CaO MnO Cr2 O3 MgO removal red mud rate, % 1 2

Fine 63.63 grinding 63.67

20.16

7.27

3.17

2.89

0.74 0.33

0.23

0.17

72.87

20.18

7.26

3.16

2.84

0.72 0.38

0.22

0.19

72.78

5.2 Direct Reduction of Red Mud for Refining Desulfurization

123

removal in Shandong Aluminum Plant are shown in Fig. 5.11. SEM photos of red mud after sodium removal under different fields of view are shown in Fig. 5.12. The XRD results of red mud after sodium removal show that Fe mainly exists in the form of Fe2 O3 (hematite) and FeO(OH) (goethite); Na mainly exists in the form of NaAlSiO4 ; Aluminum mainly exists in the form of gibbsite. Silicon mainly exists in the form of SiO2 . It can be found that there is no NaAlO2 in Na element composition, which indicates that it has been removed during sodium removal, but NaAlSiO4 has not been removed. The pH value of red mud during sodium removal with CO2 was recorded (Fig. 5.13). During the first 20 min of sodium removal, the pH value of red mud with 2 particle sizes decreased rapidly, and then decreased slowly, while the pH value of fine-grained red mud was lower. Sodium-alkali substances in fine-grained red mud are easier to leach, and the reaction rate between sodium-alkali substances and CO2 is faster. The pH value of coarse-grained red mud fluctuates around 7.1 after

Fig. 5.11 XRD analysis results of red mud after sodium removal

Fig. 5.12 SEM photos of red mud after sodium removal under different fields of view

124

5 Disposal of Solid Waste in Steelmaking Process

Fig. 5.13 Change in pH value of red mud

sodium removal time for 30 min, while the pH value of fine-grained red mud reaches a plateau after sodium removal time for 40 min. Coarse-grained red mud has larger particles, and the leaching amount of sodium-alkali substances is very small after the reaction time reaches 30 min, while the overall leaching amount of sodium-alkali substances in fine-grained red mud with smaller particles is relatively large. The red mud was treated with CO2 for sodium removal, and sodium-alkali substances in the red mud reacted with CO2 to form NaHCO3 . According to the determined process conditions and results of sodium removal, the NaHCO3 solution with a concentration of 3.86 g/100 mL, which is 4 times higher than that of red mud, can be obtained for each certain amount of red mud. The concentration of 3.86 g/100 mL NaHCO3 solution is still low for direct recovery of solid NaHCO3 , which is not economically feasible. The NaHCO3 solution can be used to remove sodium again instead of water, and the NaHCO3 solution can be increased to more than 6%, and then the solid NaHCO3 can be recovered by distillation or membrane method. It can be concluded that the pH value of red mud decreases from 10.2 to 7.1 (almost neutral) when the red mud with 1/4 solid–liquid ratio (average particle size 1.236 µm) is treated by sodium removal with CO2 for 40 min at room temperature. After treatment, the soluble sodium and alkali in the red mud were basically transformed into NaHCO3 , and the red mud after sodium removal and NaHCO3 solution were obtained by separating the red mud. Residual sodium in red mud after sodium removal exists in the form of sodium aluminosilicate insoluble in water, which does not pose a potential pollution threat to water and land resources. NaHCO3 solution can be further studied for enrichment and recovery of solid NaHCO3 .

5.2 Direct Reduction of Red Mud for Refining Desulfurization

125

The harmless treatment process of sodium removal with CO2 for high-iron red mud is simple and easy to operate, and the source of greenhouse gas CO2 is very wide. The high-iron red mud is harmless by this technology, and at the same time, the emission of greenhouse gas CO2 is reduced, and NaHCO3 can be further recovered, thus realizing the treatment of waste with waste. Therefore, it is beneficial to the recycling of resources.

5.2.2 Deep Reduction of Red Mud After Sodium Removal Red mud after sodium removal is deeply reduced in rotary hearth furnace with reducing agent. The molar ratio of C/O is 1.8, and the reduction temperature in high temperature zone is 1100 °C ± 50 °C. After mixing red mud and reducing agent in proportion, pellets were prepared by rolling pelletizing machine, and then put into rotary hearth furnace for deep roasting reduction. The flow ratio of natural gas and air should be controlled at CH4 /O2 = 7/ 4 to ensure the reducing atmosphere in the furnace. The reduction time is 2–4 h. The remaining operations are described above. 4 pilot scale tests were carried out in the experiment, and the experimental results are shown in Table 5.9. The reduction time has little effect on the reduction degree. In the experiment, the reduction degree of 4 h was less than 1% higher than that of 2 h. Long reduction time is obviously beneficial to promote the reaction, but the reduction time in experiments is longer, so the contrast effect is not obvious enough. The reduction products of red mud after sodium removal were analyzed by XRD, particle size analysis and SEM, and the basic characteristics of the reduction products were confirmed. Figure 5.14 shows the XRD analysis results of the reduction products. Because the reduction degree of red mud after sodium removal is extremely high at 1000 °C, almost all the iron elements in the reduction products exist in the form of metallic iron, so only the energy spectrum of iron elements is shown in the XRD energy spectrum. The particle size distribution of the reduction products of red mud after sodium removal is shown in Fig. 5.15. It can be found that the particle size has hardly changed before and after reduction, indicating that no sintering occurred in the reduction process. Table 5.9 Deep reduction and metallization rate Number of furnace times

Reduction time, h

Metallization rate, %

1

2

98.981

2

3

97.186

3

4

99.138

4

4

99.225

126

5 Disposal of Solid Waste in Steelmaking Process

Fig. 5.14 XRD energy spectrum of reduction products of red mud after sodium removal

Fig. 5.15 Particle size distribution of reduction products of red mud after sodium removal

The reduction products were analyzed by SEM scanning electron microscope to obtain the morphology of the reduction products of red mud after sodium removal under different fields of view (Fig. 5.16).

Fig. 5.16 Topography of reduction products of red mud after sodium removal under different fields of view

5.2 Direct Reduction of Red Mud for Refining Desulfurization

127

It can be seen from Fig. 5.16 that after reduction, the particles of red mud after sodium removal change from spherical and cylindrical to irregular briquettes, and the adhesion between particles is aggravated. From the particle size analysis, it is known that there is no sintering phenomenon between particles, so it is considered that the particles are physically bonded.

5.2.3 Desulfurization Design of Steelmaking Refining Slag Prepared from Reduction Products After the reduction of red mud after sodium removal, almost all iron elements exist in the form of metallic iron, and the reduction products are mainly composed of metallic iron and alumina. However, the magnetic separation results of the reduction products show that the metal iron and other components in the reduction products can not be effectively separated. The metallic iron in the reduction product of red mud after sodium removal is considered as iron resource, and other oxides mainly composed of alumina can be used as flux of refining slag. It is assumed that refining slag can be prepared by adding a certain proportion of CaO. When the slag is used for desulfurization in steelmaking, the metallic iron in the reduction product enters liquid steel, and the flux such as alumina and CaO form refining slag, which can effectively separate the metallic iron from other oxides in the reduction product, and finally the reduction product of red mud after sodium removal can be fully utilized as resources. The composition of common steelmaking refining slag is shown in Table 5.10, the main components of reduction products of red mud after sodium removal are listed in Table 5.11, and the main components of oxides except iron in reduction products are listed in Table 5.12. From Table 5.12, it can be seen that alumina is the main oxide component in the reduction product of red mud after sodium removal in Shandong Aluminum Plant. After adding CaO with the same quality as this oxide, the composition is shown in Table 5.13. Comparing the components in Tables 5.10 and 5.13, it can be seen that the composition of oxides in the reduction products of red mud after sodium removal after Table 5.10 Main components of common refining slag (wt%) Composition

CaO

Al2 O3

SiO2

FeO + MnO

Content

45–55

25–40

< 20

1.5 and Al2 O3 content of 25–35%, TiO2 content exceeding 1% can significantly increase the melting point of slag system. Dou Zhihe et al. studied CaO–Al2 O3 –TiO2 –CaF2 slag system and found that when TiO2 content is lower than 15%, the increase of TiO2 is beneficial to reduce the viscosity of slag. Li Youqi et al. prepared refining desulfurizer with TiO2 content of 1.86% from titanium-containing blast furnace slag. It can be seen that there is no systematic study on CaO–Al2 O3 –SiO2 –TiO2 slag system, and there are different viewpoints among related studies. In order to determine the effect of TiO2 on common refining slag, the melting performance and desulfurization ability of CaO–Al2 O3 –SiO2 –TiO2 slag system based on CaO–Al2 O3 –SiO2 –TiO2 slag system were systematically studied. It is expected that oxides in the reduction products of red mud after sodium removal can be used as flux for refining slag, which is beneficial to the resource reuse of high-iron red mud. The steelmaking temperature is generally around 1600 °C, and the liquid phase region corresponding to slag composition is 1500 °C. The area of liquid phase region at 1500 and 1600 °C under different TiO2 content (0–7%) is drawn as Fig. 5.17 by FactSage thermodynamic software. As can be seen from Fig. 5.17, under the same temperature conditions, the TiO2 content is high and the liquid phase area is large. FactSage data can be used to plot the change of liquid area in the liquid area of CaO–Al2 O3 –SiO2 –TiO2 quaternary slag system at 1500 and 1600 °C. At 1500 °C, the liquid phase region of CaO–Al2 O3 – SiO2 –TiO2 quaternary slag system can be divided into two parts: high alkalinity

5.2 Direct Reduction of Red Mud for Refining Desulfurization

129

Fig. 5.17 The effect of TiO2 Content (0–7%) on liquid phase area of CaO–Al2 O3 –SiO2 –TiO2 slag system

liquid phase region and low alkalinity liquid phase region. The change of liquid phase area of the two parts can be seen from Fig. 5.18. Under the condition of TiO2 content less than 3%, the area ratio of liquid phase in the high basicity region is basically unchanged, and the distribution of liquid phase in this part is similar to that of CaO–Al2 O3 –SiO2 slag system; when TiO2 content

130

5 Disposal of Solid Waste in Steelmaking Process

Fig. 5.18 The effect of TiO2 on liquid phase area of CaO–Al2 O3 –SiO2 –TiO2 slag system at 1500 °C

is higher than 3%, the area of liquid phase region increases significantly. High TiO2 content corresponds to a large area of low alkalinity liquid phase, so the total area of liquid phase is increased. When TiO2 content is 5% and 10%, the total area of liquid phase is 26.8% and 36.48% respectively, which is 3.44% and 13.12% higher than that of 23.6% in CaO–Al2 O3 –SiO2 slag system. When the content of TiO2 is less than 7%, the higher the content of TiO2 , the larger the area of liquid phase of CaO–Al2 O3 –SiO2 –TiO2 slag system. It can be seen that the addition of a small amount of TiO2 does not significantly reduce the melting performance of slag system, but increases the area of low melting point area to a certain extent, which can meet the melting point requirements of refining slag.

5.2.4 Calculation of Desulfurization Ability of CaO–Al2 O3 –SiO2 –TiO2 Slag System 5.2.4.1

Evaluation Index of Desulfurization Ability of Steel Slag

Desulfurization ability of steel slag is the key to achieve low sulfur content in liquid steel. The sulfur distribution coefficient L S among steel slag and sulfur capacity C S of slag are often used to evaluate desulfurization ability of slag. The sulfur distribution coefficient LS is the mass percentage ratio of sulfur in slag to sulfur in steel, namely: LS =

(%S) [%S]

(5.1)

5.2 Direct Reduction of Red Mud for Refining Desulfurization

131

where L S is the distribution coefficient of sulfur, %; (%S) is the mass percentage of sulfur content in the slag, %; [%S] is the mass percent of sulfur content in steel, %. According to the definition of sulfur capacity obtained from steel-slag reaction: CS =

K  · αO2− α[O] = (%S) rS2− α[S]

(5.2)

where CS is sulfur capacity obtained from steel-slag reaction; (%S) is sulfur content in slag, %; α[O] is oxygen activity in molten metal, %; α[S] is sulfur activity in molten metal, %.

5.2.4.2

Sulfur Capacity of CaO–Al2 O3 –SiO2 –TiO2 Slag System

The sulfur capacity of slag is often calculated by the optical basicity of slag. The ratio of the oxygen-emitting electron properties of some oxides in the spectral line with frequency to the oxygen-emitting electron properties of CaO is the optical basicity, which is related to Pauling’s electronegativity. The relationship between the two is as shown in formula (5.3). According to this formula, the optical basicity can be calculated as the theoretical optical basicity.  =

i X A X Oi · A X Oi  X A X Oi

(5.3)

where i is the number of oxygen atoms in oxide A X Oi ; X is the number of atoms of element in oxide A X Oi ; A X Oi is the optical basicity of pure oxides; X A X Oi is the number of moles of pure oxide A X Oi . In 1992, Young and Duffy obtained the relationship between sulfur capacity and optical basicity according to a large number of experimental data, which can be divided into the relationship between C S and  when the optical basicity of slag is greater than 0.8 and less than 0.8. When the optical basicity  of slag is less than 0.8, 11710 T − 0.02223(%SiO2 ) − 0.02275(%Al2 O3 )

lg CS = −13.193 + 42.84 − 23.822 −

(5.4)

When the optical basicity  of slag is more than ≥ 0.8, lg CS = −0.6261 + 0.4808 + 0.71972 + −

2587 + 0.0005144(%FeO) T

1697 T (5.5)

132

5 Disposal of Solid Waste in Steelmaking Process

Fig. 5.19 Equal sulfur capacity diagram of CaO–Al2 O3 –SiO2 –TiO2 quaternary slag system (wt%)

The sulfur capacity of CaO–Al2 O3 –SiO2 –TiO2 slag system with different TiO2 content (the area with CaO content of 35–65%) was calculated by the above method, and the calculation results were plotted to obtain the equal sulfur capacity Fig. 5.19. From Fig. 5.19, we can find that: (1) The area of slag system with sulfur capacity higher than 0.3 becomes smaller and the area of slag system with sulfur capacity lower than 0.3 becomes larger with high TiO2 content; (2) The equal sulfur capacity diagram of the quaternary slag system with 1% TiO2 content is very similar to that of the CaO–Al2 O3 –SiO2 ternary slag system, and the equal sulfur capacity diagram of the quaternary slag system with 3 and 5% TiO2 content is very similar. (3) The sulfur capacity of common refining slag ranges from 0.1 to 0.2. Comparing the sulfur capacity in the four-element slag system corresponding to the common slag range in each diagram, it can be found that the change of TiO2 content has little influence on the sulfur capacity of slag in this range.

5.2 Direct Reduction of Red Mud for Refining Desulfurization

133

It can be concluded that for CaO–Al2 O3 –SiO2 –TiO2 quaternary slag system, high TiO2 content will reduce the area of high sulfur capacity area with sulfur capacity higher than 0.3, but the influence on the area corresponding to common refining slag is small and almost negligible.

5.2.4.3

Sulfur Partition Coefficient of CaO–Al2 O3 –SiO2 –TiO2 Slag System

Equilib module of FactSage thermodynamic software can be used to simulate the reaction equilibrium of steel slag. By inputting the set mass and composition of liquid steel and slag, the mass and composition of slag and liquid steel under the condition of thermodynamic equilibrium can be calculated. According to the calculation results, the sulfur content in slag and steel under thermodynamic equilibrium state of each liquid steel and slag composition can be obtained, and then the sulfur distribution coefficient L S between slag and steel can be calculated by using formula (5.1). At present, HRB400 thread steel is used as the initial liquid steel composition, the composition is shown in Table 5.14, and CaO–Al2 O3 –SiO2 –TiO2 quaternary slag system is used as refining slag for desulfurization. When the mass ratio of slag to steel is set to 1:100, the slag of each component in CaO–Al2 O3 –SiO2 –TiO2 slag system and HRB400 thread steel liquid steel in the area with CaO content of 35–65% under different TiO2 content conditions are simulated and calculated. According to the calculation results, each L S was obtained, and the equal sulfur distribution coefficient between slag and steel was drawn (Fig. 5.20). Figure 5.20 shows that: (1) The area of high sulfur partition coefficient area in CaO–Al2 O3 –SiO2 –TiO2 quaternary slag system with sulfur partition coefficient higher than 200 decreases with high TiO2 content, but the change is very small. (2) The sulfur partition coefficient of slag composition in the corresponding diagram of common refining slag is about 100. Compared with different TiO2 content in this area, it can be seen that the change of TiO2 content has little influence on sulfur partition coefficient, which is almost negligible. In a word, the desulfurization ability of CaO–Al2 O3 –SiO2 –TiO2 quaternary slag has a tendency to decrease with the increase of TiO2 , but the increase of TiO2 content within 5% has little effect on the desulfurization ability of slag, which can be ignored. If a certain proportion of CaO is added to the reduction product of red mud after sodium removal, the slag composition composed of oxide and CaO is shown in Table 5.15. The blending ratio of CaO content is the mass percentage of CaO content Table 5.14 Composition of liquid steel (wt%) Elements in liquid steel

C

Si

Mn

P

S

Content

0.25

0.65

1.55

0.025

0.026

134

5 Disposal of Solid Waste in Steelmaking Process

Fig. 5.20 Equal sulfur distribution coefficient diagram of CaO–Al2 O3 –SiO2 –TiO2 quaternary slag system (wt%)

in the reduction product of red mud after sodium removal to the amount of refining slag prepared by this method, and the metal iron in the reduction product is not included in the amount of refining slag prepared. Table 5.15 Main components of oxides in reduction products of red mud after sodium removal after adding CaO Composition

CaO, %

Al2 O3 , SiO2 , TiO2 , Na2 O, Other, Binary Mannesmann % % % % % basicity, index, MI R2

CaO 40 blending 45 ratio, % 50

41.33 34.90

12.37 5.31

4.16

1.96

3.34

0.10

45.08 32.00

11.34 4.87

3.81

1.76

3.98

0.12

50.08 29.09

10.31 4.42

3.47

1.60

4.86

0.17

55

55.07 26.18

9.28 3.98

3.12

1.44

5.93

0.23

60

60.06 23.27

8.24 3.54

2.77

1.28

7.29

0.31

5.2 Direct Reduction of Red Mud for Refining Desulfurization

135

From the data in Table 5.15, it can be seen that the composition of refining slag prepared with CaO blending ratio of 50–55% meets the requirements of the experimental optimal slag composition range in Table 5.15, and the TiO2 content after adding CaO is within 5%, which is expected to be used as refining slag for desulfurization treatment. Viscosity of Refining Slag Prepared from Reduction Products Viscosity module of FactSage software can be used to calculate the viscosity of slag at various temperatures and compositions. The viscosity of slag can be obtained by using viscosity module of FactSage software to calculate the slag components formed by adding different proportions of CaO to oxides in the reduction products of red mud after sodium removal. Figure 5.21 shows the change of CaO ratio on the viscosity of refining slag prepared at 1500 and 1600 °C. As can be seen from Fig. 5.21, with the increase of CaO ratio, the viscosity of refining slag prepared by reducing products of red mud after sodium removal decreased at 1500 and 1600 °C. When CaO content is in the range of 50–55%, the viscosity of slag is about 0.1 Pa s at the steelmaking temperature of 1600 °C, which can meet the requirements of refining slag for steelmaking. Melting Point and Sulfur Partition Coefficient of Refining Slag Prepared from Reduction Products The melting point of slag can be calculated by using Phase diagram module of Factsgae software. This module is used to calculate the melting point of refining slag prepared by different proportions of CaO. Equilib module of Factsgae software can be used to calculate the thermodynamic equilibrium state of steel slag reaction, and obtain the composition of slag and liquid steel in equilibrium state, so as to calculate the sulfur distribution coefficient. This module can be used to calculate the sulfur partition coefficient of refining slag prepared with different CaO ratios. Set the mass ratio of slag to steel as 1:100, and the composition of liquid steel shall follow the Fig. 5.21 The effect of CaO blending ratio on viscosity of prepared refining slag

136

5 Disposal of Solid Waste in Steelmaking Process

Fig. 5.22 The effect of CaO blending ratio on melting point and sulfur partition coefficient of refining slag

composition in Table 5.14. The influence of CaO ratio on the melting point and sulfur partition coefficient of the prepared refining slag can be obtained by sorting out the calculation results as shown in Fig. 5.22. Figure 5.22 shows that: (1) The melting point of the slag is about 1330 °C, which is far lower than the requirement of refining slag for steelmaking, when the ratio of CaO is between 49 and 55%. However, when the content of CaO is higher than 55%, the melting point of slag increases rapidly, and when the content of CaO is higher than 58%, solid CaO precipitates and cannot be melted within 1600 °C. Therefore, the proportion of CaO should be 50–55%. (2) The sulfur partition coefficient of refining slag increases with the increase of CaO ratio. However, when the content of CaO is higher than 59%, the sulfur partition coefficient of slag remains unchanged because the precipitated CaO cannot be melted at steelmaking temperature. When CaO content is in the range of 50–55%, the sulfur partition coefficient of slag is in the range of 200–600, which can meet the requirements of steel-making refining slag. It can be concluded from the above theoretical calculation that the refining slag prepared under the condition of 55% CaO ratio has low melting point, low viscosity, high sulfur distribution coefficient and strong desulfurization ability. Therefore, it is theoretically feasible to prepare refining slag for steelmaking desulfurization by adding CaO to the reduction product of red mud after sodium removal.

5.2 Direct Reduction of Red Mud for Refining Desulfurization

137

5.2.5 Desulfurization Experiment of Refining Slag Prepared from Red Mud Reduction Product 5.2.5.1

Experimental Arrangement

Experimental Materials (1) Experimental steel: The experimental steel is HRB400 thread steel, the specific composition is shown in Table 5.14, and the amount of steel used in each experiment is about 93 g. (2) Experimental slag: The reduction product with a certain proportion of CaO, which is equivalent to 10% of the experimental steel. Experimental Equipment The equipment used in the experiment is a tubular resistance furnace produced by Baotou Yunjie Electric Furnace Factory, the model is SKL16-φ80×250-8, and the specific technical parameters are shown in Table 5.16. Experimental Process (1) Placement of experimental materials: First of all, the experimental steel and experimental slag are put into the alumina crucible in turn, and then the experimental crucible is placed in the center of the tubular resistance furnace with protective crucible. (2) Heating up: High purity argon gas is introduced for gas protection to prevent sample oxidation, turn on the power supply for heating for 3 h and raise the temperature to 1600 °C, and keep constant temperature for 30 min until the desulfurization experiment is over. (3) Sampling: After the desulfurization experiment is finished, the experimental crucible is directly taken out from the furnace for water quenching cooling, and the steel sample and slag sample at the end of the experiment can be directly obtained after cooling. Table 5.16 Main technical parameters of tubular resistance furnace Model

SKL16-φ80×250-8

Rated power

8 kW

Rated voltage

380 V

Phase number

2-phase

Maximum temperature

1650 °C

Rated temperature

1600 °C

Heating time of empty furnace

≤ 90 min

Heating element connection

Series connection

Silicon-molybdenum rod specification

φ6/250/280/30

Furnace size

φ80×250

138

5 Disposal of Solid Waste in Steelmaking Process

Table 5.17 Table of experimental schemes Experiment No.

CaO blending ratio, %

Steel sample quantity, g

CaO content, g Amount of reduction product, g

Amount of superimposed residue, g

1

40

93.05

3.72

12.20

9.31

2

45

93.41

4.20

11.23

9.34

3

50

92.84

4.64

10.14

9.28

4

55

93.32

5.13

9.18

9.33

5

60

93.36

5.60

8.16

9.34

6

Common refining slag

93.09

–

–

9.31

Table 5.18 Main components of experimental refining slag (wt%) Composition

CaO

Al2 O3

SiO2

TiO2

CaF2

Experimental optimal slag

55

28

8

4

5

Common refining slag

50

40

5

0

5

5.2.5.2

Experimental Scheme

According to the composition in Table 5.17, the refining slag prepared with different CaO ratios was desulfurized for 5 times, and the common refining slag (see Table 5.18 for the composition) was used for desulfurization experiment comparison, with a total of 6 experiments, as shown in Table 5.17.

5.2.5.3

Experimental Results and Analysis

Desulfurization Effect Carry out the experiment according to the above experimental arrangement, and send the steel samples obtained from the experiment to the testing center of the General Institute of Iron and Steel Research for chemical analysis. The experimental results are shown in Table 5.19 and Fig. 5.23. According to the experimental results, (1) The higher the ratio of CaO in the reduction product of red mud after sodium removal is, the higher the desulfurization rate is. However, when the ratio of Table 5.19 Experimental results of desulfurization of refining slag prepared Experiment No.

1

2

3

4

5

6 (common refining slag)

[S], ppm

160

120

81

42

31

46

Desulfurization rate ηS , %

38.46

53.85

68.85

83.85

88.08

82.31

5.2 Direct Reduction of Red Mud for Refining Desulfurization

139

Fig. 5.23 Desulfurization experiment results

CaO is higher than 55%, the increasing trend of desulfurization rate of the prepared refining slag slows down. High CaO content is naturally beneficial to desulfurization, but too high CaO content will lead to the decline of slag fluidity, which is not conducive to desulfurization reaction. (2) The ratio of CaO added to the reduction product of red mud after sodium removal is about 55%, the desulfurization rate of refining slag prepared by adding CaO to red mud after sodium removal is over 80%, and the sulfur in liquid steel is removed below 50 ppm. Comparing the desulfurization effect of common refining slag under the same conditions, it can be seen that the desulfurization ability of this slag has reached or even exceeded that of common refining slag. It can be concluded that the desulfurization capacity of refining slag prepared by adding about 55% CaO to the reduction product of red mud after sodium removal is equivalent to that of common refining slag, and it can be used as refining slag. Separation Effect of Iron Slag, a Reduction Product of Red Mud After Sodium Removal At high temperature, CaO combined with oxides in the reduction product of red mud after sodium removal to form refining slag, and iron in the reduction product of red mud after sodium removal melted into liquid steel. The changes of materials before and after the experiment are shown in Table 5.20. The mass of steel obtained from the experiment can be weighed, but the mass of slag obtained from the experiment cannot be weighed because all the slag obtained from the experiment cannot be collected. The data in Table 5.20 are estimated. It can be seen from Table 5.20 that during the desulfurization reaction, the yield of iron in the reduction product of red mud after sodium removal is about 99.9%, which shows that almost all the metallic iron enters the steel, thus realizing good

140

5 Disposal of Solid Waste in Steelmaking Process

Table 5.20 Changes of materials before and after desulfurization experiment Experiments Initial steel quantity, g

Amount CaO of content, reduction g products of red mud after sodium removal, g

Iron content in reduction products of red mud after sodium removal, g

End Feeding Yield Slag point steel of quantity, steel quantity iron, g quantity, % g

1

93.05

12.20

3.85

6.74

99.79

6.74

99.94 9.31

2

93.41

11.23

4.32

6.20

99.60

6.19

99.86 9.35

3

92.84

10.14

4.74

5.60

98.43

5.60

99.90 9.29

4

93.32

9.18

5.23

5.07

98.38

5.07

99.92 9.34

5

93.36

8.16

5.68

4.51

97.85

4.50

99.72 9.35

separation of metallic iron and other oxides in the reduction product of red mud after sodium removal.

References

1. Gordon, Yakov, and Sunil Kumar. 2013. Selection of ironmaking technology: Principles and risks. Transactions of the Indian Institute of Metals 66 (5): 501–513. 2. Cheng, Zhilong, Zhoutuo Tan, Zhigang Guo, Jian Yang, and Qiuwang Wang. 2020. Recent progress in sustainable and energy-efficient technologies for sinter production in the iron and steel industry. Renewable and Sustainable Energy Reviews 131: 110034. 3. Buergler, Thomas, and Irmela Kofler. 2017. Direct reduction technology as a flexible tool to reduce the CO2 intensity of iron and steelmaking. BHM Berg-und Hüttenmännische Monatshefte 162 (1): 14–19. 4. Gorai, Bipra, R.K. Jana, and Premchand. 2003. Characteristics and utilisation of copper slag—A review. Resources, Conservation and Recycling 39 (4): 299–313. 5. Zain, M.F.M., M.N. Islam, S.S. Radin, and S.G. Yap. 2004. Cement-based solidification for the safe disposal of blasted copper slag. Cement and Concrete Composites 26 (7): 845–851. 6. Shi, Caijun, Christian Meyer, and Ali Behnood. 2008. Utilization of copper slag in cement and concrete. Resources, Conservation and Recycling 52 (10): 1115–1120. 7. Liu, Zhanhua, Wenliang Chen, Yingui Ding, Zhicheng Cao, Daohong Wu, and Wen Yu. 2019. Study on recovery of iron and zinc from copper slag by direct reduction process in rotary hearth furnace. Metal Mine 48 (5): 183–187. 8. Michailova, Irena, and D. Mehandjiev. 2010. Characterization of fayalite from copper slags. Journal of the University of Chemical Technology and Metallurgy 45 (3): 317–326. 9. Zhou, Huihui, Guijian Liu, Liqun Zhang, and Chuncai Zhou. 2021. Mineralogical and morphological factors affecting the separation of copper and arsenic in flash copper smelting slag flotation beneficiation process. Journal of Hazardous Materials 401 (1): 123293. 10. Zuo, Zongliang, Qingbo Yu, Siyi Luo, Jingkui Zhang, and Enze Zhou. 2020. Effects of CaO on two-step reduction characteristics of copper slag using biochar as reducer: Thermodynamic and kinetics. Energy & Fuels 34 (1): 491–500. 11. Yang, Shuo, Amirhomayoun Saffarzadeh, Takayuki Shimaoka, and Takashi Kawano. 2014. Existence of Cl in municipal solid waste incineration bottom ash and dechlorination effect of thermal treatment. Journal of Hazardous Materials 267 (28): 214–220. 12. Wong, Guojing, Xiaohui Fan, Min Gan, Zhiyun Ji, Hengdi Ye, Zhian Zhou, and Zhaocai Wang. 2020. Resource utilization of municipal solid waste incineration fly ash in iron ore sintering process: A novel thermal treatment. Journal of Cleaner Production 263: 121400. 13. Huang, Binbin, Min Gan, Zhiyun Ji, Xiaohui Fan, Dan Zhang, Xuling Chen, Zengqing Sun, Xiaoxian Huang, and Yong Fan. 2022. Recent progress on the thermal treatment and resource utilization technologies of municipal waste incineration fly ash: A review. Process Safety and Environmental Protection 159: 547–565. 14. Xue, Yang, and Xiaoming Liu. 2021. Detoxification, solidification and recycling of municipal solid waste incineration fly ash: A review. Chemical Engineering Journal 420: 130349. © Metallurgical Industry Press 2023 K. Dong et al., Resource Utilization of Solid Waste by Thermometallurgy in Steel Processes, https://doi.org/10.1007/978-981-99-5655-5

141

142

References

15. De Casa, G., T. Mangialardi, A.E. Paolini, and L. Piga. 2007. Physical-mechanical and environmental properties of sintered municipal incinerator fly ash. Waste Management 27 (2): 238–247. 16. Wang, Kuensheng, Changjung Sun, and Chungchi Yeh. 2002. The thermotreatment of MSW incinerator fly ash for use as an aggregate: A study of the characteristics of size-fractioning. Resources, Conservation and Recycling 35 (3): 177–190. 17. Lam, Charles H.K., Alvin W.M. Ip, John Patrick Barford, and Gordon Mckay. 2010. Use of incineration MSW ash: A review. Sustainability 2 (7). 18. Liu, Yangsheng, Liting Zheng, Xiaodong Li, and Shaodong Xie. 2009. SEM/EDS and XRD characterization of raw and washed MSWI fly ash sintered at different temperatures. Journal of Hazardous Materials 162 (1): 161–173. 19. Hu, Hongyun, Huan Liu, Qiang Zhang, Pingan Zhang, Aijun Li, Hong Yao, and Ichiro Naruse. 2016. Sintering characteristics of CaO-rich municipal solid waste incineration fly ash through the addition of Si/Al-rich ash residues. Journal of Material Cycles and Waste Management 18 (2): 340–347. 20. Xue, Yang, Wenzhou Yu, Jie Mei, Weiyan Jiang, and Xuewei Lv. 2019. A clean process for alumina extraction and ferrosilicon alloy preparation from coal fly ash via vacuum thermal reduction. Journal of Cleaner Production 240: 118262. 21. Bai, Kaikai, Jiangwei Shen, Zhenglu Zhu, Haibin Zuo, Yuzhu Pan, Jingsong Wang, and Qingguo Xue. 2019. Effect of Al2 O3 on the formation of calcium ferrite in the solid state. Metals 9 (6). 22. Long, Hongming, Xuejian Wu, Tiejun Chun, Zhanxia Di, and Bin Yu. 2016. Assimilation behavior of calcium ferrite and calcium diferrite with sintered Al2 O3 and MgO. Metallurgical and Materials Transactions B 47 (5): 2830–2836. 23. Patrick, Tim R.C., and Mark I. Pownceby. 2002. Stability of silico-ferrite of calcium and aluminum (SFCA) in air-solid solution limits between 1240 °C and 1390 °C and phase relationships within the Fe2 O3 -CaO-Al2 O3 -SiO2 (FCAS) system. Metallurgical and Materials Transactions B 33 (1): 79–89. 24. Wang, Xiaozhe, Zhiwu Yan, Rongrong Wang, Donghui Liu, Jianliang Zhang, and Zhengjian Liu. 2018. Study of the enrichment characteristics of sinter by alkali metal vapors. Metallurgical Research & Technology 115 (3). 25. Meng, Fei, Qingca Liu, and Guangtai Zhu, et al. 2017. Experimental study on replacing part of sintering flux with steel slag magnetic separation tailings in sintering. Journal of Central South University: Natural Science Edition48 (1): 8. 26. Liu, Shouping, Guangyuan Wen, and Qionghua Zhu, et al. 2001. Experimental study on the use of converter steel slag in the production of sintered ore. Sichuan Metallurgical 23 (1): 3. 27. Li, Shuzhuang. 1991. Practice of using open hearth slag for sintering production. Steel 026 (002): 65–69. 28. Wu, He, and Yue Long. 2014. Experimental study on the basic properties of gas quenched steel slag sintering. Hebei Metallugy (02): 8–11. 29. Shen, Delin. 1991. Feasibility study on returning steel slag to sintering as a flux at baosteel. Baosteel Technology (1): 6. 30. Yu Mingxing, Sun Yin, and Jingdong Liu, et al. 2011. Production practice of steel slag sintering. Comprehensive Utilization of Resources in China 29 (9): 2. 31. Grilo, João P.F., Daniel A. Macedo, Rubens M. Nascimento, and Fernando M.B. Marques. 2020. Performance of GDC with alkali metal carbonates as sintering aids. Solid State Ionics 346 (C). 32. Min, Y., C. Qin, P. Shi, et al. 2017. Effect of municipal solid waste incineration fly ash addition on the iron ore sintering process, mineral phase and metallurgical properties of iron ore sinter. ISIJ International 57 (11): 1955–1961. 33. Peng, Z., R. Weber, Y. Ren, et al. 2020. Characterization of PCDD/Fs and heavy metal distribution from municipal solid waste incinerator fly ash sintering process. Waste Management 103: 260–267. 34. Zhan, X., L.A. Wang, L. Wang, et al. 2021. Co-sintering MSWI fly ash with electrolytic manganese residue and coal fly ash for lightweight ceramisite. Chemosphere 263: 127914.

References

143

35. Zeng, L., H.J. Sun, T.J. Peng, et al. 2019. The sintering kinetics and properties of sintered glass-ceramics from coal fly ash of different particle size. Results in Physics 15: 102774. 36. Chou, Sunyu, Shanglien Lo, Chinghong Hsieh, and Chinglung Chen. 2009. Sintering of MSWI fly ash by microwave energy. Journal of Hazardous Materials 163 (1): 357–362. 37. Fan, Xiaohui, Yannan Wang, Min Gan, Zhiyun Ji, Yang Zhou, and Xuling Chen. 2019. Thermodynamic analysis and reaction behaviors of alkali metal elements during iron ore sintering. Journal of Iron and Steel Research International 26 (6): 558–566. 38. Zeng, Li, Hongjuan Sun, Tongjiang Peng, and Tao Hui. 2021. Effect of glass content on sintering kinetics, microstructure and mechanical properties of glass-ceramics from coal fly ash and waste glass. Materials Chemistry and Physics 260: 124120. 39. Decker, B.F., and J.S. Kasper. 2010. The structure of calcium ferrite (correction). Acta Crystallographica 11 (4). 40. Min, Yi, Yutao Feng, Chongda Qin, Wen Li, Peiyang Shi, and Heng Wei. 2017. Effect of municipal solid waste incineration fly ash addition on property of iron ore sinter. Journal of Iron and Research Steel 29 (8): 610–615. 41. Bethanis, S., C.R. Cheeseman, and C.J. Sollars. 2002. Properties and microstructure of sintered incinerator bottom ash. Ceramics International 28 (8): 881–886. 42. Zhao, Jun, Hang Su, Haibin Zuo, Jingsong Wang, and Qingguo Xue. 2020. The mechanism of preparation calcium ferrite from desulfurization gypsum produced in sintering. Journal of Cleaner Production 267: 122002. 43. Sas, Z., W. Sha, M. Soutsos, et al. 2019. Radiological characterisation of alkali-activated construction materials containing red mud, fly ash and ground granulated blast-furnace slag. Science of the Total Environment 659: 1496–1504. 44. Wang, Yici, Jianliang Zhang, Fang Zhang, and Guoping Luo. 2011. Formation characteristics of calcium ferrite in low silicon sinter. International Journal of Iron and Steel Research 18 (10): 1–7. 45. Yin, Jiaqing, Xuewei Lv, Shenglin Xiang, Chenguang Bai, and Bin Yu. 2013. Influence of CaO source on the formation behavior of calcium ferrite in solid state. ISIJ International 53 (9): 1571–1579. 46. Matsumura M., M. Hoshi, and T. Kawaguchi. 2005. Improvement of sinter softening property and reducibility by controlling chemical compositions. Isij International—ISIJ INT 45: 594– 602. 47. Yasuda N., K. Nishioka, and S. Nomura. 2020. Formulation of shrinkage rate of sinter during softening process. ISIJ International 60: 1487–1494. 48. Kumar V, D. Patil, K. P. Mrunmaya, et al. 2020. Impact of sinter basicity and alumina on softening melting behavior in blast furnace. Transactions of the Indian Institute of Metals 73(2): 365–375.