Blasting Technology for Underground Hard Rock Mining 9819926440, 9789819926442

Citation preview

Vivek Kumar Himanshu A. K. Mishra M. P. Roy P. K. Singh

Blasting Technology for Underground Hard Rock Mining

Blasting Technology for Underground Hard Rock Mining

Vivek Kumar Himanshu · A. K. Mishra · M. P. Roy · P. K. Singh

Blasting Technology for Underground Hard Rock Mining

Vivek Kumar Himanshu CSIR-Central Institute of Mining and Fuel Research Dhanbad, Jharkhand, India M. P. Roy CSIR-Central Institute of Mining and Fuel Research Dhanbad, Jharkhand, India

A. K. Mishra Department of Mining Engineering Indian Institute of Technology (Indian School of Mines) Dhanbad, Jharkhand, India CSIR-Central Institute of Mining and Fuel Research Dhanbad, Jharkhand, India P. K. Singh CSIR-Central Institute of Mining and Fuel Research Dhanbad, Jharkhand, India

ISBN 978-981-99-2644-2 ISBN 978-981-99-2645-9 (eBook) https://doi.org/10.1007/978-981-99-2645-9 © The Editor(s) (if applicable) and The Author(s), under exclusive license to Springer Nature Singapore Pte Ltd. 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 publisher, 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 publisher 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 publisher remains 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

The mineral industries have shown a significant shift from openpit to underground method of mining due to extinction of shallow depth ore deposits. The openpit mining is preferred for the faster rate of production. So, maintaining the production pace from underground mining to meet the rising raw material demands is the main challenge for the mining practitioners. The large-scale blasting is the only alternative to meet the target production with faster pace. Accordingly, the scientists, practitioners and planners have devised various methods of underground blasting. These methods have been devised to deal with the practical challenges of the mining. The methods are based on the scientific predictions using advanced numerical simulation and statistical analysis. The devised methods were further implemented at the site, and the desired outcomes were achieved. This book consists of the discussions on these blasting methodology, scientific analysis for underground blast, instrumentations during blasting operation, etc. The book encompasses the detailed discussion on the recent advancements in the underground blasting for hard rock mining under Indian geomining conditions. This book consists of nine chapters. Chapter 1 introduces the readers regarding the need for underground blasting. The importance of drilling and blasting has been briefed in this chapter with the help of statistical data. The progressive advancements in underground blasting have also been described in this chapter. Chapter 2 encompasses a brief discussion about underground metalliferous mining methods. The methods used in Indian mineral industries have been described in detail in this chapter. The discussion regarding the role of rock–explosive interaction on blastinduced damages has been made in Chap. 3. This chapter gives a brief outlook to the readers regarding the mechanism of rock blasting and parameters affecting the blasting outcomes. The underground blasting for hard rock mining progresses in three major steps as drivage development, stope opening and production blasting. The discussions regarding the science and applicability of these three steps have been made in Chaps. 4, 5 and 6. The blasting operation has associated hazards in terms of ground vibration, air overpressure, flyrocks, etc. The ground vibration is the major hazard which affect structural safety. The discussion regarding the prediction and control of induced ground vibration from underground blasting has been done v

vi

Preface

in Chap. 7. Some of the special blasting techniques used in the Indian mines to overcome the safety and productivity-related challenges have been discussed in Chap. 8. Chapter 9 provides a summary of this book. The future challenges with underground blasting and way forward to deal with the future challenges have also been discussed in this chapter. We hope that this book would be a knowledgebase to the mining professionals. The methodology used in the book may be replicated in new mines with similar site conditions. The analytical part of the book will give the research professionals an outlook to deal similar problems with scientific approach. The academicians and students of the related discipline will also be benefitted with the diversified content of this book. Dhanbad, India

A. K. Mishra

Acknowledgements

It would not have been possible to write this book without the help and support of the kind people around us, to only some of whom it is possible to give a particular mention here. Authors would like to express their sincere thanks and gratitude to Dr. C. Sawmliana, Chief Scientist and Head of Section of Rock Excavation Engineering Research Group of CSIR-CIMFR, Dhanbad, for providing necessary technical inputs as and when required during experimentation at different study sites. We are also thankful to the other members of this research group for their help and cooperation during the experimentation at the study sites. We express our special thanks to Mr. Ashish Kumar Vishwakarma, Technical Officer, CSIR-CIMFR, Dhanbad, for his help in editing the figures for this book. Our heartfelt thanks goes to Dr. P. Pal Roy, Former Outstanding Scientist, Dr. M. M. Singh, Former Chief Scientist and Dr. B. M. P. Pingua, Former Chief Scientist of CSIR-CIMFR, Dhanbad, Prof. U. K. Dey, Former Director of BIT Sindri and others for their support and encouragements. Authors are also thankful to the mine management of M/s Hindustan Zinc Limited (Vedanta) and M/s Hindustan Copper Limited for their support and cooperation during field studies. We would also like to thank Mr. Praveen Sharma, Chief Operating Officer of M/s HZL, Mr. Arun Kumar Shukla, CMD of M/s Hindustan Copper Limited and Mr. Prabhat Kumar, Director General, DGMS, Dhanbad. Last but not the least, we express special thanks to our family members. We received their continuous support, constant care, prayers and countless love which always prompted us to work dynamically. We thank God for showering divine blessings, though abstract in nature but potent enough to be felt of its efficacy, without which nothing can perhaps be accomplished in this world. Vivek Kumar Himanshu

vii

Contents

1 Introduction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1.1 Preamble . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1.2 Underground Blasting for Metalliferous Mining . . . . . . . . . . . . . . . . . 1.3 Advancements in Underground Production Blasting . . . . . . . . . . . . . 1.4 Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . References . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

1 1 3 3 6 7

2 Overview of Underground Metalliferous Mining . . . . . . . . . . . . . . . . . . 2.1 Introduction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.2 Stages in Underground Metalliferous Mining . . . . . . . . . . . . . . . . . . . 2.3 Stoping Methods Practised in India . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.3.1 Sublevel Stoping . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.3.2 Longhole Stoping . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.3.3 Room and Pillar Stoping . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.3.4 Cut and Fill Stoping . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.4 Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . References . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

9 9 10 12 13 15 20 22 23 23

3 Rock–Explosive Interaction During Underground Blasting . . . . . . . . . 3.1 Introduction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.2 Rock Breakage Mechanism . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.3 Rock Mass Damages Under Dynamic Loading . . . . . . . . . . . . . . . . . . 3.3.1 Constitutive Models Used for Numerical Simulation of Rock Blasting . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.4 Rock Parameters and Their Role on Damage Characteristics . . . . . . 3.4.1 Nearfield Vibration and Rock Breakage Characteristic . . . . . 3.5 Influence of Explosive Parameters on Rock Mass Damage . . . . . . . . 3.5.1 Commercial Explosive and Its Principle of Explosion . . . . . 3.5.2 Explosive Parameters . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.5.3 Equation of State for Explosive . . . . . . . . . . . . . . . . . . . . . . . . 3.6 Impacts of Blast Design Parameters on Rock Mass Damage . . . . . . 3.6.1 Blasthole Diameter . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

25 25 26 27 30 30 33 34 34 35 39 40 40 ix

x

Contents

3.6.2 Burden . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.6.3 Spacing . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.6.4 Stemming . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.6.5 Subgrade Drilling . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.6.6 Charging Parameters . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.6.7 Delay Timings . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3.7 Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . References . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

41 42 43 43 43 43 44 44

4 Drivage Excavation Using Drilling and Blasting . . . . . . . . . . . . . . . . . . . 4.1 Introduction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 4.2 Elements of Burn-Cut Blast Design . . . . . . . . . . . . . . . . . . . . . . . . . . . 4.3 Measurement of Outputs from Underground Drivage Blasting . . . . . 4.4 Optimization of Drivage Blasting Pattern . . . . . . . . . . . . . . . . . . . . . . 4.4.1 Numerical Simulation and Prediction of Damages from Drivage Blasting . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 4.4.2 Designing of Delay Pattern for Drivage Blasting Using Nearfield Ground Vibration Monitoring Results . . . . . 4.5 Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . References . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

49 49 51 52 54

5 Box-Cut Excavation and Stope Opening . . . . . . . . . . . . . . . . . . . . . . . . . . 5.1 Introduction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 5.2 Principle of Underground Slot Raise Blasting . . . . . . . . . . . . . . . . . . . 5.3 Drilling Pattern for Slot Raise Blasting . . . . . . . . . . . . . . . . . . . . . . . . 5.4 Charging and Initiation of Blastholes in Slot Raise Blasting . . . . . . . 5.5 Delay Sequence of Blastholes for Slot Blasting Pattern . . . . . . . . . . . 5.6 Special Slot Raise Blasting Pattern . . . . . . . . . . . . . . . . . . . . . . . . . . . . 5.7 Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . References . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

65 65 67 67 68 69 70 71 73

6 Underground Ring Blasting . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 6.1 Introduction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 6.2 Drilling Pattern for Underground Ring Blasting . . . . . . . . . . . . . . . . . 6.3 Charging of Ring Blastholes . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 6.3.1 Suitability of Explosive for Different Rock Strata . . . . . . . . . 6.3.2 Charge Factor and Dimensional Parameters for Ring Blasting . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 6.4 Connection and Firing Sequence of Charged Ring Blastholes . . . . . 6.4.1 Optimization of Delay Sequence and Delay Timing for Production Blasting . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 6.4.2 Delay Timing for Multi-Ring Blasting . . . . . . . . . . . . . . . . . . 6.5 Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . References . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

75 75 76 78 80

56 58 61 62

82 84 85 88 89 90

Contents

xi

7 Blast-Induced Hazards . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 93 7.1 Introduction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 93 7.2 Standards on Ground Vibration Limits . . . . . . . . . . . . . . . . . . . . . . . . . 94 7.3 Sensitive Structures Prone to Ground Vibration Hazards . . . . . . . . . 96 7.4 Optimisation of Ring Blasting Parameters for Structural Safety . . . 98 7.4.1 Parameters Affecting Blast-Induced Ground Vibration . . . . . 98 7.4.2 Case Studies on Designing Controlled Blasting Pattern for Underground Ring . . . . . . . . . . . . . . . . . . . . . . . . . . 99 7.5 Safety Precautions While Blasting . . . . . . . . . . . . . . . . . . . . . . . . . . . . 103 7.6 Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 103 References . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 104 8 Innovative Blasting Practices for Underground Hard Rock Mining . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 8.1 Introduction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 8.2 Technique for Simultaneous Excavation of Slot Raise and Rings in a Single Blasting Round . . . . . . . . . . . . . . . . . . . . . . . . . 8.3 Drilling and Blasting Methodology for Extraction of Narrow Vein Ore Deposits . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 8.4 Multi-Ring Blasting Pattern for Rock Excavation in Ore-Waste Combination . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 8.5 Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . References . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 9 Challenges and Way Forward . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 9.1 Overview . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 9.2 Future Challenges of Underground Blasting and Way Forward . . . . 9.3 Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . References . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

107 107 108 111 114 116 117 119 119 120 122 122

About the Authors

Dr. Vivek Kumar Himanshu is Senior Scientist at CSIR-Central Institute of Mining and Fuel Research, Dhanbad. He did his Ph.D. in Mining Engineering from Indian Institute of Technology (Indian School of Mines), Dhanbad. He graduated in Mining Engineering from B.I.T Sindri and M.Tech. in Mine Safety Engineering from Academy of Scientific and Innovative Research, CSIR-CIMFR, Dhanbad. He joined Hindustan Copper Limited as Graduate Engineer Trainee (Mining) and served for one year in various capacity of underground mine supervision, mine planning, EIA/EMP study, etc. He served as Trainee Scientist at CSIR-CIMFR, Dhanbad, during his period of post-graduation and worked for various research projects of metalliferous mining and its simulation design. He joined National Institute of Technology Rourkela after his M.Tech. and worked there as Assistant Professor for 8 months. He gathered teaching, research, and academic experience during this period. Currently, he is working at Rock Excavation Engineering Research Group of CSIR-CIMFR, Dhanbad. He has received Young Engineers Award 2022 from The Institution of Engineers (India) for his contributions in mining engineering. His area of research includes rock excavation by blasting, controlled blasting, blast vibration prediction, blast simulation, etc. He has published more than 30 papers in different journals and proceedings of seminar. Prof. A. K. Mishra is Mining Engineer with M.Tech. and Ph.D. in Mining Engineering from Indian Institute of Technology (Indian School of Mines), Dhanbad, having teaching, industrial, and administrative experience of 31 years. Presently, He is Director of CSIR-Central Institute of Mining and Fuel Research, Dhanbad. He is also Professor en-lien in the Department of Mining Engineering, IIT(ISM), Dhanbad. After graduation, he joined Coal India Limited as Mining Engineer and served for three years. He then joined the IIT(ISM), Dhanbad, as Faculty Member in 1992. He has also served as Technical Services Manager in Orica Mining Services, Australia, and headed the technical services function of Indian Business during 2005 to 2009. He worked as Chair Professor (Uranium) sponsored by Department of Atomic Energy, GoI. His areas of specialization include opencast and underground mining technology, drilling and blasting technology, rock excavation engineering, xiii

xiv

About the Authors

tunnelling, geomechanics, ground control, and strata management. Prof Mishra has to his credit, 193 research publications in peer-reviewed reputed journals, international/national conferences, and symposia. He is actively involved in solving real-life problems of mining and allied industry by undertaking R&D and industry sponsored projects in mining and allied areas and has completed more than 139 such projects. He has guided 10 Ph.D. dissertations, while another 15 are ongoing. He has completed 29 Management and Executive Development Programs for various mining, IT, and other allied industries. He has been awarded with the prestigious National Geoscience Award from the Ministry of Mines, GoI by Honorable President of India in 2016. He is also Recipient of the Abheraj Baldota Memorial Gold Medal Award—2012 (Young Mining Engineer of the year 2012) from Mining Engineer’s Association of India. Dr. M. P. Roy is Senior Principal Scientist at CSIR-Central Institute of Mining and Fuel Research, Dhanbad. He has made notable contributions in blasting science and technology to solve practical problems in the mining industry encompassing the fields of blast vibration and associated correlation with structural damage, wall control, rock fragmentation, blast design in mines and tunnels, cast blast designs in coal mining operations, and controlled blasting in civil construction projects. He has worked as Project Leader in two Mega S&T Project titled “Characterization of rock and explosive parameters for optimal explosive energy utilization in opencast blasting” and “Blast Design and Fragmentation Control-Key to Productivity” and significantly contributed in another five (5) R&D Projects of Ministry of Coal, Government of India, in different area of rock blasting. He has published more than 90 scientific papers in international and national journals and conferences. He has received CSIR Technology Awards 2019. Dr. P. K. Singh is Former Director of CSIR-Central Institute of Mining and Fuel Research, Dhanbad. He is a Doctorate in Engineering from Institute of Mining Engineering, Technical University, Clausthal, Germany. His areas of research interest include rock mechanics, mining methods, rock excavation engineering, coal characterization, and clean coal technologies. He has published more than 200 papers in scientific journals and proceedings of seminars. He has also authored and edited 13 books. He has been Project Leader and Advisor in about 20 S&T projects sponsored by Government of India. Dr. Singh received different prestigious awards including National Mineral Award, Raman Research Fellowship, German Academic (DAAD) Fellowship, CSIR Technology award, etc. He is Fellow of the Indian Academy of Sciences (FASc), Bengaluru, and National Academy of Sciences (FNASc), Allahabad.

Chapter 1

Introduction

Abstract Industrial revolution has increased the mineral/metal consumption during previous decades. Accordingly, there was fast depletion of shallow depth mineral deposits. The excavation of deep seated deposits is possible only using underground mining, as openpit mining for such deposit will not be economical. Maintenance of production pace from underground mining as par with openpit mining is a challenging task. However, with the technological advancements, it has been possible to do largescale excavation from underground as well. Such technological innovations have addressed the associated operational and safety related issues of underground mining. In this chapter, the introductory details of the need of the mineral, advancements in underground mining and advancements in technologies to address the challenges of underground mining have been discussed.

1.1 Preamble The mineral and energy consumption is one of the major precursors of the socioeconomic growth of a country. Esen and Bayrak (2017) found in a study that energy consumption has significant contribution in the economic growth of a country. Mudakkar et al. (2013) conducted the study to correlate Gross Domestic Products (GDP), Foreign Direct Investment (FDI) and other financial development indicators with energy consumption for South Asian Association for Regional Cooperation (SAARC) countries. The study concluded energy consumption as a major contributor in GDP and promoting FDI. The study on the global status of mineral consumption further reveals that per capita mineral consumption for developed countries are higher than the developing nations (US Department of Interiors, accessed on 04.10.2021). India has targeted to be a USD 5 trillion economy by 2024–25 (The Economic Times, accessed on 04.10.2021). The mineral and energy consumptions will eventually rise to fulfil this target. Such consumption demand can be addressed by enhancing the mineral and energy production or accelerating the imports. Mining plays the significant role in bridging the gap between demand and supply of the minerals and energy. In Indian context, mining and quarrying sector accounts for 2.5% of country’s Gross Domestic Product (GDP). The contributions of mining © The Author(s), under exclusive license to Springer Nature Singapore Pte Ltd. 2023 V. K. Himanshu et al., Blasting Technology for Underground Hard Rock Mining, https://doi.org/10.1007/978-981-99-2645-9_1

1

2

1 Introduction

sector in GDP of the developed countries are more than 10% (National Mining Association, accessed on 04.10.2021). Drilling and blasting is a dominant method for excavation of coal and minerals throughout the globe to date. The rock mass is fractured using the explosive energy in this method. The dominancy of this method can be gauged from the data of global explosive consumption. The total cost of global consumption of industrial explosives was $1.7 billion in 2019, which is projected to be $10.9 billion by 2027 (Alliedmarketsearch, accessed on 15.07.2021). The coal mining sector is the largest consumer of the industrial explosives. It consumes around 40% of the total industrial explosives. The second largest consumer is metalliferous mining sector with 33% consumption of the explosives. The consumption of explosives in stone quarries and limestone mining are around 16% (Ihsmarkit, accessed on 15.07.2021). In India, drilling and blasting contributes to more than 95% excavation of coal and 100% excavation of hard rock metalliferous deposits. As per DGMS (2015a, b), altogether 571 coal mines, 19 lignite mines and 1183 metalliferous mines use drilling and blasting based techniques for mining in India. The coal cutting machines are used in nine Indian coal mines. The total consumption of industrial explosives during 2015 for coal and non-coal mining in India were 591,099 tonnes and 105,428 tonnes, respectively (DGMS 2015a, b). Moreover, the excavation of hard rocks in various civil construction works, viz. tunnel construction, roadway construction, foundations for industrial establishments, etc., is also carried out by drilling and blasting. Despite the significant advancements in blasting science and technology, effective utilization of explosive energy is still a challenge, as only a small fraction of the energy (20–30%) is helpful for rock breakage, and the rest is considered waste energy (Monjezi et al. 2011). This is well known that waste energy contributes to undesirable phenomena such as ground vibration, flyrocks, air blast, noise, over breaks, undesirable throw, etc. (Parida and Mishra 2015; Rebello et al. 2016). Various controllable and uncontrollable parameters impact the blasting performance and consequently the utilization of explosive energy. The controllable parameters include the blast design parameters like—the amount of explosive charge, detonation velocity of explosive, density of explosive, delay timing between different blastholes, the accuracy of delay detonators, etc. The uncontrollable parameters consist of rock mass conditions at the blasting site, isotropy and homogeneity of the rock mass, blast vibration propagation media, etc. The prior assessment of uncontrollable parameters and thereby designing a blast can lead to achieve optimum utilization of explosive energy. This may be accomplished by the assessment of the rock–explosive interactions under a defined geomining condition. Saharan et al. (2006) suggested that the engineering system of rock–explosive interaction consists of rock properties, explosive properties, and boundary conditions. The suggested system is shown in Fig. 1.1. The rock properties in this system included geological and geotechnical parameters. The explosive properties included energy parameters of explosive and stress wave propagation characteristics. The boundary condition mainly included the status of the free face condition. The boundary conditions will vary depending on the mining methods. The rock and explosive properties also show a great extent of variability. The consideration of suitable boundary condition becomes more vital, when the excavation has to

1.3 Advancements in Underground Production Blasting

3

be carried out by underground. The challenges with the underground blasting are also different from the bench blasting or foundation blasting for civil construction works. To overcome these challenges, the significant advancements in the prediction method, engineering and operational technique of underground blasting have occurred during last decades. This book presents the discussion regarding these advancements.

1.2 Underground Blasting for Metalliferous Mining Underground mining is preferred for the exploitation of minerals/coal situated at greater depth. The method is advantageous in different other aspects also such as enhancement of land use, reduction in environmental pollution, etc. The large-scale underground mining of coal is mainly done by mechanized excavations. Continuous miners and shearers are the foremost excavators for the underground mining of coal. However, the drilling and blasting-based method is used for the underground mining of hard rock deposits. The hard rock excavations include the exploitation of metalliferous ore deposits. Underground mining is practised in India for the exploitation of minerals such as Copper, Lead–Zinc, Chromite, Manganese, Uranium, etc. These deposits are excavated by different stoping methods such as cut and fill, room and pillar, post-pillar, sublevel stoping and its variants. The ore deposits of steeper inclination are exploited using cut and fill, room and pillar or post-pillar stoping methods. The rate of excavation in these stoping methods is relatively slow. The method includes small diameter drilling and blasting. The nearly vertical dipping ore deposits are exploited using the sublevel stoping method and its variants. The variants of this method include—longitudinal longhole stoping, transverse longhole stoping, vertical retreat mining, etc. The advanced mining using the longhole stoping method uses large diameter drilling with higher consumption of explosives for blasting. The method has resulted in achieving a faster production pace in underground mining. This stoping method is used for the mining of Lead–Zinc ore and Copper deposits in India. The steps in underground excavation using longhole stoping include excavation of drivages, slot raises and rings. The major challenges with underground blasting include—pull enhancement from drivage (face) blasting, overbreak control in face blasting, damage optimization in slot raise excavation and fragmentation control in ring blasting. The blast design parameters are optimized to overcome these challenges.

1.3 Advancements in Underground Production Blasting The significant advancements have occurred in underground production blasting during last decades. The advancements were a need to increase the production pace in order to meet the rising demand of the raw materials. The advancements have also occurred to deal with the operational and safety challenges of the mine. The

Fig. 1.1 Engineering system of rock–explosive interaction (Saharan et al. 2006)

4 1 Introduction

1.3 Advancements in Underground Production Blasting

5

state-of-the-art prediction methodologies and instrumentations were used to reach the operational advancement. The underground ring blasting under Indian condition was started with the excavation of only four holes of a ring. The number of holes was restricted as the surface residential structures were in the close proximity (within 300 m) from the blasting face. However, the rising demand of production could not be made by blasting smaller number of holes in a round. With the advent of more accurate delay detonators, it was found that the larger number of holes separated with long delay timing can be blasted, even by restricting the ground vibration near structures within safe limit. Further, it was felt that the development of deeper level interval is possible, when the orebody is nearly vertical. In such case, the level interval of more than 50 m was developed for some of the mines, in order to reduce the development cost. Now, the complete extraction of the deeper ring blastholes was a challenging task. It was expected that maximum explosive charge per delay (MCPD) would increase while taking complete length of a ring blasthole in a single round, and thereby ground vibration would increase near the structures. The statistical analysis of the recorded vibration data was carried out to compute the optimum MCPD. It was thought to go for decking to reduce MCPD. The optimum deck length was decided by studying the breakage characteristic of rock. The deck length was chosen, so that it should not lead to the formation of boulders. The suitable delay timing, plugging system, etc. were also decided to maximize the utility of explosive energy. Based on the outcomes of the analysis, the multi-decking ring blasting methodology was devised to control the ground vibration along with taking complete level of a stope in one go. Further, the multi-ring blasting was also used to increase the productivity. It was found that the firing of larger quantity of explosive in a blasting round lead to superposition of blast vibration wave from different delay, specifically at far-field distances. The intensity of ground vibration in such case increased near the far-field underground structures, viz. shaft, decline, underground workshop, etc. This challenge was undertaken by selecting the optimal delay time between two rings. The selection was made such that the initiation of second ring had started only after the complete dampening of vibration from first ring. Working in shear zone was another challenge at one of the mines. The manpower deployment after different round of blasting was a risky task for the faces with shear zone. Accordingly, a special method was devised so that there would be a need of manpower deployment in only three rounds of blasting. The first round of blasting in this method had to be carried out from lower level to upper level using uphole charging pattern. The human resources would be safe under a specially designed canopy in uphole charging pattern. In second round, the complete extraction of slot raise, slot raise expansion and ring holes of hangingwall direction was taking place. Almost 60–70% of the stope was excavated in the second round. In final round, only the rings of footwall side were blasted. Since, the hangingwall zone was already blasted in second round, the risk to the working manpower in third round would be minimal. This method has also resulted into further enhancement in productivity by conducting larger blast in one go. This would have happened due the advent of special type of high viscosity explosive for uphole charging. Such explosive is compatible

6

1 Introduction

Table 1.1 Technological advancements in underground blasting for metalliferous mining Issues addressed

Methodology

Technological output

Production enhancement

Numerical simulation, experimental trials on simulation results

Ring blasting with larger number of holes

Safety of surface structures

Experimental trials with various MCPD, statistical analysis, experimental trial with reduced MCPD

Multi-deck ring blasting

Safety of underground structures

Experimental trials, waveform Delay design for multi-ring analysis and subsequently blasting dampening of the effect of total explosive charge on blast vibration

Working in shear zone

Prospects to extract complete ring in a single go was examined. Experimental trials and analysis

Making space for accumulation of blasted muckpile of complete stope

Prospects for uphole charging Uphole charging and blasting and blasting was examined and pattern analysed

Simultaneous extraction of slot, slot expansion and ring in one go

to stick to the blastholes and thereby works as a resistance to fall under gravity. The advancements in the blasting methodologies for underground metalliferous mining are summarized in Table 1.1. The detailed discussion on these advancements along with their principle and practices has been discussed in different chapters of this book.

1.4 Summary The enhancement of production from underground is the need of the hour. Accordingly, the larger underground blast is the prime need of the mining industry. The operation of large-scale underground blasting comes with different safety and site specific operational challenges. The scientific design, advanced data analysis and utilization of advanced operational techniques are the solution of these challenges. Indian mining industry has seen progressive technological advancements in rock blasting from last decade. These advancements have enabled the mine management to conduct underground blast with total explosive consumption of 22 tonnes in a blasting round. The advancements in prediction methodologies and designing techniques have also enabled the management to achieve longer pull, reduced overbreak, improved fragmentation and optimized muckpile distribution.

References

7

References Alliedmarketsearch. https://www.alliedmarketresearch.com/industrial-explosives-market. Accessed on 15 July 2021 DGMS (2015a) Statistics of mines in India Volume-I (Coal). http://dgms.gov.in/writereaddata/Upl oadFile/Coal_2015a.pdf DGMS (2015b) Statistics of mines in India Volume-II (non-Coal). http://dgms.gov.in/writereaddata/ UploadFile/VOLUME_II_NON_COAL_2015b.pdf Esen Ö, Bayrak M (2017) Does more energy consumption support economic growth in net energyimporting countries? J Econ Fin Admin Sci 22(42):75–98. https://doi.org/10.1108/JEFAS-012017-0015 Ihsmarkit. https://ihsmarkit.com/products/explosives-and-blasting-chemical-economics-han dbook.html. Accessed on 15 July 2021 Monjezi M, Ghafurikalajahi M, Bahrami A (2011) Prediction of blast-induced ground vibration using artificial neural networks. Tunn Undergr Space Technol 26(1):46–50. https://doi.org/10. 1016/j.tust.2010.05.002 Mudakkar SR, Zaman K, Shakir H, Arif M, Naseem I, Naz L (2013) Determinants of energy consumption function in SAARC countries: balancing the odds. Renew Sustain Energy Rev 28:566–574 National Mining Association. https://nma.org/wp-content/uploads/2016/09/Economic-Contribut ions-of-Mining-in-2015-Update-final.pdf. Accessed on 4 Oct 2021 Parida A, Mishra MK (2015) Blast vibration analysis by different predictor approaches-a comparison. Procedia Earth Planet Sci 11:337–345. https://doi.org/10.1016/j.proeps.2015.06.070 Rebello NE, Shivashankar R, Sastry VR (2016) Response of strata and buildings to blast induced vibrations in the presence and absence of a tunnel. Geotech Geol Eng 34(4):1013–1028. https:// doi.org/10.1007/s10706-016-0021-y Saharan MR, Mitri HS, Jethwa JL (2006) Rock fracturing by explosive energy: review of state-ofthe-art. Fragblast 10(1–2):61–81. https://doi.org/10.1080/13855140600858792 The Economic Times. https://economictimes.indiatimes.com/news/economy/policy/govt-sticksto-usd-5-trillion-economy-target-emphasis-on-infra-aimed-at-achieving-goal-dea-secretary/art icleshow/80686511.cms?from=mdr. Accessed on 4 Oct 2021 US Department of Interiors. https://www.doi.gov/ocl/global-mineral-consumption. Accessed on 4 Oct 2021

Chapter 2

Overview of Underground Metalliferous Mining

Abstract Various stoping methods are practised across the globe for the final extraction of mineral deposits. The selection of a feasible mining method for a mine is dependent on the technical suitability and economical profitability. The technical suitability of a method for a site is assessed by evaluation of the thickness, strength and inclination of the ore deposit as well as the strength of the wall rocks. The economic profitability assessment includes the evaluation of capital investment, pace of production and additional cost in account of safety and environmental concern. In Indian condition, room and pillar stoping, cut and fill stoping, sublevel stoping and their variants are used for exploitation of Copper, Lead–Zinc, Mica, Chromite and other minerals. The brief discussions on these methods have been done in this chapter.

2.1 Introduction The mineral industries have seen significant shift towards the underground mining due to extinction of shallow depth ore deposits (Carvalho 2017). The exploitation of mineral deposits by underground is advantageous in many other aspects as well. It is comparatively safer for the nearby resident in respect of the hazards associated with flyrock ejections during blasting (Raina et al. 2014). It is also environmentally viable due to absence of dust emission and propagation under the circumstances of openpit blasting. It also provides the opportunity for simultaneous land utilisation while mining (Chugh 2018; Mborah et al. 2015; Whittle et al. 2018). In Indian mining scenario, underground mining is practised for exploitation of Copper, Lead–Zinc– Silver, Gold, Uranium, Chromite and Mica deposits (TERI 2019; Tewari et al. 2020; Verma and Chaudhari 2017). During the recent years, the complete transition from openpit to underground mining has been seen for Copper and Lead–Zinc deposits in India (Singh et al. 2015). The major minerals which are excavated by openpit mining includes Iron ore and manganese. The openpit mining for these deposits are feasible till date due to their higher cut off grades. However, with the faster excavation to fulfil the rising raw material needs, there is prospect for transition of mining methods for the excavation of these mineral deposits as well. There is existence of underground iron © The Author(s), under exclusive license to Springer Nature Singapore Pte Ltd. 2023 V. K. Himanshu et al., Blasting Technology for Underground Hard Rock Mining, https://doi.org/10.1007/978-981-99-2645-9_2

9

10

2 Overview of Underground Metalliferous Mining

ore mines viz. Kiruna underground mine in global scenario (Dineva and Boskovic 2017; Newman et al. 2006). To cope with the faster excavation targets from underground is one of the major challenges for mineral industries. Selection of suitable mining method plays an important role over here. The conventional way for selecting suitable mining method is based on the physicomechanical properties of host rock and orebody (Deshmukh 2009; Hartman 2002). The need today however is to optimize the suitability of mining method based on the physicomechanical properties and production demand. The details regarding the selection of suitable mining method have been discussed in this chapter. The discussion has also been made regarding the basic overview of underground metalliferous mines, associated terminologies and different mining methods practised in India.

2.2 Stages in Underground Metalliferous Mining The underground mining for mineral exploitation comprises of mainly three stages viz. making access to the orebody, initial/box-cut excavation and final excavation of ore. The primary access to the underground mine is in the form of shaft, incline, decline or adit. These accesses are made from the surface. The mine is further connected to different secondary access viz. drive/drivage, cross-cuts through primary access. The secondary access works as approach to the orebody. The location for primary access to the mine is optimised so that to minimise the developments of secondary access during the advancement of mining operations. A mine should have minimum two primary access. Sometimes, tertiary access is also made as chute, ore pass or manway. The overview of different access to an underground mine is shown in Fig. 2.1. Once the access to the mine is made, the excavation of the orebody is carried out. The ore extraction takes place using different stoping methods. The dominant stoping methods practised under Indian geomining conditions have been discussed in different subsections of this chapter. The major terminologies used in underground metalliferous mining are defined below: Dip: Angle of inclination of a deposit measured from the horizontal. Strike: Horizontal bearing of a tabular deposit at its surface intersection. Country Rock/Host Rock: Rocks lying in adjacent to the mineral deposit. If such rock is lying under the deposit, it is termed as footwall. If the rock is lying above the ore deposit, it is termed as hangingwall. Incline: Inclined opening driven to connect surface from underground. Shaft: Vertical opening connecting the surface with underground openings. Decline: Primary access to the mine driven in nearly helical fashion. Adit: Primary access to the mine, which is horizontal opening to the underground deposit. Level: System of horizontal openings connected to a shaft.

2.2 Stages in Underground Metalliferous Mining

11

Fig. 2.1 Outlook of terminologies for an underground mine (Himanshu et al. 2021a)

Drift/Drive/Drivage: Access made parallel to the orebody. The access exposing footwall side of the ore body is termed as footwall drive/drivage. The access exposing hangingwall side of the orebody is termed as hangingwall drive. Hangingwall and footwall drivages together work to expose the ore body. They are connected with crosscut. Crosscut: Access connecting drivages from shaft/decline or access connecting two drivages. Haulageway: Horizontal opening used primarily for materials handling. Generally, footwall drivages are used as haulageway. Manway: Compartment of a raise or a vertical or near vertical opening intended for personnel travel between two levels. Ore pass: Vertical or near vertical opening through which rock material flows by gravity. Chute: Opening from a draw point, utilizing gravity flow to direct excavated rock from ore pass to load a conveyance. Draw point: Loading point beneath a stope, utilizing gravity to move rock material downward and into a conveyance. Ramp: Openings driven to connect levels usually in a downward direction and used for haulage. Raise: Vertical or near vertical opening driven upward from one level to another. Winze: Vertical or near vertical opening driven downward from one level to another. Stoping: Final excavation of mineral. Stope: Area comprising the exposed minerals for final exploitation.

12

2 Overview of Underground Metalliferous Mining

Stope Back: Roof, top or overlying surface of excavated stope. Cap Rock: Overlying material or burden above an underground stope. Pillar: Unexcavated portion of an ore deposit which is left to ensure stability of host rock and overlying burden during excavation. Crown Pillar: Pillar which is left above the excavated stope to ensure stability of drivages. Rib Pillar: The pillar left between two stopes. This pillar is used to ensure the stability of wall rock after excavation. Sill Pillar: Pillar which is left below the excavated stope to ensure stability of drivages. Slot: Initial/box-cut opening in the ore body to make free face for the subsequent round of stope excavation.

2.3 Stoping Methods Practised in India Variety of mineral and coal deposits are present across different parts of India. These deposits are excavated dominantly by drilling-blasting method (DBM). However, mechanical excavators using shearer, continuous miner, surface miner and high-wall mining equipment are used at few places for the excavation of coal deposits. Most of the precious metals like Copper, Lead–Zinc–Silver, Gold, etc. are excavated dominantly by underground mining method. However, openpit mining is still dominant for the excavation of Iron ore deposits. Chromite and Mica deposits are excavated by both openpit as well as underground excavation methodology. An overview of excavation methodology for different coal and metalliferous deposits is shown in Fig. 2.2. The charge factor suggested by many researchers is mostly for the excavation using open pit mining method. The charge factor requirement for underground excavation will be different specially in the case of hard rock excavations, as the mining method in such cases is devised to extract orebodies only. Different excavation methods are used for exploitation of mineral deposits using underground access. The selection of excavation methodology is based on the nature of orebody and country rock as well as the expected production demand. The suitability of excavation methodology for different combinations of nature of orebodies and wall rock has been outlined in Table 2.1. Different stoping methods have various advantages and disadvantages. Open stoping methods are cost economic, as they do not require additional expenditure on support system. However, minerals are locked in such methods to provide support to the wall rocks and levels. Stoping methods with artificial supports viz. backfill material, timber support, rock bolting are practised to increase the mineral recovery. Timber-based supports are very costly; hence, the applicability of square set stoping is limited. Backfilling using the wastes from mill tailings is in trend nowadays. The applicability of caving-based methods is also difficult. These methods are only applicable when the strength of the wall rock is weak. However, induced caving is also

2.3 Stoping Methods Practised in India

13

Fig. 2.2 Overview of excavation methodologies for coal and metalliferous deposits in India

practised sometimes, when the strength of the wall rock is sufficiently high. Accordingly, these stoping methods are practised for the exploitation of mineral deposits with an aim to optimise between the technical applicability and economic profitability. A summary of these stoping methods being practised at different underground mines in global scenario is given in Table 2.2. The dominant excavation methodology used under Indian conditions is—cut and fill stoping, room and pillar mining and sublevel/longhole/vertical crater retreat (VCR) stoping. Some variants of room and pillar and cut and fill stoping are also employed at different mines in India. The excavation method for these dominant methods with the ways to estimate the charge factor has been discussed in various subsections.

2.3.1 Sublevel Stoping In this method of stoping, sublevels are driven between two main levels. These sublevels work as drivages. The number of such drivages at a particular sublevel may be more than one, depending upon the width of the orebody. These sublevels are connected by driving raises. These raises may work as manway, ore pass or

14

2 Overview of Underground Metalliferous Mining

Table 2.1 Summary of mining method suitability for underground excavations in hard rock (Deshmukh 2009; Hartman 2002; Hustrulid and Lu 2003) Stoping methods

Applicability criteria Thickness of the orebody

Strength of ore

Dip of orebody

Strength of wall rocks

Open stoping with Thin or thick casual pillars

Weak or strong

Flat

Strong

Room and pillar stoping

Thin or thick

Strong

Flat

Strong

Room and pillar longwall

Thin or thick

Weak or strong

Flat

Strong

Cut and fill stoping Thin

Weak or strong

Flat

Weak or strong

Post-pillar stoping Thick

Weak or strong

Flat

Weak or strong

Sublevel open stoping

Thin or thick

Weak or strong

Steep

Strong

Sublevel stoping and their variants with backfilling

Thin or thick

Weak or strong

Steep

Weak or strong

Shrinkage stoping

Thin or thick

Strong

Steep

Strong

Square set stoping

Thin or thick

Weak

Flat

Strong

Top slicing

Massive

Weak or strong

Flat

Weak

Sublevel caving

Massive

Weak or strong

Steep

Weak

Block caving

Massive

Weak or strong

Flat or Steep

Weak

ventilation network. The blocks for the extraction of orebody are also divided by different raises. One of the raises which work as ore pass has chute in the bottom. This chute works for the ore transfer by gravity. Chute is opened to dislodge the excavated ore directly on the dumpers/tubs. Chute arrangement is made on the main level connected with the primary access to the mine (shaft/incline/decline). One “slusher drift level” is also developed above main level and below the sublevels. This slusher drift level is connected by funnel shaped “finger raises”. The excavated ore at sublevels is put inside these finger raises, and they get transferred to the slusher drift level. Ore is further transferred from this drift level to main level using chute. Once the development work for sublevel stoping is over, the final excavation of the ore starts. Initially, a slot is opened either in the middle of the orebody or on one side of the orebody. The excavation starts from this slot and retreats towards the drivages. The ore excavation may be done at different sublevels simultaneously to achieve the productivity. However, the designing of proper support system is important. Sometimes, ore may be left as pillars to provide stability to the overlying strata. The sequence of stoping operations for sublevel stoping method is shown in Fig. 2.3.

2.3 Stoping Methods Practised in India

15

Table 2.2 Summary of stoping methods practised at different underground mines of world S. No.

Stoping method

Name of Mine

References

01

Room and pillar stoping

• Mineral Hill Mine, Jardine, Montana, USA • Immel mine, East Tennessee, USA

• Kahraman (2002) • Hustrulid et al. (2001)

02

Cut and fill mining

• Nellore Mica belt, India • Balaghat underground manganese mine, India

• Tewari et al. (2020) • Manekar et al. (2017)

03

Post pillar stoping

• Mosaboni mine, India

• Singh et al. (1995)

04

Sublevel open stoping

• Salaman Gold Deposit, Spain

• e Faria et al. (2022)

05

Sublevel stoping

• Rampura Agucha Mine, India • Zawar Group of Mine, India

• Himanshu et al. (2021a) • Barton and Pandey (2011)

06

Shrinkage stoping

• Gold Road Mine, Arizona, USA • Nixen Fork Mine, Alaska, USA

• Silver (1997) • Kral (1997)

07

Square set stoping

• Homestake Gold Mine, South Dakota, USA • Bunker Hill Silver Mine, Idaho, USA

• Hartman and Mutmansky (1987) • Songstad (1982)

08

Top slicing

• Negaunee Mine, Michigan, USA • Rožná Uranium mine, Czech Republic

• Russell, (2007) • Sasaoka et al. (2015)

09

Sublevel caving

• The Ridgeway SLC Gold • Brunton et al. (2010) operation—Australia • Zhang and Wimmer (2018) • Malmberget mine—Sweden • Chengchao Iron Mine—Hubei • Cao et al. (2016) Province, China

10

Block caving

• El Teniente Mine—Chile • Tongkuangyu mine—China

• Palma and Agarwal (1973) • Xia et al. (2021)

2.3.2 Longhole Stoping Longhole stoping is a high production and low-cost mining method for metalliferous mine. It is a very popular method chosen when open pit mining activities are no longer economical and mines move to underground operations. This method is a variant of sublevel stoping in which longer blastholes with larger diameters (100–165 mm) are used. Depth of hole may reach up to 100 m. In this method, a vertical slot is created at one end of the stope and then works in the sublevels to drill a radial pattern of drill holes. This radial pattern of drill holes is known as ring/fan drilling (Kushwaha et al. 2014).

16

2 Overview of Underground Metalliferous Mining

(a) Sectional view of Development for a sublevel stoping method

(b) Development in sublevel stoping method to extend opening raise from hangingwall to footwall Fig. 2.3 Sequence of operations for sublevel stoping method

After a set of these holes are loaded, blocks of ore body are blasted into open stope. There are two main variations of this method as longitudinal longhole stoping and transverse longhole stoping. The advancement of stoping operation takes place along the strike of the orebody in longitudinal longhole stoping method. Accordingly, the stope progress takes place between two drivages. In transverse longhole stoping method, the stoping operation advances along dip of the orebody. The operation advances from hangingwall to footwall side along the crosscut driven between

2.3 Stoping Methods Practised in India

17

(c) Progress of stoping operation from opening raise to Manway Fig. 2.3 (continued)

hangingwall and footwall drivages. The diagrammatic comparison of these two variants of longhole stoping method is shown in Fig. 2.4. The selection of longitudinal or transverse longhole stoping method is significantly based on the width of the orebody. The transverse longhole stoping is approached for the wider orebody, whereas the longitudinal longhole stoping method is used for narrow orebodies. In addition to this, the strength properties of the rock are also a deciding factor for the selection of different variants of longhole stoping method. The optimum thickness of the orebody to employ transverse longhole stoping is based on the strength properties of the host rock (Himanshu et al. 2021b). In longhole stoping method, the main access to the orebody is made through a shaft or decline. The main access is connected to footwall drive through cross-cuts on each stoping level. The orebody under excavation is exposed by driving drivages as footwall drive and extraction drive (Himanshu et al. 2021b). The longhole stoping method is practised extensively at different Lead–Zinc underground mines, Gold mine and Copper mines in India. At Rampura Agucha underground Lead–Zinc mine, for longhole stoping, the cross-cuts are made at the interval of 30 m each. The first 15 m spans between two cross-cuts are excavated as primary stope, and the next 15 m are left as a pillar. The left pillars are excavated as secondary stope after the completion of extraction and subsequent backfilling of the primary stope. Excavated stopes are backfilled using cemented rock fill (CRF) or paste fill. The excavation between the cross-cuts is made by opening slot raises (Himanshu et al. 2021a). The plan and sectional view of the mining method used at this mine is shown in Fig. 2.5.

18

2 Overview of Underground Metalliferous Mining

Fig. 2.4 Diagram showing difference between longitudinal and transverse longhole stoping method (Kushwaha et al. 2014)

Fig. 2.5 Plan and sectional view showing the sequence of stope excavation at Rampura Agucha underground Lead–Zinc mine (Himanshu et al. 2021b; Roy et al. 2022)

2.3 Stoping Methods Practised in India

19

Fig. 2.6 Schematic layout for stope excavation at Rampura Agucha underground mine (Himanshu et al. 2021a)

The stoping operation for extraction of minerals at this mine is executed by drilling-blasting technology. The complete operation of stope extraction includes the opening of slot raise, slot raise expansion and ring blasting. The initial excavation in a stope is made by opening a slot raise. The slot raises are opened through the cross-cuts. Since the slot raise is an initial opening in the stope, it lacks sufficient space for the movement of the blasted rock mass. Accordingly, the free face in the slot raise blast design is provided using relief/reamer holes. Once the excavation for slot raise is completed, the blastholes of slot raise expansion and ring are blasted along the respective free face. Sometimes, the initial excavation from the lower level to the upper level of a mine is performed to make the space for the accumulation of blasted rock in the subsequent stages of excavation. The operation is termed sliping. The excavation of sliping holes is done by uphole drilling and charging (Himanshu et al. 2021a). The schematic layout of stope excavation operation at Rampura Agucha underground mine is shown in Fig. 2.6. The major variations in drilling and blasting pattern are possible in this stoping method. Different methods adopted under Indian geomining conditions have been dealt in subsequent chapters. For longhole stoping at Malanjhkhand underground copper mine, slot is opened at the hangingwall and footwall end. Slot is excavated using drop raise method. Once the slot is excavated, the slot expansion holes are blasted. The blasting of slot expansion holes takes place along the excavated slot raise. The firing of ring holes takes place along the excavated slot. Accordingly, the movement of the blasted rock takes place along both the directions, i.e. along hangingwall as well as footwall side. A schematic of excavation pattern practised at Malanjhkhand underground mine is shown in Fig. 2.7. At Malanjhkand underground copper mine, a level interval of about 50 m is made. About 16 m from lower level to upper level is excavated as trough. Rest 35 m is drilled as main ring from upper level to lower level. Trough holes are of small diameter of 34 mm. Small diameter cartridges or compressed ANFO explosive is used for the firing of trough holes. Once the trough holes are blasted, mouth of the main ring holes get opened. The holes of main rings have hole diameter of 115 mm. These holes are charged with cartridge explosives of 83 mm diameter or ANFO. Sometimes, a ring is also blasted in two parts, in order to reduce maximum charge per delay (MCPD). Section of drilling pattern used at this mine is shown in Fig. 2.8.

20

2 Overview of Underground Metalliferous Mining

Fig. 2.7 Schematic drill design for stope excavation at an Indian copper mine

2.3.3 Room and Pillar Stoping This method of stoping is very common for ore bodies with lesser width and having mild gradient. In this method of stoping, raises are driven between two levels. Raises work for different purposes including transport of materials, ventilation and layout of water pumps. Sill pillar is left up to certain extent from lower level, and stope is widened along both sides from the raises. Sufficient crown pillar is left to support upper level. Rib pillar is left between two stopes. Dimensions of sill pillar, crown pillar and rib pillar depend on characteristic of orebody and surrounding rock strata. Excavated stope is supported with the help of roof bolts or rock bolts. Diagrammatic layout of room and pillar stoping method is shown in Fig. 2.9. This method of mining is used for exploitation of narrow width ore deposits of Copper and Uranium in India. The method is extensively used in Singhbhum zone in Jharkhand, India. The drilling-blasting process in this method of mining consists of drifting and production drilling. The initial drifts are driven to open the stope using horizontal drilling. The stoping operation progresses with horizontal or inclined/uphole drilling of the blastholes. The holes are drilled in rectangular/scattered pattern. The horizontal/vertical openings work as the free face for the blastholes. In Singhbhum zone of mines, the drilling operation is carried out using jack hammer drills. The charge factor for the blastholes is decided based on the hardness and massiveness of the rock strata. Charge factor in the range of 0.55–0.8 kg/m3 is used for the exploitation of rock mass at Singhbhum group of mines. The dimensional parameters for drifting operation are similar to that of the tunnelling, whereas for the production blasts, the dimensional parameters viz burden and spacing are taken in the line of openpit excavations.

2.3 Stoping Methods Practised in India

Fig. 2.8 Section of drill pattern used at Malanjhkhand underground copper mine

Fig. 2.9 Plan and sectional views of room and pillar stoping method

21

22

2 Overview of Underground Metalliferous Mining

2.3.4 Cut and Fill Stoping In this method of stoping, raises are driven between two levels. Sufficient sill pillar is left above lower level to support lower level. A sill drive is opened in the region above sill pillar; to open the stope, heightening of sill drive is further done as a part of extraction of material up to designed height. Excavated part is backfilled with sand filling material to work as platform for further excavation. This cut and fill operation propagates towards upper level. Crown pillar is left below upper level to support it. Ore passes are made which may be equipped with chute at the bottom to transfer material at loading level by gravity. Design of sill pillar and crown pillar depends on rock mass strength parameters of rock and accordingly stress analysis due to excavation. Raises serve the purpose for ventilation, and sometimes, it may also be equipped to serve as manways. Sometimes, pillars are left in this method when ore body is wider comparatively. This variation of cut and fill stoping is called post-pillar stoping. Design of pillar dimension is also done on the basis of stress analysis. Diagrammatic layout of cut and fill stoping method is shown in Fig. 2.10. This method of stoping is used for excavation of steeply inclined ore deposits of Copper, Uranium, Manganese, Mica, etc. in India. The drilling-blasting process in this method of mining also consists of drifting and production drilling. The drifting is carried out using the horizontal drilling. The stoping operation progresses with horizontal or inclined/uphole drilling of the blast holes. The holes are drilled in rectangular/scattered pattern similar to that of room and pillar stoping method.

Fig. 2.10 Diagrammatic layout of cut and fill stoping method

References

23

2.4 Summary An overview of the underground metalliferous mining has been presented in this chapter. The summary of the discussions made in this chapter is as follows: • Room and pillar stoping, cut and fill stoping, sublevel stoping and sublevel caving methods are used for the exploitation of mineral deposits in India. • The variants of sublevel stoping are used for the faster pace of mineral exploitation. This method has advantage of large diameter and deep hole drilling and blasting. • The selection of a stoping method is done on the basis of technical suitability and cost economics.

References Barton N, Pandey SK (2011) Numerical modelling of two stoping methods in two Indian mines using degradation of c and mobilization of ϕ based on Q-parameters. Int J Rock Mech Min Sci 48(7):1095–1112 Brunton ID, Fraser SJ, Hodgkinson JH, Stewart PC (2010) Parameters influencing full scale sublevel caving material recovery at the Ridgeway gold mine. Int J Rock Mech Min Sci 47(4):647–656 Cao S, Song W, Deng D, Lei Y, Lan J (2016) Numerical simulation of land subsidence and verification of its character for an iron mine using sublevel caving. Int J Min Sci Technol 26(2):327–332 Carvalho FP (2017) Mining industry and sustainable development: time for change. Food Energy Secur 6(2):61–77. https://doi.org/10.1002/fes3.109 Chugh YP (2018) Concurrent mining and reclamation for underground coal mining subsidence impacts in China. Int J Coal Sci Technol 5(1):18–35. https://doi.org/10.1007/s40789-018-0189-2 Deshmukh DJ (2009) Elements of mining technology, vol. 2. Denett & Company Dineva S, Boskovic M (2017) Evolution of seismicity at Kiruna Mine, 125–139. https://doi.org/10. 36487/ACG_rep/1704_07_Dineva e Faria MF, Dimitrakopoulos R, Pinto CLL (2022) Integrated stochastic optimization of stope design and long-term underground mine production scheduling. Resources Policy 78:102918 Hartman HL (2002) Introductory mining engineering. Wiley Hartman HL, Mutmansky JM (1987) Introductory mining engineering. John Wiley & Sons, New Jersey Himanshu VK, Mishra AK, Roy MP, Vishwakarma AK, Singh PK (2021a) Numerical simulation based approach for assessment of blast induced deformation pattern in slot raise excavation. Int J Rock Mech Min Sci 144:104816. https://doi.org/10.1016/j.ijrmms.2021.104816 Himanshu VK, Roy MP, Shankar R, Mishra AK, Singh PK (2021b) Empirical approach based estimation of charge factor and dimensional parameters in underground blasting. Min Metall Explor 38(2):1059–1069. https://doi.org/10.1007/s42461-020-00374-8 Hustrulid W, Lu W (2003) The Lu-Hustrulid approach for calculating the peak particle velocity caused by blasting. In: Explosives and blasting technique. Taylor & Francis, pp 291–300. https:// doi.org/10.1201/9781439833476.ch36 Hustrulid WA, Hustrulid WA, Bullock RL (eds) (2001) Underground mining methods: Engineering fundamentals and international case studies. SME, 83–87 Kahraman SA˙IR (2002) Estimating the direct P-wave velocity value of intact rock from indirect laboratory measurements. Int J Rock Mech Min Sci 39(1):101–104 Kral S (1997) Mining industry beginning to rediscover Alaska. In: Mining engineering, 49

24

2 Overview of Underground Metalliferous Mining

Kushwaha A, Himanshu VK, Sinha A (2014) Design of stoping parameters and support system for long-hole stoping method by numerical modelling. In: Proceedings of 5th Asian mining congress and exhibition, Kolkata Manekar GG, Shome D, Chaudhari MP (2017) Prediction of subsidence parameters & 3-D analysis at Balaghat underground manganese mine of MOIL limited, India. Procedia Eng 191:1075–1086 Mborah C, Bansah KJ, Boateng MK (2015) Evaluating alternate post-mining land-uses: a review. Environ Poll 5(1):14. https://doi.org/10.5539/ep.v5n1p14 Newman AM, Martinez M, Kuchta M (2006) A review of long- and short-term production scheduling at Lkab’s Kiruna Mine. In: Herrmann JW (eds) Handbook of production scheduling. International series in operations research & management science, vol 89. Springer, Boston, MA. https://doi.org/10.1007/0-387-33117-4_11 Palma R, Agarwal R (1973) A study of the cavability of primary ore at the El Teniente Mine. Technical Report from Colombia University, New York, (niepublikowane) Raina AK, Murthy VMSR, Soni AK (2014) Flyrock in bench blasting: a comprehensive review. Bull Eng Geol Env 73(4):1199–1209. https://doi.org/10.1007/s10064-014-0588-6 Roy MP, Himanshu VK, Kaushik AP, Singh PK (2022) Influence of ring blasting pattern on the safety of nearby underground structures. S¯adhan¯a 47(4):192. https://doi.org/10.1007/s12046022-01968-2 Russell L (2007) Negaunee Mine, top slicing operation Sasaoka T, Hamanaka A, Wahyudi S, Shimada H, Kebo V (2015) Numerical study on effect of rock mass stability on deep deposit slicing at uranium mine in Czech Republic. Electron J Geotech Eng 20(12):4379–4386 Silver DB (1997) Gold road mine: anatomy of a turnaround. Min Eng 49(8):28–32 Singh UK, Jain PN, Prasad M (1995) Post-pillar behaviour at deep levels in a copper mine. Int J Rock Mech Min Sci Geomech Abstracts 32(6):585–593 Singh PK, Roy MP, Paswan RK, Dubey RK, Drebenstedt C (2015) Blast vibration effects in an underground mine caused by open-pit mining. Int J Rock Mech Min Sci 80:79–88. https://doi. org/10.1016/j.ijrmms.2015.09.009 Songstad J (1982) Square-set timber in load-haul-dump stopes at the Bunker Hill Mine, Kellogg, Idaho. In: Underground mining methods handbook, pp 744–748 TERI (2019) Exploration and mining in India: time for a deeper look. The Energy and Resources Institute, August, 1–40 Tewari S, Himanshu VK, Porathur JL, Bhattacharjee R, Das AJ, Mandal PK (2020) Exploitation of mica deposits at Nellore mica belt, Andhra Pradesh, India. Curr Sci 118(4). https://doi.org/ 10.18520/cs/v118/i4/593-602 Verma S, Chaudhari S (2017) Safety of workers in Indian mines: study, analysis, and prediction. Saf Health Work 8(3):267–275. https://doi.org/10.1016/j.shaw.2017.01.001 Whittle D, Brazil M, Grossman PA, Rubinstein JH, Thomas DA (2018) Combined optimisation of an open-pit mine outline and the transition depth to underground mining. Eur J Oper Res 268(2):624–634. https://doi.org/10.1016/j.ejor.2018.02.005 Xia ZY, Tan ZY, Zhang L (2021) Instability mechanism of extraction structure in whole life cycle in block caving mine. Geofluids Zhang ZX, Wimmer M (2018) A case study of dividing a single blast into two parts in sublevel caving. Int J Rock Mech Min Sci 104:84–93

Chapter 3

Rock–Explosive Interaction During Underground Blasting

Abstract The rock–explosive interaction behaviour influences the blasting outputs in terms of breakage, fragmentation and induced hazards. The blast designers can optimize the design parameters by assessment of rock–explosive interaction to achieve the desired output. The expected outcomes from underground blasting would be different under varying geomining conditions. The optimum breakage of rock mass with enhanced pull and reduced overbreak is expected while underground drivages blasting. The blasting under box-cut excavation scenario comes with the expectation to achieve the optimum pull and maximized movement. The expectations from the production blasting are to optimize the rock fragmentation. The geological conditions viz. presence of joints, shear zone, etc. also influence the blasting outputs. In these cases, the design parameters need to be optimized using the assessment of rock–explosive interaction behaviour. This chapter presents a theoretical overview of rock–explosive interactions.

3.1 Introduction The quality of a blast is measured in three significant aspects as—extent of rock mass damage, rock fragmentation and blasting hazards (Silva et al. 2019). The blast design parameters are optimized to improve the quality of a blast in these aspects. The review of the literatures on rock–explosive interactions suggests that the rock parameters, explosive parameters and mining conditions influence the blasting outputs (Singh et al. 2016). Different empirical and numerical approaches have been used to identify the impacts of these parameters on blasting quality over the years (Zhang et al. 2020; Gorai et al. 2021). Most of these approaches are for openpit excavations. The concerns relating to the quality of a blast for underground excavation would be different. Despite different concerns under varying geomining conditions, the science behind rock breakage under dynamic/blast loading would be the same. So, the discussions have been made in this chapter on the rock breakage mechanism. The various theories concerning the rock breakage mechanism have also been discussed. Under the impact of explosive energy, the rock mass experiences a high impact of more than 4000 m/s (Gómez et al. 2020; Leng et al. 2021; Chiquito et al. 2019). © The Author(s), under exclusive license to Springer Nature Singapore Pte Ltd. 2023 V. K. Himanshu et al., Blasting Technology for Underground Hard Rock Mining, https://doi.org/10.1007/978-981-99-2645-9_3

25

26

3 Rock–Explosive Interaction During Underground Blasting

The deformation behaviour of rock mass under such impact loading would be plastic or hydrodynamic. The physical behaviour of any engineering material under a set of loading conditions is described by constitutive models. Different constitutive models have been used by the researchers over the years to evaluate the rock–explosive interaction behaviour under different practical situations (Baranowski and Małachowski, 2018; Yang et al. 1996). Some of these constitutive models have also been used in numerical simulation in successive chapters. The damages in rock mass are analysed using these constitutive models. It is analysed with various descriptors viz. induced strain, factor of safety, damage index, etc. (Coggan et al. 2012; Gui et al. 2017; Hu et al. 2015; Ma and An 2008; Zhang et al. 2020; Shukla and Mishra, 2020; Liu et al. 2016; Tan et al. 2020; Maji and Adugna 2016). Researchers like Arora and Dey (2010), Gómez et al. (2020), Hustrulid and Lu (2003), Yilmaz and Unlu (2014), etc. have correlated damage with nearfield peak particle velocity (PPV). Rock parameters, explosive parameters and blast design parameters have the foremost impact on nearfield PPV. This chapter includes the discussions regarding the rock breakage mechanism, assessing the influences of rock and explosive parameters on rock breakage, and the constitutive models for rock breakage under dynamic loading. The explosive interaction behaviour during underground blasting has been summarized in the end of this chapter.

3.2 Rock Breakage Mechanism Blast-induced rock mass breakage consists of three major steps as—shock/stress wave-induced cracking, crack propagation and rock movement by gas pressurization. Damage/cracking induced by shock/stress wave in the immediate adjacent to the borehole results from crushing of borehole wall due to extremely high detonation pressure (ISEE Blasters’ Handbook 2011). It is found that the development of cracks and damage has a significant time lag behind the shock wave. Shockey et al. (1974) stated that the crack propagation speed is three times slower than the shock wave velocity. The pressure exerted on the borehole wall induces compressive shock waves in the rock mass. The wave gets reflected after reaching free face/discontinuity. This reflected wave is termed as secondary wave. The secondary wave may be compressive or tensile in nature, depending on the difference in characteristic acoustic impedance of the interface. There is an interface between rock and air when the stress wave gets reflected from the free face. The nature of the secondary wave in such a case will be tensile. The fracturing in rock mass occurs when the superimposed values of the incident compressive stress and reflected tensile stress exceed the tensile strength of the rock mass. However, the strength of the rock mass also changes with variations in the strain rate (Prasad et al. 2000). So, considering the strain rate dependency, the rock fracturing will occur when the superimposed values of the incident compressive stress and reflected tensile stress would exceed the characteristic strength of the rock, which is the result of strain rate-dependent response at that point under the existing

3.3 Rock Mass Damages Under Dynamic Loading

27

Fig. 3.1 View of blast-induced stress wave propagation through rock media (Pal 2015)

state of stress. A schematic diagram of rock breakage mechanism under the influence of stress wave is shown in Fig. 3.1. Bhandari (1997) classified the zones of influence in the rock mass under the action of the stress wave. The zones were crush zone, fracture zone and seismic zone. The crush zone is expected up to 4 times the radius of the blasthole. The fracture zone is up to 50 times the radius of the blasthole. The zone beyond the fracture zone is considered the seismic zone. Apart from shock/stress waves, explosive detonation also produces high-pressure gaseous products at very high temperatures. These high-pressure products also play a key role in the fracturing of the rock mass. Researchers such as Ash (1963), Kutter and Fairhurst (1971), Langefors and Kihlstrom (1963), Persson et al. (1970), Porter and Fiarhurst (1970), Sengani (2020), etc. discussed the role of gaseous pressure in rock breakage mechanism. As per literature, the high-pressure gaseous detonation products flows through the pre-existing cracks and widens them, thereby enhancing the fragmentation. Dally et al. (1975) observed through the experimental trials that the gas pressure helps in extending cracks in the vicinity of the blasthole. The findings suggested that the containment of gaseous products increases the amplitude of the tensile wave and results in more fractures and fracture length. Bhandari and Vutukuri (1974) found in the experimental trials that the fractures near the blast holes are formed by quasi-static gas pressure, and the fractures near the free face are formed by stress wave reflection.

3.3 Rock Mass Damages Under Dynamic Loading The physical response of the material under varying loading conditions is described by constitutive models. The rock mass behaviour under static and dynamic loading has also been represented by different constitutive models. Selection of a proper constitutive model to study the breakage mechanism of rock mass under dynamic

28

3 Rock–Explosive Interaction During Underground Blasting

Table 3.1 Deformation behaviour of materials under different loading (Jaiswal 2018) Impact velocity (m/s)

Strain rate (/s)

Deformation behaviour

< 10–5

Static/creep

< 50

10–5

50–1000

10–1 to 101

Elasto-plastic

1000–3000

101 to 106

Primarily plastic (pressure level ≥ material strength)

3000–12,000

106

Hydrodynamic

> 12,000

> 108

to

to

10–1

108

Elastic

Vaporization of colliding solids

loading is very difficult, as the rock mass shows variations in mechanical behaviour under such loading. A brief about the deformation behaviour of materials under different impact loading is shown in Table 3.1 (Jaiswal 2018). The table shows that using detonation velocity and strain rate of the explosive, the material behaviour of rock mass under blast loading can be taken as primarily plastic or hydrodynamic. However, minimal literature is available on the constitutive models to assess the physical behaviours of rock mass under dynamic loading. But, significant developments have been made in the constitutive models for concrete and other similar materials. Cui et al. (2017) discussed about the different constitutive theories of concrete and emphasized that it shows theories of—elasticity, plasticity, damage, visco-plasticity and fracture mechanics under the dynamic loading. A stress–strain plot is a fundamental way to study the physical behaviour of rock under a defined loading condition. The rock shows proportional relation between stress and strain up to the attainment of yield strength under uniaxial compression loading conditions. The rock behaviour is ductile up to the yield strength. Its behaviour changes to brittle from ductile after the attainment of the strength of the rock. There is a decrement in stress with the increasing strain after attaining the uniaxial compressive strength (UCS) of the rock. The rock shows inelastic or plastic behaviour when it is loaded further after the yield point. The state of rock when the stress–strain curve continues to rise after the yield stress, but with a shallower slope is known as the plastic behaviour with strain hardening. Such rock material shows unloading behaviour along the curve parallel to the original stress–strain curve. The rock that shows no increment in stress with increasing strain after attaining yield strength is elastic-perfectly plastic. The stress–strain behaviour under the unloading condition in this rock will be parallel to the elastic stress–strain behaviour. A material is considered as rigid-perfectly plastic if it shows no increment in strain with loading conditions. However, the stress will change with the cumulative loading. The unloading behaviour will be parallel to the stress axis for this type of material. The rock behaviour under high strain rate loadings would be different from that due to static loading. Researchers viz. Beser and Aydiner (2019), Gong et al. (2019a, b), Pan et al. (2021), Renjie et al. (2020), Wen et al. (2020), etc. have studied the rock behaviour under high strain rate loadings. The stress–strain plot from the findings of Gong et al. (2019a, b) is shown in Fig. 3.2. The plot shows the rock behaviour under uniaxial compressional loading conditions with different strain rates. It can be

3.3 Rock Mass Damages Under Dynamic Loading

29

Fig. 3.2 Stress–strain behaviour of rock in uniaxial compression and with different strain rate loadings (Gong et al. 2019a, b)

drawn from this plot that UCS of rock at the strain rate of 135/s is about 2.5 times of UCS at the strain rate of 45/s. Apart from the uniaxial compression, the rock mass may be subjected to biaxial or tri-axial loading conditions. The behaviour of rock and associated stress–strain plot would be different under confined loading conditions. The confined strength of the rock would be higher than UCS. The amplification in strength would be dependent on the magnitude of confinement. The stress–strain relation is obtained from the response of a material under different mechanical and/or thermal loading conditions. This stress–strain relation, together with the conservation laws and kinematic relations, constitutes the constitutive models. These constitutive models can be divided into equation of states (EOSs), strength models and failure models. The EOSs comprise the relation between the pressure to volume and internal energy or temperature. The strength models relate the deviatoric stress to the deviatoric strain. Additionally, a failure criterion is required to identify the onset and describe the evolution of material failure (Wu et al. 2017). Babu et al. (2010) classified the constitutive models for the concrete material as linear elastic models, nonlinear elastic models and plasticity-based models. Plasticity-based models follow the yield criterion, flow rules or hardening rules. The yield criteria are pressure-dependent and pressure-independent. The standard pressure-independent yield criterion-based plasticity constitutive models are— Tresca or maximum shear stress criterion, von Mises or maximum distortion energy criterion, Rankine or maximum tensile stress criterion, Mohr–Coulomb criterion, etc. The pressure-dependent plasticity yield criteria are—Drucker–Prager criterion, Mises-Schleicher criterion, etc. (Babu et al. 2005).

30

3 Rock–Explosive Interaction During Underground Blasting

3.3.1 Constitutive Models Used for Numerical Simulation of Rock Blasting Researchers have used different constitutive models for predictions of blasting outputs under practical conditions. A summary of some of such research works is given in Table 3.2. These research works included simulation for predictions of vibration and rock fragmentation.

3.4 Rock Parameters and Their Role on Damage Characteristics The discussions on constitutive models show that the rock parameters affect the rock mass damage characteristic under dynamic loading. The impacts of rock parameters have also been assessed in various research works using experimental field data. Most of these research works have presented rock mass damage as a function of nearfield PPV. In this context, the critical vibration limit to initiate the rock mass damages has been proposed by researchers under varying conditions. Bauer and Calder (1970) have defined the peak particle velocity (PPV) limits of the rock mass for initial fracturing, tensile slabbing, radial cracking and complete breakup. The PPV limits for different nature of fracturing in the rock mass are shown in Table 3.3 (Silva et al. 2019; Bauer and Calder 1970). The suggested limits by Bauer and Calder (1970) are based on the nature of fracturing due to blast vibration. However, the inconsistency in the fracturing in different rock types has not been discussed under this classification. The literature suggests the P-wave velocity of the rock mass as one of the suitable parameters to define the rock mass damage under the influence of detonation wave. Forsyth (1993) suggested a relationship of critical PPV with P-wave velocity (Silva et al. 2019; Forsyth 1993). The suggested relationship is based on the general law of elasticity, under which the PPV and P-wave velocity of the rock mass has been considered as the change in velocity and original velocity to compute the strain. The damage has been considered against the tensile strength of the rock mass. The relationship correlating the rock properties with critical PPV is shown in Eq. 3.1. PPVcritical =

σt Vp Y

(3.1)

where PPVcritical = Critical peak particle velocity limit to initiate damage in the rock mass σ t = Tensile strength of the rock Y = Youngs modulus of elasticity of the rock V p = P-wave velocity of the rock.

3.4 Rock Parameters and Their Role on Damage Characteristics

31

Table 3.2 Summary of the constitutive model used by researchers in numerical simulation of rock blasting Research work

Title

Constitutive model

Simulation code

Li et al. (2017)

Numerical simulation of blast Plastic kinematic vibration and crack forming strengthened constitutive effect of rock-anchored beam model excavation in deep underground caverns

Dynamic FEM

Wang et al. (2018)

Johnson–Holmquist-II(JH-2) Johnson–Holmquist-II(JH-2) constitutive model for rock constitutive model materials: parameter determination and application in tunnel smooth blasting

FEM

Shahrin Numerical simulation of rock Elastic et al. (2019) fragmentation by blasting using discrete element method and particle blast method

DEM

Baranowski Experimental testing and et al. (2020) numerical simulations of blast-induced fracture of dolomite rock

FEM and SPH

Yang et al. (2020)

Johnson–Holmquist-II(JH-2) constitutive model

Numerical simulation of Elastic blasting in confined fractured rocks using an immersed-body fluid–solid interaction model

Experimental and numerical He and Yang (2019) investigations of dynamic failure process in rock under blast loading

Johnson–Holmquist constitutive model

DFN

FEM

Gao et al. (2019)

Effect of initiation location User-defined damage model on distribution and utilization of explosion energy during rock blasting

FEM

Trivino and Mohanty (2015)

Assessment of crack initiation Visco-elastic Kelvin model and propagation in rock from explosion-induced stress waves and gas expansion by cross-hole seismometry and FEM–DEM method

FEM-DEM

Hu et al. (2015)

Numerical simulation of the Tensile-compressive damage complete rock blasting model response by SPH-DAM-FEM approach

SPH-DAM-FEM

(continued)

32

3 Rock–Explosive Interaction During Underground Blasting

Table 3.2 (continued) Research work

Title

Constitutive model

Simulation code

Pramanik and Deb (2015)

Implementation of smoothed particle hydrodynamics for detonation of explosive with application to rock fragmentation

Graddy and Kipp damage model

SPH

Himanshu Numerical simulation-based et al. (2021) approach for assessment of blast-induced deformation pattern in slot raise excavation

Elastic

FEM Explicit Dynamics

Himanshu Explicit dynamics-based et al. (2022) numerical simulation approach for assessment of impact of relief hole on blast-induced deformation pattern in an underground face blast

Elastic

FEM Explicit Dynamics

FEM Finite element method; DEM Discrete element method; DFN Discrete fracture network; SPH Smooth particle hydrodynamics

Table 3.3 Vibration limit for different nature of fracture in the rock mass (Bauer and Calder 1970)

PPV (mm/s)

Effect

< 250

No fracture of intact rock

250–635

Minor tensile slabbing will occur

635–2540

Strong tensile and some radial cracking

> 2540

Complete breakup of rock mass

Although there are merits of this P-wave velocity-based damage assessment method, Fleetwood et al. (2009) have highlighted some of the major constraints of this method. The main constraint is the concept that considers the variations in the strength of the rock by changing the loading rate. This consideration has not been taken into account under this criterion. Despite these limitations, it is clear from this criterion that the rock mass damage is a function of tensile strength, P-wave velocity and elastic modulus. Researchers have also correlated other rock properties, viz. elastic modulus, Poisson’s ratio, density, RQD, RMR, GSI, etc., with PPV. The statistical algorithms were used in these research works to identify the influence of these parameters. The summary of research work that identifies the influence of rock parameters on PPV is given in Table 3.4.

3.4 Rock Parameters and Their Role on Damage Characteristics

33

Table 3.4 Summary of research work correlating rock parameters with PPV Study

Input rock parameters

Technique

R2

Hajihassani et al. (2015a)

P-wave velocity, Young’s modulus of elasticity

ICA-ANN

0.97

Hajihassani et al. (2015b)

Rock quality designation (RQD)

PSO-based ANN

0.85

Khandelwal and Singh (2009)

P-wave velocity, Young’s ANN modulus of elasticity, Poisson’s ratio

0.98

Rezaeineshat et al. (2020)

RQD

ICA-ANN

0.90

Amin Shokravi et al. (2018)

Rock mass rating (RMR)

PSO-Linear

0.95

Kumar et al. (2016)

Uniaxial compressive strength (UCS), RQD, geological strength index (GSI)

Statistical Analysis

0.78

Kumar et al. (2014)

Poisson’s ratio

Statistical Analysis

0.907

ICA Imperialist competitive algorithm; ANN Artificial neural network; PSO Particle swarm optimization

3.4.1 Nearfield Vibration and Rock Breakage Characteristic Based on the assumption of rock breakage under critical vibration limit, nearfield vibration can be taken as the important parameter to assess breakage characteristics. The review of vibration propagation characteristics shows that the amplification and attenuation of blast vibration depend on the propagating media. The propagating media can be considered to be homogeneous in the nearfield. So, the assessment of nearfield blast vibration can give the extent of damage under the blast loading. Holmberg and Persson predictor is one of the globally accepted nearfield vibration predictors. The predictor is a nonlinear relationship between distance and explosive charge. The relationship under this predictor is shown in Eq. 3.2. The schematic diagram of the parameters used in this predictor is shown in Fig. 3.3 (Onederra and Esen 2004). ⎡ PPV = K ⎣l

x s +Δx

xs

⎤α dx  2 β/2α ⎦ r0 + (x − x0 )2

(3.2)

where l is the linear charge concentration (kg/m). dx is the element of charge contributing to the PPV at point P. This relation can be simplified by assuming β = 2α. The simplified relation is shown in Eq. 3.3.

34

3 Rock–Explosive Interaction During Underground Blasting

Fig. 3.3 Schematic diagram of Holmberg and Persson parameters to model near field PPV (Onederra and Esen 2004)



PPV = k

l r0





 α Δx + xs − x0 x0 − xs arctan + arctan r0 r0

(3.3)

K and α in these relations are site constants. The site constants are dependent on the nature of rock strata and discontinuity in the strata. This nearfield predictor makes the following presumptions: • • • • •

The blast wave obeys the charge weight scaling law. PPV due to each small element of charge within the blasthole is additive. The velocity of detonation (VOD) of the explosive charge has been neglected. The impact of the free face boundary has been neglected. PPV is proportional to the dynamic strain experienced by the rock mass.

3.5 Influence of Explosive Parameters on Rock Mass Damage The rock breakage pattern is also influenced by explosive parameters viz. explosive density, detonation velocity, detonation pressure, etc. (Jimeno et al. 1987). The choice of suitable explosive as per the strata demand is necessary to get the optimum output from the blast.

3.5.1 Commercial Explosive and Its Principle of Explosion The explosive is a reactive substance that instantaneously releases a huge amount of energy. The energy is released in the form of shock or stress waves and gases. When

3.5 Influence of Explosive Parameters on Rock Mass Damage

35

the explosive is initiated by shock, the molecules of the explosive mass convert into a gaseous substance. These gaseous molecules propagate through the adjacent media. The speed of the first gasified molecule is so high that it does not lose heat through the conductivity to the unreacted zone. Instead, the unreacted zone is initiated by the transmission of shock from the reacted zone. So, once the explosive is initiated, the shock/stress wave is generated, propagating through its own mass. At the same time, the reacting explosive mass releases a large quantity of high-temperature gases. The pressure due to this high-temperature gas is known as secondary pressure. If this secondary pressure acts upon the undetonated explosive part, its effect is added to the primary pressure. The combined effect is capable of transforming the process of deflagration to detonation. If the secondary pressure acts against the primary pressure in the undetonated portion, it slows down the explosive reaction. It may stop the further energization of the rest of the explosive mass, and the explosive may remain undetonated. The variation of pressure concerning the time at different points in an explosive column is shown in Fig. 3.4 (Zhang 2016). In the frame of the leading shock of the detonation wave, gases enter at supersonic velocity and are compressed through the shock. This sudden change in pressure initiates the release of chemical energy. This released energy again accelerates the flow back at sonic speed. The phenomenon leads to the development of a pressure gradient at a plane. This plane is known as Chapman–Jouguet (C–J) plane. A schematic of C–J plane in the detonation of the explosive is shown in Fig. 3.5. The C–J plane moves at a very high speed. The velocity of the C–J plane is known as velocity of detonation (VOD). The speed of movement of explosion products is one-fourth of the VOD (Jimeno et al. 1987). So, the detonation pressure exerted by the explosive is taken as per Eqs. 3.4 and 3.5. Pd = ρe (VOD)Ue

(3.4)

where Pd = Detonation pressure ρ e = Density of explosive VOD = Velocity of detonation of explosive U e = Velocity of explosion products. Pd =

1 ρe (VOD)2 4

(3.5)

3.5.2 Explosive Parameters Based on the principle of explosive detonation, its composition and output behaviour, different parameters of the explosive have been identified. The influence of these parameters on the rock breakage characteristic has been discussed in this section.

36

3 Rock–Explosive Interaction During Underground Blasting

Fig. 3.4 Pressure profile at different locations in a blasthole (Zhang 2016)

Researchers have also studied the interdependency among these explosive parameters (Mishra et al. 2018; Yunoshev et al. 2012).

3.5.2.1

Density

The density of an explosive is defined as the mass of explosive per unit volume. It contributes to the borehole pressure on the wall of the blasthole. It is recommended to use higher density explosives for a denser rock formation. The recommendation is based on the impedance match theory. The higher is the ratio of the characteristic impedance of explosive to that of rock, and the rock breakage will be more. The characteristic impedance for explosive is defined as the product of the explosive density and its velocity of detonation (VOD). The characteristic impedance for the rock is defined as the product of the rock density and its P-wave velocity. The density

3.5 Influence of Explosive Parameters on Rock Mass Damage

37

Fig. 3.5 Schematic diagram showing C–J plane in the detonation of explosive (Jimeno et al. 1987)

of explosives generally lies in the range of 0.5–1.7 g/cc (Jimeno et al. 1987; Mishra et al. 2019).

3.5.2.2

Velocity of Detonation (VOD)

It is defined as the speed with which the detonation wave moves through the explosive column. The different factors which affect the detonation velocity are explosive type, explosive diameter, confinement of explosive, etc. The impedance ratio gets affected dominantly by the velocity of detonation of explosives during rock fracture process. The velocity of detonation generally lies in the range of 1500–6700 m/s. The explosives such as emulsion and slurry, having higher VOD, produce more shock waves upon detonation than explosives with lesser VOD (He and Yang 2019; Brinkmann 1990).

3.5.2.3

Detonation Pressure

It is defined as the pressure associated with the reaction zone of a detonating explosive. Silva et al. (2019) and Torbica and Lapcevic (2014) highlighted the role of detonation pressure in the process of rock damage by blasting.

3.5.2.4

Energy

The explosive energy is the amount of actual expansion work carried out by the explosive. This is an important parameter along with the other explosive parameters.

38

3 Rock–Explosive Interaction During Underground Blasting

It is expected that the high energy explosive will give better performance in rock breaking. The computation of explosive energy is difficult. The most common method of computing explosive energy is an underwater test. Although the values from this test may not give real value while blasting in rock, the heat of explosion is also sometimes considered for computation of explosive energy. However, it is a poor descriptor of the strength of the explosive as the mechanical work performed by the explosive on rock is a better descriptor of the strength (Sanchidrián and López 2006).

3.5.2.5

Sensitivity

It is the property of explosive which defines the ease of explosive to get detonated by fire, impact, shock, heat or any other source. It is related to the requirement of input energy for the detonation of explosives. To enhance the ease of detonation, explosives are categorized as cap sensitive and non-cap sensitive. A standard detonator is required to detonate cap sensitive explosives, whereas a detonator cannot initiate non-cap sensitive explosives, as they require primer or booster with the detonator. The sensitivity of explosives gets affected by different factors such as water, temperature and charge diameter.

3.5.2.6

Flammability

It is the property of explosive which define the ease of explosive to get detonated by spark, fire, heat or flame. In other words, it is the measure of the amount of heat required to cause the burning of explosives instead of detonation. For any explosive, the required minimum level of heat must be achieved to get detonated instead of burning. It is an important property in terms of safety during the storage and transportation of explosives. Nowadays, maximum explosive compounds are inflammable due to alteration in its component, but they must be taken into account for safety concerns (Jimeno et al. 1987).

3.5.2.7

Water Resistance

It is the ability of an explosive to withstand exposure to water without suffering any detrimental effects in its performance. Explosive generally has internal resistance among its composition, which shows hydrophilic or hydrophobic nature while coming in contact with water. The explosives which are hydrophobic in nature can directly be poured into the borehole without any detrimental effect in nature. The explosives, which are hydrophilic in nature, when come in contact with water, start dissolving some of their ingredients and hence decreases their strength. When a blast is to be taken in the watery blastholes, a good and fair water-resistant explosive must be selected. When blast for sleeping holes needs to be carried out, then also the explosive having good water resistance should be used (Jimeno et al. 1987).

3.5 Influence of Explosive Parameters on Rock Mass Damage

3.5.2.8

39

Fume

The fume is the measure of the amount and type of toxic gases produced after the detonation of the explosive. They are generated due to oxygen dis-balance in the explosive. Fume production is a significant concern in underground mines due to insufficient ventilation. Toxic fumes are produced due to inadequate priming, poor water resistance of explosives, lesser confinement, critical charge diameter, etc. (Jimeno et al. 1987).

3.5.3 Equation of State for Explosive The equation of state is a thermodynamic relation that describes the state of a matter under a set of physical conditions. The basic relation for the EOS is shown in Eq. 3.6. The relation states that pressure (P), volume (V ) and temperature (T ) parameters in a reaction are related to each other. Ideal gas law is the basic equation of state. The other EOSs are modifications of this relation under a set of physical scenarios. f (P, V , T ) = 0

(3.6)

The physical properties of the detonation products such as explosives follow the Jones–Wilkins–Lee (JWL) equation of state (EOS) (Artero-Guerrero et al. 2017; Castedo et al. 2018; Hu et al. 2015; Pramanik and Deb 2015 José A. Sanchidrián and López 2006). The EOS is a relationship among pressure, volume and energy. The expression for the JWL equation of state is shown in Eq. 3.7.  



ω ω ωE e−R1 V + B 1 − e−R2 V + P = A 1− R1 V R2 V V

(3.7)

where A, B, R1 , R2 and ω are constants P = Pressure V = Specific volume E = Energy per unit volume. Cylindrical tests are conducted to estimate the parameters of JWL EOS (Baker et al. 2012). Sanchidrián et al. (2015) estimated JWL EOS parameters of the explosives of different densities and diameters and validated the result with the numerical simulation. Castedo et al. (2018) estimated JWL EOS parameters for emulsion explosives of different diameters using cylindrical tests. Davis and Hill (2001) estimated the JWL parameters for Ammonium Nitrate–Fuel Oil (ANFO) explosives using cylindrical tests.

40

3 Rock–Explosive Interaction During Underground Blasting

3.6 Impacts of Blast Design Parameters on Rock Mass Damage The different blast design parameters such as drilling accuracy and free face condition have the foremost impact on the rock breakage. The free face boundary works to reflect the stress wave. The reflected tensile wave works to initiate dislodging of the broken rock mass. The discussions on some of these parameters influencing rock breakage have been made in this section. An outline of different blast design parameters in a bench blasting is shown in Fig. 3.6. For the underground metalliferous mining, the excavated slot works as the free face. The blasting in such cases takes place in ring/fan pattern against the excavated slot. An outline of the different blast design pattern for underground ring blasting is shown in Fig. 3.7.

3.6.1 Blasthole Diameter The charge concentration at a point is controlled by this parameter. Jimeno et al. (1987) have suggested the optimum blasthole diameter as 0.00016 times the depth of the blasthole. The suggestion has been made to optimize the fragmentation, powder factor and cost. Such rules of thumb also consider that the larger blastholes tend to increase deviation. However, the development of the recent drilling equipment has made it possible to drill larger blastholes of small diameter with greater accuracy. Accordingly, the longer blastholes of small diameter are also in practice nowadays. On the other hand, sometimes, the large diameter blastholes are drilled even in the smaller bench height when the rock strata are very hard. This is done to cope up with the inability of the drilling equipment to drill holes of smaller diameter in the harder Fig. 3.6 Outline of different blast design parameters used in bench blasting

3.6 Impacts of Blast Design Parameters on Rock Mass Damage Fig. 3.7 Outline of different blast design parameters for underground ring blasting

Upper level

41

Holes of 2nd Ring Holes of 1st Ring

Burden Burden

Free Face or Slot

Stope

Toe Spacing Lower level

rock formation. So, the selection of blasthole diameter nowadays is mainly based on the strata condition, economy and safety. The chances of deviation of blasthole are more in the case of underground blasting, as the deeper holes with inclination are drilled in such case. The control on drill deviation is achieved by the use of accurate drilling machine. At the same time, the measurement of drill deviation before explosive charging is also important. The charging parameters can be modified to achieve the desired outputs from the blasts, if the deviation in blastholes is known to the blast designers.

3.6.2 Burden Burden is defined as the shortest distance to relief at the time of detonation of holes. Burden also signifies inter-row distance. It is the most critical parameter in any blasting operation. When the burden is very small, problems such as high air blast level, production of excessive fines and the throw of rock to large distances arise. If the burden is too much, it produces severe back break and shattering of back wall (Pal Roy 2007). The optimum burden for a bench blast is determined based on the assessment of rock parameters, explosive parameters and operational parameters. Researchers have proposed some empirical relations to compute the optimum burden

42

3 Rock–Explosive Interaction During Underground Blasting

for bench blasting. The relation proposed by Konya and Walter (1990) is given in Eq. 3.8.

 SGe B= 2 + 1.5 .φe SGr

(3.8)

where B = Burden in ft SGe = Specific gravity of explosive SGr = Specific gravity of rock φ e = Diameter of explosive in inch. The empirical relation for burden proposed by Pal Roy (2005) is given in Eq. 3.9. √ 5.93 Le φe . + (0.37) B=H φ RQD K

(3.9)

where B = Burden (m) H = Bench height (m) φ e = Diameter of explosive (mm) φ = Diameter of blasthole (mm) RQD = Rock quality designation L e = Loading density of explosive (kg/m) K = Charge factor (Kg/m3 ). In case of ring blasting, toe burden/burden is the distance of the ring blastholes from the excavated slot portion. It is also the distance between two rings. Very limited literature is available regarding the selection of optimum burden for ring blasting. Rustan (1992) proposed a relationship between the blasthole diameter and burden. The relationship is shown in Eq. 3.10. B(in m) = 11.8 × φ 0.63

(3.10)

where φ = Blasthole diameter (in m).

3.6.3 Spacing Spacing is defined as the distance between two adjacent holes, measured perpendicularly to the burden and parallel to the free face. Spacing and burden form grid patterns if their values are equal, which is applicable for massive rock breaking. Twisting and tearing of rock is more, and splitting and back break are lesser if the burden is small and spacing is large. If the spacing is smaller than the burden, splitting

3.6 Impacts of Blast Design Parameters on Rock Mass Damage

43

occurs between blastholes and back break is observed. As a general guideline used by Pal Roy (2007), spacing should be 1.2–1.5 times the burden in bench blasting. The term “toe spacing” is used in the case of underground ring blasting. It can be defined as the distance between two blastholes of a ring at the toe point. Rustan (1992) suggested to keep toe spacing as 1.5–2.0 times of the burden for ring blasting.

3.6.4 Stemming Stemming is the inert material (mud, clay, sand or drill cutting) used to cover the hole after the drill hole is packed with explosives. It works to give confinement to the propagating shock wave while blasting. Stemming increases blasting efficiency by enhancing fragmentation. It also reduces air shock waves and high-pressure gases venting into the atmosphere. The preferable stemming material to reduce flyrock ejection is 3–4-mm rock chips.

3.6.5 Subgrade Drilling Drilling of the blastholes beyond the planned grade lines is termed subgrade drilling. The subgrade drilling causes a larger zone of maximum tension, which must be sheared for a good rock breakage (Pal Roy 2007).

3.6.6 Charging Parameters Quantity of explosive charge has a dominant impact on rock breakage patterns. The charging parameters are expressed in different terms as explosive charge per hole, explosive weight per delay and total explosive charge in a firing round. The variations of these parameters affect the quantitative blast output in terms of fragmentation, backbreak and throw, and blasting hazards in terms of ground vibration, air overpressure, etc.

3.6.7 Delay Timings The rock breakage profile in blasting is mainly dependent on delay timings. The delay timings work to separate the firing of blastholes, thereby increasing the utility of the explosive energy. The delay interval between the blastholes should be such that the burden from the previously fired holes has enough time to move out and provide adequate relief to the subsequent holes.

44

3 Rock–Explosive Interaction During Underground Blasting

3.7 Summary The overview of role of rock mass properties, explosive parameters and blast design parameters on blasting outputs has been discussed in this chapter. The summary of discussions made in this chapter is as follows: I.

The rock breakage during blasting takes place under the influence of shock wave and gaseous pressure. The selection of proper explosive based on the strata demand is necessary to maximize the explosive energy utilization. II. The blasting outputs have foremost influence of rock properties. The proper assessment of rock parameters and geological discontinuities is important to decide optimal explosive and design suitable blasting pattern. III. The explosive parameters viz. detonation velocity, detonation pressure and density also influence the blasting results. The blast design parameters are optimized based on the assessment of rock–explosive interactions.

References Arora S, Dey K (2010) Estimation of near-field peak particle velocity: a mathematical model. J Geol Min Res 2(4):68–73. http://www.academicjournals.org/jgmr Artero-Guerrero J, Pernas-Sánchez J, Teixeira-Dias F (2017) Blast wave dynamics: the influence of the shape of the explosive. J Hazard Mater 331:189–199. https://doi.org/10.1016/j.jhazmat. 2017.02.035 AminShokravi A, Eskandar H, Derakhsh AM, Rad HN, Ghanadi A (2018) The potential application of particle swarm optimization algorithm for forecasting the air-overpressure induced by mine blasting. Eng Comput. 34(2): 277–285. https://doi.org/10.1007/s00366-017-0539-5 Ash RL (1963) The mechanics of rock breakage, standards for blasting design. Pit Quarry 56(3):118– 122 Babu RR, Benipal G, Singh AK (2005) Constitutive modeling of concrete: an overview. Asian J Civ Eng 6(4):211–246 Babu RR, Benipal GS, Singh AK (2010) Constitutive model for bimodular elastic damage of concrete. Latin Am J Solids Struct 7(2):143–166. https://doi.org/10.1590/S1679-782520100 00200003 Baker EL, Murphy D, Stiel LI, Wrobel E (2012) Theory and calibration of JWL and JWLB thermodynamic equations of state. 60, 41–52.https://doi.org/10.2495/978-1-84564-750-6/05 Baranowski P, Małachowski J (2018) Possibilities of rock constitutive modelling and simulations, 130005. https://doi.org/10.1063/1.5019135 Baranowski P, Mazurkiewicz Ł, Małachowski J, Pytlik M (2020) Experimental testing and numerical simulations of blast-induced fracture of dolomite rock. Meccanica 55(12):2337–2352. https:// doi.org/10.1007/s11012-020-01223-0 Bauer A, Calder PN (1970) The influence and evaluation of blasting on stability in open pit mining. Society of Mining Engineers of the American Institute of Mining, Metallurgy and Petroleum Engineers Beser MH, Aydiner K (2019) Effect of stress level on the compressive strength of the rock samples subjected to cyclic loading. IOP Conf Ser: Earth Environ Sci 221:012023. https://doi.org/10. 1088/1755-1315/221/1/012023 Bhandari S (1997) Engineering rock blasting operations

References

45

Bhandari S, Vutukuri VS (1974) Rock fragmentation with longitudinal charges. In: 3rd International congress on rock mechanics, Denever, United States, pp 1337–1342 Brinkmann JR (1990) An experimental study of the effects of shock and gas penetration in blasting. In: 3rd International symposium on rock fragmentation by blasting, pp 55–66 Castedo R, Natale M, López LM, Sanchidrián JA, Santos AP, Navarro J, Segarra P (2018) Estimation of Jones-Wilkins-Lee parameters of emulsion explosives using cylinder tests and their numerical validation. Int J Rock Mech Min Sci 112:290–301. https://doi.org/10.1016/j.ijrmms.2018.10.027 Chiquito M, Castedo R, López LM, Santos AP, Mancilla JM, Yenes JI (2019) Blast wave characteristics and tnt equivalent of improvised explosive device at small-scaled distances. Defence Sci J 69(4):328–335. https://doi.org/10.14429/dsj.69.12637 Coggan J, Gao F, Stead D, Elmo D (2012) Numerical modelling of the effects of weak immediate roof lithology on coal mine roadway stability. Int J Coal Geol 90–91:100–109. https://doi.org/ 10.1016/j.coal.2011.11.003 Cui J, Hao H, Shi Y (2017) Discussion on the suitability of concrete constitutive models for highrate response predictions of RC structures. Int J Impact Eng 106:202–216. https://doi.org/10. 1016/j.ijimpeng.2017.04.003 Dally JW, Fourney WL, Holloway DC (1975) Application of dynamic photoelasticity to excavation technology. Recent Adv Eng Sci 7:19–28 Fleetwood KG, Villaescusa E, Li J (2009) Limitations of suing PPV damage models to predict rock mass damage. International Society of Explosive Engineers Forsyth WW (1993) A discussion on the blast induced overbreak around underground excavations.Proceedings of 4th Int. Symp on Rock Fragmentation by Blasting 161–166 Gao Q, Lu W, Yan P, Hu H, Yang Z, Chen M (2019) Effect of initiation location on distribution and utilization of explosion energy during rock blasting. Bull Eng Geol Env 78(5):3433–3447. https://doi.org/10.1007/s10064-018-1296-4 Gómez S, Sanchidrián JA, Segarra P (2020) Near-field vibration from blasting and rock damage prediction with a full-field solution. Int J Rock Mech Min Sci 134:104357. https://doi.org/10. 1016/j.ijrmms.2020.104357 Gong F-Q, Si X-F, Li X-B, Wang S-Y (2019a) Dynamic triaxial compression tests on sandstone at high strain rates and low confining pressures with split Hopkinson pressure bar. Int J Rock Mech Min Sci 113:211–219. https://doi.org/10.1016/j.ijrmms.2018.12.005 Gorai AK, Himanshu VK, Santi C (2021) Development of ANN-based universal predictor for prediction of blast-induced vibration indicators and its performance comparison with existing empirical models. Mining Metallurgy Explorat. https://doi.org/10.1007/s42461-021-00449-0 Gong F, Zhao G, Zhang Q, Wu W (2019b) Dynamic failure characteristics and behavior of rock materials. Adv Civ Eng 2019:1–2. https://doi.org/10.1155/2019/6972053 Gui YL, Zhao ZY, Zhou HY, Goh ATC, Jayasinghe LB (2017) Numerical simulation of rock blasting induced free field vibration. Procedia Eng 191:451–457. https://doi.org/10.1016/j.proeng.2017. 05.203 Hajihassani M, Jahed Armaghani D, Marto A, Tonnizam Mohamad E. (2015a) Vibrations au sol prédiction dans quarry dynamitage à travers un réseau neural artificiel optimisé par une concurrence impérialiste algorithme. Bull Eng Geo Environ 74(3):873–886. https://doi.org/10.1007/ s10064-014-0657-x Hajihassani M, Jahed Armaghani D, Monjezi M, Mohamad ET, Marto A (2015b) Blast-induced air and ground vibration prediction: a particle swarm optimization-based artificial neural network approach. Environ Earth Sci 74(4):2799–2817. https://doi.org/10.1007/s12665-015-4274-1 He C, Yang J (2019) Experimental and numerical investigations of dynamic failure process in rock under blast loading. Tunn Undergr Space Technol 83(3):552–564. https://doi.org/10.1016/j.tust. 2018.08.047 Himanshu VK, Mishra AK, Roy MP, Vishwakarma AK, Singh PK (2021) Numerical simulation based approach for assessment of blast induced deformation pattern in slot raise excavation. Int J Rock Mech Min Sci 144: 104816. https://doi.org/10.1016/j.ijrmms.2021.104816

46

3 Rock–Explosive Interaction During Underground Blasting

Himanshu VK, Mishra AK, Vishwakarma AK, Roy MP, Singh PK (2022) Explicit dynamics based numerical simulation approach for assessment of impact of relief hole on blast induced deformation pattern in an underground face blast. Geomech Geophy for Geo-Energy and Geo-Resour 8:19. https://doi.org/10.1007/s40948-021-00327-5 Hu Y, Lu W, Chen M, Yan P, Zhang Y (2015) Numerical simulation of the complete rock blasting response by SPH-DAM-FEM approach. Simul Model Pract Theory 56:55–68. https://doi.org/ 10.1016/j.simpat.2015.04.001 Hustrulid W, Lu W (2003) The Lu-Hustrulid approach for calculating the peak particle velocity caused by blasting. In: Explosives and blasting technique. Taylor and Francis, pp 291–300. https://doi.org/10.1201/9781439833476.ch36 International Society of Explosives Engineers (2011) ISEE blasters’ handbook. International Society of Explosives Engineers, Cleveland, Ohio Jaiswal A (2018) Ansys training material. https://www.scribd.com/document/388146000/ANSYSExplicit-Dynamics-Ashish-Jaiswal-pdf Jimeno Carlos L, Jimeno Emilio L, Ayala Carcedo Francisco J, Yvonne Visser de R (1987) Drilling and blasting of rocks, 1st ed. Routledge. https://doi.org/10.1201/9781315141435 Konya CJ, Walter EJ (1990) Surface blast design. Prentice Hall Khandelwal M, Singh TN (2009) Prediction of blast-induced ground vibration using artificial neural network. Int J Rock Mech Min Sci 46(7): 1214–1222. https://doi.org/10.1016/j.ijrmms.2009. 03.004 Kutter HK, Fairhurst C (1971) On the fracture process in blasting. Int J Rock Mech Min Sci Geomech Abstracts 8(3):181–202. https://doi.org/10.1016/0148-9062(71)90018-0 Kumar R, Choudhury D, Bhargava K (2014) Prediction of Blast-Induced Vibration Parameters for Soil Sites. Int J Geomech. 14(3):04014007. https://doi.org/10.1061/(asce)gm.1943-5622.000 0355 Kumar R, Choudhury D, Bhargava K (2016) Determination of blast-induced ground vibration equations for rocks using mechanical and geological properties. J Rock Mech Geotech Eng 8(3):341–349. https://doi.org/10.1016/j.jrmge.2015.10.009 Langefors U, Kihlstrom B (1963) The modern technique of rock blasting. Wiley Leng Z, Sun J, Lu W, Xie X, Jia Y, Zhou G, Chen M (2021) Mechanism of the in-hole detonation wave interactions in dual initiation with electronic detonators in bench blasting operation. Comput Geotech 129:103873. https://doi.org/10.1016/j.compgeo.2020.103873 Li XP, Huang JH, Luo Y, Dong Q, Li YH, Wan Y, Liu TT (2017) Numerical simulation of blast vibration and crack forming effect of rock-anchored beam excavation in deep underground caverns. Shock Vib. https://doi.org/10.1155/2017/1812080 Liu Y, Wei J, Ren T (2016) Analysis of the stress wave effect during rock breakage by pulsating jets. Rock Mech Rock Eng 49(2):503–514. https://doi.org/10.1007/s00603-015-0753-7 Ma GW, An XM (2008) Numerical simulation of blasting-induced rock fractures. Int J Rock Mech Min Sci 45(6):966–975. https://doi.org/10.1016/j.ijrmms.2007.12.002 Maji V, Adugna A (2016) Numerical modeling of tunneling induced ground deformation and its control. Int J Min Geo-Eng 50(2):183–188. https://doi.org/10.22059/ijmge.2016.59827 Mishra AK, Rout M, Singh DR, Jana SP (2018) Influence of Gassing Agent and Density on Detonation Velocity of Bulk Emulsion Explosives. Geotech Geol Eng 36(1):89–94. https://doi.org/ 10.1007/s10706-017-0308-7 Onederra I, Esen S (2004) An alternative approach to determine the Holmberg-Persson constants for modelling near field peak particle velocity attenuation. Fragblast 8(2):61–84. https://doi.org/ 10.1080/13855140412331336151 Pal A (2015) https://www.slideshare.net/abhijitpal00/research-on-mean-partical-size-after-drillingblasting-by-abhijit-pal Pal Roy P (2005) Rock blasting effects and operations. Oxford and IBH Publishing Company Pvt. Ltd. Pal Roy P (2007) Technical guidelines for controlled blasting

References

47

Pan B, Wang X, Xu Z, Guo L, Wang X (2021) Experimental and numerical study of fracture behavior of rock-like material specimens with single pre-set joint under dynamic loading. Materials 14(10):2690. https://doi.org/10.3390/ma14102690 Persson PA, Lundborg N, Johansson SH (1970) The basic mechanism of blasting. 2nd Cong ISRM, Belgrade 3(5):19–33 Porter DD, Fairhurst C (1970) A study of crack propagation produced by the sustained borehole pressure in Blastin. In: Dynamic rock mechanics, AIME, New York Pramanik R, Deb D (2015) Implementation of smoothed particle hydrodynamics for detonation of explosive with application to rock fragmentation. Rock Mech Rock Eng 48(4):1683–1698. https://doi.org/10.1007/s00603-014-0657-y Prasad U, Mohanty B, Nemes JA (2000) Dynamic fragmentation of selected rocks under impact loading. In: North American rock mechanics symposium, Seattle, Washington, pp 577–581 Renjie W, Haibo L, Xiaofeng L, Xiang X, Liwang L (2020) Experimental study and numerical simulation of the dynamic behavior of transversely isotropic phyllite. Int J Geomech 20(8):4020105. https://doi.org/10.1061/(ASCE)GM.1943-5622.0001737 Rezaeineshat A, Monjezi M, Mehrdanesh A, Khandelwal M (2020) Optimization of blasting design in open pit limestone mines with the aim of reducing ground vibration using robust techniques. Geomech Geophy for Geo-Energy and Geo-Resour 6(2). https://doi.org/10.1007/s40948-02000164-y Rustan RA (1992) Burden, spacing and borehole diameter at rock blasting. Int J Surf Min Reclam Environ 6(3):141–149. https://doi.org/10.1080/09208119208944329 Sanchidrián JA, López LM (2006) Calculation of the energy of explosives with a partial reaction model. Comparison with cylinder test data. Prop Explos Pyrotech 31(1):25–32. https://doi.org/ 10.1002/prep.200600003 Sanchidrián JA, Castedo R, López LM, Segarra P, Santos AP (2015) Determination of the JWL constants for ANFO and emulsion explosives from cylinder test data. Central Eur J Energ Mater 12(2):177–194 Sengani F (2020) Fundamental principles of rock fracturing at the vicinity of preconditioned blast hole. Arch Min Sci 65:769–786. https://doi.org/10.24425/ams.2020.134146 Shahrin MI, Abdullah RA, Jeon S, Jeon B, Sa’Ari R (2019) Numerical simulation of rock fragmentation by blasting using discrete element method and particle blast method. IOP Conf Ser: Mater Sci Eng 527(1). https://doi.org/10.1088/1757-899X/527/1/012032 Shockey DA, Curran DR, Seaman L, Rosenberg JT, Petersen CF (1974) Fragmentation of rock under dynamic loads. Int J Rock Mech Min Sci Geomech Abstracts 11(8):303–317. https://doi. org/10.1016/0148-9062(74)91760-4 Shukla N, Mishra MK (2020) Experimental evaluation of failure characteristics of coal using 2D digital image correlation approach. Arab J Geosci 13(20). https://doi.org/10.1007/s12517-02006044-9 Singh PK, Roy MP, Paswan RK, Sarim M, Kumar S, Jha RR (2016) Rock fragmentation control in opencast blasting. J Rock Mech Geotec Eng 8(2):225–237. https://doi.org/10.1016/j.jrmge. 2015.10.005 Silva J, Worsey T, Lusk B (2019) Practical assessment of rock damage due to blasting. Int J Min Sci Technol 29(3):379–385. https://doi.org/10.1016/j.ijmst.2018.11.003 Tan L, Ren T, Yang X, He X (2020) Numerical study on the fracture characteristics and failure mode of hard coal under coupled static and dynamic loads. E3S Web Conf 192. https://doi.org/ 10.1051/e3sconf/202019204002 Torbica S, Lapcevic V (2014) Model for estimating blasted rock fragmentation. Int Multidiscip Sci GeoConf Surv Geol Min Ecol Manag, SGEM 3(1):379–386. https://doi.org/10.5593/sgem2014/ b13/s3.050 Trivino LF, Mohanty B (2015) Assessment of crack initiation and propagation in rock from explosion-induced stress waves and gas expansion by cross-hole seismometry and FEM–DEM method. Int J Rock Mech Min Sci 77:287–299. https://doi.org/10.1016/j.ijrmms.2015.03.036

48

3 Rock–Explosive Interaction During Underground Blasting

Wang J, Yin Y, Luo C (2018) Johnson-Holmquist-II(JH-2) constitutive model for rock materials: parameter determination and application in tunnel smooth blasting. Appl Sci (Switzerland) 8(9). https://doi.org/10.3390/app8091675 Wen S, Zhang C, Chang Y, Hu P (2020) Dynamic compression characteristics of layered rock mass of significant strength changes in adjacent layers. J Rock Mech Geotech Eng 12(2):353–365. https://doi.org/10.1016/j.jrmge.2019.09.003 Wu S, Zhang S, Guo C, Xiong L (2017) A generalized nonlinear failure criterion for frictional materials. Acta Geotech 12(6):1353–1371. https://doi.org/10.1007/s11440-017-0532-6 Yang R, Bawden WF, Katsabanis PD (1996) A new constitutive model for blast damage. Int J Rock Mech Min Sci Geomech Abstracts 33(3):245–254. https://doi.org/10.1016/0148-9062(95)000 64-X Yunoshev AS, Plastinin AV, Sil’vestrov VV (2012) Effect of the density of an emulsion explosive on the reaction zone width. Combust Explos Shock Waves 48(3):319–327 Yang P, Lei Q, Xiang J, Latham JP, Pain C (2020) Numerical simulation of blasting in confined fractured rocks using an immersed-body fluid-solid interaction model. Tunnell Undergr Space Technol 98(July 2019):103352. https://doi.org/10.1016/j.tust.2020.103352 Yilmaz O, Unlu T (2014) An application of the modified Holmberg-Persson approach for tunnel blasting design. Tunn Undergr Space Technol 43:113–122. https://doi.org/10.1016/j.tust.2014. 04.009 Zhang ZX (2016) Rock fracture and blasting. Elsevier. https://doi.org/10.1016/C2014-0-01408-6 Zhang Z, Gao W, Li K, Li B (2020) Numerical simulation of rock mass blasting using particle flow code and particle expansion loading algorithm. Simul Model Pract Theory 104:102119. https:// doi.org/10.1016/j.simpat.2020.102119

Chapter 4

Drivage Excavation Using Drilling and Blasting

Abstract The drivage excavation in underground mine is done to make the access to the orebody. The excavation is dominantly performed using drilling and blasting. The burn-cut blasting pattern is used in all the Indian metalliferous mines to make drivages. The pattern consists of a set of holes to be blasted against the free face provided by relief/reamer holes. The design parameters in burn-cut face blasting include drilling pattern (burden, spacing, blasthole diameter, relief hole diameter, number of blastholes, number of relief holes, number of dummy holes, etc.), charging pattern and firing pattern. These design parameters need to be designed based on the assessment of rock mass properties at the site. The numerical simulation and nearfield ground vibration waveform analysis may be advantageous in the designing of blasting pattern of drivage blasting to obtain optimum breakage. The delay pattern and maximum charge weight pre-delay for cut and perimeter of the blasting face can be optimized using these tools. With the technological advancements, the bulk emulsion explosives with varying densities have evolved in the recent days. Such explosives provide flexibility to the blast designers to optimize the charging pattern to get optimum damage.

4.1 Introduction The major challenge with the blasting for the excavation of tunnels or drivages of an underground metal mine is to achieve the optimum pull with the reduction in over break. The blasting for a drivage excavation also comes with the operational challenges due to the presence of limited free face in single direction only. The principle and methodology for the drivage excavation and tunnel excavation are similar. The tunnel excavation is carried out using the wedge-cut or burn-cut pattern. In wedge-cut pattern, the horizontal holes are drilled at an angle of about 60°, such that they came closer at one end to form a wedge (Mei et al. 2021; Ramulu 2012; Rana et al. 2022; Venkatesh et al. 2018). All the drivages/face blasting operation in Indian underground metalliferous mines are carried out using burn-cut pattern. This pattern has a set of cut holes containing empty (relief/reamer) and blastholes. The relief holes work to provide additional free face to ensure tensile breakage in this pattern. © The Author(s), under exclusive license to Springer Nature Singapore Pte Ltd. 2023 V. K. Himanshu et al., Blasting Technology for Underground Hard Rock Mining, https://doi.org/10.1007/978-981-99-2645-9_4

49

50

4 Drivage Excavation Using Drilling and Blasting

This is required to maximize the explosive energy utilization and thereby obtaining the optimum pull from the face blast (Adhikari 1994; Gupta et al. 1991; Murthy and Dey 2002; Verma et al. 2018). The diameter of relief holes is generally kept more than that of the blastholes. Allen (2014) emphasized in the research findings that the lengthen relief holes will result into the improvement of pull-in burn-cut pattern. Singh (1995) further suggested that the dimension of reamer holes, including their diameter, number, depth and relief area, plays a significant role in achieving the optimum damage due to blasting. Further, the optimum dimensions of relief hole diameter would be different under varying physico-mechanical properties of rock (Liu and Qiu 2020; Salum and Murthy 2019; Murthy and Dey 2002). Sharma (2005) studied the impact of multiple reamer holes on the results of a face blast. It was concluded in this study that the multiple number of relief holes will be more effective than the single large diameter hole to prevent freezing in the spongy rock mass. The rock mass properties, viz. uniaxial dynamic compressive strength, uniaxial dynamic tensile strength, elastic modulus, Poisson’s ratio, etc., have foremost impact on the induced damage profile from a face blast. The prior assessment of these parameters helps the blast designers in estimating the optimum blast design pattern. The P-wave and S-wave velocities of the rock strata are another important parameter. The optimum detonation velocity of the explosive for blasting in a rock is decided by the prior assessment of the ultrasonic velocities (Verma et al. 2018). The results of drivage blasting for underground metalliferous mining are also influenced by the insitu stresses. Researchers found in their study that the pre-stressed strata shows higher overbreak under the similar conditions of the explosive loading (Abdel-Meguid et al. 2003; Mandal and Singh, 2009; Xiao et al. 2019). Mandal and Singh (2009) suggested to do excavations in small sections and phase-wise to reduce the overbreak under such conditions. Verma et al. (2016) highlighted the role of the discontinuities in the rock strata in the blasting results from a tunnel blast. Authors correlated the parameters such as Q system, rock mass rating (RMR), rock quality designation (RQD) and core recovery (CR) with the resulting overbreak from the blast. The optimum explosive parameters are selected for a face blast based on the results of the assessment of the rock mass. The suitable explosive quantity and quality are selected for maximizing the pull and minimizing the overbreak. Researchers have outlined the role of charging parameters and initiation sequence on the blast-induced damage pattern. Mandal et al. (2005) found that the magnitude of overbreak enhances with the decrement in the delay interval. Bullock (2013) highlighted the role of the velocity of detonation (VOD) of the explosive on induced damage pattern from blasting. Authors suggested to use the low VOD explosives for the blast in the spongy (plastic) nature of rock compared to the brittle nature of the rock. The excess scattering in the delay detonators significantly influences the profile of breakage in the face blast. The scattering leads to the improper detonation of the charged blast holes and thereby influences the creation of the required cut for the progressive initiation round. Vishwakarma et al. (2020) found that the scattering in the delay detonators impacts the resulting pull as well as enhances the overbreak. The profile of excavation in the face blast is affected by several other parameters such as the

4.2 Elements of Burn-Cut Blast Design

51

magnitude of deviation in the drill holes and hole spacing (Cardu and Seccatore 2016; Singh 2018). The detailed discussion regarding the burn-cut face blasting pattern, measurement of face blasting outputs and optimization of face blasting pattern has been made in different sections of this chapter.

4.2 Elements of Burn-Cut Blast Design The burn-cut face blasting design consists of a set of blastholes and reamer holes. The reamer holes are mostly of larger diameter than that of the blastholes and are kept uncharged during the process of blasting. The blastholes of the burn-cut pattern are classified as—cut holes, easers, lifters and perimeter holes. The blastholes surrounding the reamer holes are termed as cut holes. These cut holes are fired against the free face provided by reamer holes. The judicious drill design and firing pattern for cut holes are important to achieve the maximum pull from the face blast. Once the firing of cut holes is completed, the easers are fired along the free face created by the cut portion. The lifters are placed in the bottom of the face. They are fired either in the last or before the firing of the perimeter holes. The perimeter holes play important role in the reduction of overbreak during face blasting. Accordingly, the longer delay interval is provided in the perimeter holes. The longer delay interval works to provide sufficient relief to the blastholes and thereby ensure tensile breakage of the rock mass. Sometimes line drilling is also done along the perimeter holes to reduce the overbreak. These line drilling holes are kept uncharged, and work for reducing the superposition of stress waves from different blastholes. A sample face blasting design with the positions of reamer, cut, easer, perimeter and line drilling holes is shown in Fig. 4.1. The position of reamer and cut holes in a drivage blasting face of an underground mine are shown in Fig. 4.2. The arrangement of blastholes is changed as per the site requirement. The rules of thumb for burn-cut face blasting pattern have been developed over the years by blasting practitioners. These rule of thumbs states that the burden/spacing for easer holes should be in the range of 16–22 times of the blasthole diameter. It should be in the range of 13–17 times the blasthole diameter for perimeter holes. The reamer holes should be chosen such that it provides void of about 15% of the overall drivage face (Dyno Nobel Underground Manual 2004). However, the blasting practitioners judiciously change these parameters to achieve the optimum output from the face blast.

52

4 Drivage Excavation Using Drilling and Blasting

Fig. 4.1 View of arrangements of blastholes and reamer holes in burn-cut face blasting pattern

Fig. 4.2 View of arrangement of cut holes and reamer holes in an underground drivage

4.3 Measurement of Outputs from Underground Drivage Blasting The outputs from a drivages blasting are measured in terms of fragmentation, pull achieved and overbreak. The measured output is analysed, and parametric variations in blast design are done to improve the output. The fragmentation analysis from the

4.3 Measurement of Outputs from Underground Drivage Blasting

53

face blast is carried out using image analysis. The images of blasted muckpile are taken for this purpose. The size distribution of the fragmented rock is analysed. A view of fragmentation output from an underground drivage face blast is shown in Fig. 4.3. The optimum fragmentation output helps the mine management in reducing the cost of downstream processes, viz. loading and breaking. The fragmentation is not an issue in most of the cases of face blasting. This is because the charge concentration in a small area of face blasting is very high. The pull and overbreak measurements from the face blasting are done using survey equipment. The conventional way to assess the pull and overbreak is by using total station-based survey. During recent times, the scanning systems have evolved, which are capable of measuring these parameters more accurately and efficiently. The comparison of drivage profile before and after blasting can be done using these scanners. A view of drivage profiling after blast using cavity monitoring system (CMS) scanner for an underground mine is shown in Fig. 4.4. The reduced pull from the blast also pose safety threats. Sometimes, when the pull from a blast face is below 50%, then there is formation of socket in the post-blast faces. There may be a possibility of the presence of unburnt explosive cartridge or detonator inside such sockets. These unburnt cartridges may detonate while hole cleaning operation performed by a miner. So, the adequate inspection of sockets

Fig. 4.3 Rock fragmentation output from a drivage face blast

(a) Development face blast

(b) Haulage Drive

Fig. 4.4 View of drivage profiling using cavity monitoring system

54

4 Drivage Excavation Using Drilling and Blasting

Fig. 4.5 View of assessment of socket in a blasted face

after the blasts is required. The sockets should be inspected physically. A view of the inspection of socket in a blasted drivage face is shown in Fig. 4.5.

4.4 Optimization of Drivage Blasting Pattern The optimum output from a face blasting pattern can be achieved by assessing the rock–explosive interaction behaviour and thereby designing the blasting pattern. The rock mass damage in this operation is optimized to enhance pull and reduce overbreak. For the purpose of pull enhancement, the requirement of detonation pressure/borehole pressure in cut portion of the face blast is more. To enhance the borehole pressure in this region, the explosive charge with high density, high velocity of detonation (VOD) and better coupling are used. Sometimes two or more blastholes are detonated simultaneously to increase the resultant tensile breakage in the cut portion. However, the care is required while deciding the number of blastholes firing simultaneously in the cut portion. As the firing of larger number of holes simultaneously may lead to over-damage in the cut portion, and thereby the explosives of the easer holes would be unutilized, which may result into the reduction of pull and enhancement in overbreak. The number and diameter of reamer holes also affect the resultant pull from the cut portion of the drivage blasting. Himanshu et al. (2022) found in the study that the extent of damage in the cut portion of the face blasting increases with the increment in number and diameter of reamer holes. Authors suggested that the blast face shows more controlled deformation while using multiple number of relief holes of optimum diameter as compared to a single large diameter relief hole. The comparison of extent of deformation due to cut hole blasting under relief provided by varying diameter and numbers of reamer holes is shown in Fig. 4.6. Jimeno et al. (1987) also emphasized that the single large diameter relief hole provides more relief

Fig. 4.6 Comparison of extent of deformation in rock mass for burn-cut development face blast under different variations of relief holes (Himanshu et al. 2022)

Extent of damage in perephery (m2)

4.4 Optimization of Drivage Blasting Pattern

55

25 20 15 10

1 relief hole 2 relief holes

5

3 relief holes 4 relief holes

0

0

50

100

150

200

Diameter of relief holes (mm)

than the multiple numbers of relief holes. Authors proposed a relationship for equivalent diameter of a single relief hole compared to the multiple number of relief holes. The relationship is shown in Eq. 4.1 eq

φR =

√

n × φR ,

(4.1)

where eq

φR = equivalent diameter of relief hole n = number of relief holes φR = diameter of relief holes. The overbreak control in drivage face blasting pattern is mainly dependent on the damage induced by perimeter holes. Therefore, designing of charging pattern and delay sequence of perimeter holes are important to reduce overbreak from the face blasting. The insufficient breakage of rock mass in cut or easer portion of the burn-cut pattern may also result into the overbreak. This is mainly because of unavailability of sufficient free face for the perimeter holes. To reduce the overbreak, the decoupled explosive charge is used in the perimeter holes of the burn-cut pattern. The firing sequence is also designed to reduce the maximum charge weight per delay, and thereby resulting vibration in the periphery of the face. The vibration is required to be reduced below the critical vibration limit to control damage in order to reduce overbreak. The low-density explosives are also used in perimeter holes to reduce the resultant damage beyond the line of breakage. With the advancements in explosive products, the bulk emulsion explosives are being charged in drivage blasting faces of different underground mine. Such bulk emulsion explosives come with different density and velocity. The density and VOD of bulk emulsion explosive used at Indian Underground Lead–Zinc mine are given in Table 4.1 (Saw and Dasgupta 2022).

56

4 Drivage Excavation Using Drilling and Blasting

Table 4.1 Properties of bulk emulsion explosive used at Indian Lead–Zinc underground mine (Saw and Dasgupta 2022) Properties Density (gm/cc)

0.8

0.9

1.0

1.1

1.2

Critical diameter (mm)

38

38

38

38

42

Velocity of detonation (m/s)

4500

4900

5300

5700

6200

Relative weight strength (RWS)

72

78

85

92

98

Relative bulk strength (RBS)

72

89

106

127

147

Sleep time

7 days

The variability in the density and VOD gives the flexibility to the blast designers to better plan the blast. The detonation pressure exerted by the explosive is a function of density and VOD. The relationship among the detonation pressure, density and VOD of explosive is given in Eq. 4.2.   ρe × VOD2 × 106 , Pd = 4

(4.2)

where Pd = Detonation pressure (MPa) ρe = Density of explosive (kg/m3 ) VOD = Velocity of detonation (m/s). The detonation pressure gets transferred as borehole pressure on the blasthole wall during the process of blasting. The transfer amount is dependent on the rock type, explosive coupling and confinement. The explosive charge with high detonation pressure is charged in the cut portion, and with low detonation pressure is charged in the perimeter holes of the burn-cut pattern. This ensures the generation of low borehole pressure in the periphery region and thereby reduces the overbreak.

4.4.1 Numerical Simulation and Prediction of Damages from Drivage Blasting The blast design pattern for optimum damage from a drivage blasting would be different under varying physico-mechanical properties of rock strata. Numerical simulation may be a comprehensive tool to optimize the blast design pattern. In the simulation study, the numerical model with input rock mass parameters and explosive parameters is prepared. The blast design pattern is varied in different models, and output damages are evaluated. The design pattern is optimized to get enhanced pull and reduced overbreak under the prevailing rock mass condition.

4.4 Optimization of Drivage Blasting Pattern

57

Himanshu et al. (2022) optimized the cut hole blasting pattern for drivage blasting in an Indian Lead–Zinc underground mine using numerical simulation. The drivages at this mine are usually of 4.0 m × 5.0 m dimension. The blastholes are drilled in the burn-cut pattern to make the drivages. The burn-cut pattern at the mine incorporates drilling of the blastholes of diameter 40 mm and drilling length of 4.0 m. The existing burn-cut blasting pattern at the site consists of 56 charged blastholes of 40 mm diameter with 04 relief holes of 89 mm diameter. The existing drilling and blasting pattern practised at this mine site are shown in Fig. 4.7. To optimize the blast design pattern of this site, the numerical simulation with varying diameter and numbers of relief holes was carried out. The extent of damage was assessed under different parametric variations. A view of the extent of damage due to blasting of cut holes with four relief holes of different diameter is shown in Fig. 4.8 (Himanshu et al. 2022). There were issues of socket formation as well as overbreak from the face blast at the site using the existing blast design pattern. The identification of sockets was made to investigate the cause of its formation. Most of the sockets were in the holes near the cut portion, which reveals that there was overbreak while blasting the cut portion. This overbreak might have caused to restrict the detonation of the nearby charged blastholes, which have resulted in the socket formation. The overbreak may be due to the unavailability of proper void for relief. This might have led to excessive

Fig. 4.7 Existing drilling and blasting patterns practised at the experimental development faces of the study site (Himanshu et al. 2022)

58

4 Drivage Excavation Using Drilling and Blasting

Fig. 4.8 Damage due to firing of cut holes against four relief holes of different diameter (Himanshu et al. 2022)

damage around the cut portion resulted into improper blast. The extent of damage output from the numerical model has been compared to estimate the optimum cut blasting pattern to reduce the damage due to the blast of cut holes. The extent is due to the simultaneous detonation of thirteen blastholes. The numerical simulation with the practical delay timing is not possible with the Ansys-Explicit Dynamics module. Hence, simultaneous detonation was provided in the model to all the cut blastholes. Based on the dependency of rock breakage on critical peak particle velocity (Holmberg and Persson 1978), the blast-induced damage can be considered proportional to the maximum charge per delay (MCPD). Accordingly, the damage while firing of two blastholes of cut simultaneously will be 1/6.5th of that of the simulation results. The extent of damage using this computation for the firing of two cut holes simultaneously against four relief holes of 115 mm diameter will be 2.4 square metres. Accordingly, the deformation will be 0.1 m more than the cut boundary extent using this pattern. Hence, this pattern can be considered optimum for the blasting face of the study site. Based on this result, the cut hole pattern of the blast design for the site was revised. The revised pattern consisted of four relief holes of 115 mm diameter. The number of cut holes firing simultaneously was also reduced to two in the revised pattern (Himanshu et al. 2022). The revised blast design pattern for the study site is given in Fig. 4.9.

4.4.2 Designing of Delay Pattern for Drivage Blasting Using Nearfield Ground Vibration Monitoring Results The nearfield ground vibration waveform is also helpful in the designing of optimum blasting pattern for face blasting. The ground vibration for such case needs to be monitored as near from the blasting face as possible. The ground vibration waveform at distant locations is due to the superposition of the waves from different delays. There is possibility of superposition due to reflection and refraction of waves due to change in media at far-field distances as well. So, the attempt needs to be made to record the vibration upto the distance where heterogeneity is not at a larger scale. To compare the nearfield ground vibration waveform, the geophones must be placed in the similar fashion in each case. The optimization of delay pattern for the

4.4 Optimization of Drivage Blasting Pattern

59

Fig. 4.9 Revised blast design based on the results of the numerical simulation and waveform analysis (Himanshu et al. 2022)

blast design shown in Fig. 4.7 was carried out by monitoring the nearfield ground vibration. For this purpose, the geophones were placed on the same level of the mine at which blast was conducted. It was placed using Plaster of Paris (PoP). A view of vibration monitoring at different locations in the underground is shown in Fig. 4.10. The waveform analysis of recorded nearfield vibration data was carried out to explore the possibility of design modifications for the reduction of overbreak. The recorded waveform for an experimental development face blast is shown in Fig. 4.11. The analysis of the recorded waveform shows two sharp peaks of vibration. One peak is due to the blast of cut holes, and another is due to the blast of perimeter holes. The vibration peaks have been compared with the face blasting pattern shown in Fig. 4.7. The peak due to the blast of cut holes is because of the firing of four cut holes simultaneously against the insufficient free face generated by relief holes of 89 mm diameter. The peak due to the firing of perimeter holes can be considered the main reason behind the overbreak due to the blast. The review of the existing blast design reveals that the number of holes blasted at delay no. 21 and 22 is 16 and 15, respectively, which increased the charge weight per delay, thereby increasing vibration in the nearby line of extraction in the development face. The increased level of vibration will result in enhanced over break. So, the design needed to be modified to distribute the delay sequence such that the charge weight per delay along periphery holes should be reduced. Accordingly, the maximum of 8 blastholes was suggested to be fired in the blast design, thereby reducing the MCPD by half in the revised blast design pattern shown in Fig. 4.9. This waveform was recorded at a distance of 30 m from the blast face (Himanshu et al. 2022).

60

4 Drivage Excavation Using Drilling and Blasting

Fig. 4.10 A view of nearfield blast vibration monitoring at different underground locations

Fig. 4.11 Recorded nearfield vibration waveform from a drivage blasting face using the blast design pattern shown in Fig. 4.7 (Himanshu et al. 2022)

4.5 Summary

61

Fig. 4.12 Recorded waveform for the drivages blasting using blast design pattern shown in Fig. 4.9 (Himanshu et al. 2022)

The waveform at a distance of 30 m from the blast face was also recorded after the blast using the revised blast design pattern shown in Fig. 4.9. The resulting waveform is shown in Fig. 4.12. The waveform analysis shows that the vibration magnitude has come down to 10 mm/s compared to the 40 mm/s vibrations shown in Fig. 4.11. The waveform analysis also reveals that the variation in magnitude of vibration while blasting of different cuts is relatively uniform. This leads to the controlled deformation of the rock mass. Hence, the redesigned blasting pattern is optimal for the study site (Himanshu et al. 2022).

4.5 Summary The techniques of underground drivage blasting along with their blast design optimization have been discussed in this chapter. Following are the summary of the discussions made in this chapter: I.

Drivage excavation in Indian condition is done using burn-cut drilling and blasting pattern. II. The cut hole pattern in burn-cut influences the induced pull significantly. Hence, the optimization of cut portion including the dimensions of relief and blastholes is necessary. Such optimization may be carried out using numerical simulation and waveform analysis of nearfield ground vibrations. III. The optimization of delay pattern for periphery holes is necessary to reduce blast-induced overbreak. The waveform analysis of nearfield ground vibration may be helpful in optimizing delay patterns for the periphery holes.

62

4 Drivage Excavation Using Drilling and Blasting

References Abdel-Meguid M, Rowe RK, Lo KY (2003) Three-dimensional analysis of unlined tunnels in rock subjected to high horizontal stress. Can Geotech J 40(6):1208–1224. https://doi.org/10.1139/ t03-057 Adhikari GR (1994) Controlled blasting in tunnel—some issues. Tras Inst Engineers (I), MN 75:56 Allen MR (2014) An analysis of burn cut pull optimization through varying relief hole depths. Thesis submitted to Missouri University of Science and Technology Bullock RL (2013) Tunneling and underground construction Cardu M, Seccatore J (2016) Quantifying the difficulty of tunnelling by drilling and blasting. Tunn Undergr Space Technol 60:178–182. https://doi.org/10.1016/j.tust.2016.08.010 Dyno Nobel Underground Manual (2004) Gupta RN, Singh RB, Adhikari GR (1991) Controlled blasting for underground excavations. Int J Rock Mech Min Sci Geomech Abstracts 28(2–3):A101. https://doi.org/10.1016/0148-906 2(91)92477-g Himanshu VK, Mishra AK, Vishwakarma AK, Roy MP, Singh PK (2022) Explicit dynamics based numerical simulation approach for assessment of impact of relief hole on blast induced deformation pattern in an underground face blast. Geomech Geophys Geo-Energy Geo-Resour 8(1):19. https://doi.org/10.1007/s40948-021-00327-5 Holmberg R, Persson P 1978. The Swedish approach to contour blasting. In: Proceedings of 4th annual conference on explosives and blasting research. International Society of Explosives Engineers, pp 113–127 Jimeno Carlos L, Jimeno Emilio L, Ayala Carcedo Francisco J, Yvonne Visser de R (1987) Drilling and blasting of rocks, 1st ed. Routledge. https://doi.org/10.1201/9781315141435 Liu K, Qiu J (2020) Investigation of burn cut parameters and model for one-step raise excavation based on damage evolution mechanisms. Geofluids 2020:1–16. https://doi.org/10.1155/2020/ 8879477 Mandal SK, Singh MM (2009) Evaluating extent and causes of overbreak in tunnels. Tunnell Undergro Space Technol 24(1):22–36. https://doi.org/10.1016/j.tust.2008.01.007 Mandal SK, Singh MM, Bhagat NK (2005) Causes of overbreak and influence of blast parameters for smooth undamaged wall. In: International symposium on advances in mining technology and management, IIT, Kharagpur, pp 49–58 Mei J, Zhang W, Xu B, Zhu Y, Wang B (2021) Experimental study of blasting excavation for large cross-section tunnel in horizontal layered rock mass. Preprint: Research Square. https://doi.org/ 10.21203/rs.3.rs-180296/v1 Murthy VMSR, Dey K (2002) Development of predictive models for blast-induced rock damage assessment (BIRD) in tunnels. MHRD Project No: MHRD (27)/99-00/111/ME:12–43 Ramulu M (2012) Blast optimisation with In Situ rock mass characterization by seismic profiling at an opencast coal mine in India. February, pp 387–396 Rana A, Bhagat NK, Singh A, Singh PK (2022) Predicting blast-induced pull using regression tree. Arab J Geosci 15(2):173. https://doi.org/10.1007/s12517-022-09452-1 Salum AH, Murthy VMSR (2019) Optimising blast pulls and controlling blast-induced excavation damage zone in tunnelling through varied rock classes. Tunn Undergr Space Technol 85:307– 318. https://doi.org/10.1016/j.tust.2018.11.029 Saw PK, Dasgupta S (2022) Selection of explosives in fully mechanised underground metaliferous mines. Indian Min Eng J 61(02):34–37 Sharma PD (2005) Tunnel blasting—emulsion explosives and proper blast design are the prerequisite for better efficiency. Indian J Mines Met Fuel 10 Singh SP (1995) Mechanism of cut blasting. Trans Inst Min Metall 104(A1):134–138 Singh S (2018) Overbreak control in underground mines. Rock Mech Open Access J 1(1):11–17 Venkatesh HS, Balachander R, Gopinath G (2018) Drilling and blasting of tunnels in Himalayan geology. Tunnel Asso India (TAI) J 7(01):21–28

References

63

Verma HK, Samadhiya NK, Singh M, Prasad VVR, Goel RK (2016) Investigations of rock mass damage induced by blasting. J Rock Mech Tunnell Technol (JRMTT) 22(1):2016–2065. www. isrmtt.com Verma HK, Samadhiya NK, Singh M, Goel RK, Singh PK (2018) Blast induced rock mass damage around tunnels. Tunn Undergr Space Technol 71:149–158. https://doi.org/10.1016/j.tust.2017. 08.019 Vishwakarma AK, Himanshu VK, Kumar S, Roy MP (2020) Overbreak control in development face blasting of underground metal mine—a case study. In: Proceedings of national conference on advances in mining (AIM-2020), pp 473–482 Xiao S-Y, Su L-J, Jiang Y-J, Liu Z-X (2019) Numerical analysis of hard rock blasting unloading effects in high in situ stress fields. Bull Eng Geol Env 78(2):867–875. https://doi.org/10.1007/ s10064-017-1067-7

Chapter 5

Box-Cut Excavation and Stope Opening

Abstract The initial excavation of orebody in longhole stoping is done by driving slot raise. Blasting for excavation of slot raise is different from the bench/ring blasting due to the absence of sufficient free face for the movement of the blasted rock mass. To address this issue, slot raises are excavated step-wise using drop raise method. However, the drop raise method is time intensive and hence not feasible in order to enhance the pace of production. So, there is utmost need for a technique to excavate complete slot raise in a single shot. The excavation of such slot raises comes with the challenges of face jamming, boulder generation and restricted face movements. The additional free face using uncharged reamer holes is provided to address these issues during the slot raise excavation. The blast design encompassing the drilling pattern, delay design, etc., for slot raise excavation is done based on the assessments of the rock–explosive interactions. Numerical simulation is a comprehensive tool for such assessments. Various techniques of slot raise excavation along with the methods for designing a slot raise blasting pattern have been discussed in this chapter. The advanced guided slot raise drilling and blasting technique has also been explained and compared with the conventional techniques.

5.1 Introduction The initial excavation in the ore body is done to open the stope. Since such excavation is to be carried out without free face, the excavation may be termed as box-cut. In cut and fill method of stoping, the box-cut excavation is done by driving sill drivages. These sill drivages are further extended in the operation of stoping. In longhole stoping method, the box-cut excavation is done by driving raises. These raises work as the free face for the accumulation of rock in the subsequent excavation stages. There are other purposes for driving the raises such as creation of ventilation raise, formation of raises for material/ore transfer and cable layout. The raises used as the box-cut opening for the subsequent round of blasting are termed as slot raises (Tatiya 2005). Such slot raises are located either on the side or in the centre of a stope (Villaescusa 2014).

© The Author(s), under exclusive license to Springer Nature Singapore Pte Ltd. 2023 V. K. Himanshu et al., Blasting Technology for Underground Hard Rock Mining, https://doi.org/10.1007/978-981-99-2645-9_5

65

66

5 Box-Cut Excavation and Stope Opening

The raise development is done using the mechanized method or by drilling and blasting. The mechanized raise development is carried out using Raise Borer Machines (Vishwakarma et al. 2022). Since slot raise development needs to be done regularly for stope opening and that too at different places of the mine, so mechanized slot raise development is a difficult task. The mechanized slot raise development will require the frequent shifting of the machinery, which is a time- and cost-consuming task. The mechanized slot raise development method is also not feasible to cope up with the faster production demands. Hence, the drilling and blasting is the best-suited alternative for the slot raise development. The drilling- and blasting-based slot raise development methods have added advantage of lower capital cost investments. The blasting for slot raise excavation is different from the bench/ring blasting due to the unavailability of a free face. The free face during the slot raise excavation is available only along the excavated drivage. To address this limitation of the limited free face, the slot raises are excavated in multiple steps. The technique to excavate slot raise in multiple steps is termed as drop raising or inverse drop raising (Baase et al. 1982; Ferreira 2003; Liu and Tran 2000; Liu et al. 2019). However, this technique gives limited progress. So, there is an utmost need for the complete slot raise excavation in single shot. This is required to enhance the production pace. The challenges for the deep slot excavation are more severe. The availability of the limited free faces while deep slot hole excavation restricts the movements of the blasted rock and thereby leads to the issues of face jamming, freezing and production of oversize boulders. Generally, the additional free faces under such blasting conditions are provided by drilling relief/reamer holes (Singh 1995; Sharma 2005; Allen 2014). These reamer holes work as the free face for the cut holes of the slot raise. After the excavation of cut holes of the raise, subsequent rows are allowed to be blasted against the cavity made by cut holes. This cavity works as the swelling space for the blasted rocks. There is very limited literature available on the slot raise blasting pattern. The slot raise blasting concepts have been explained in some of the research papers, viz. Chandrakar et al. (2023), Dobran (1996), Gertsch and Bullock (1998), Mann (1998), Pakalnis and Hughes (2011), etc. These research papers signify the similarity in the concept of the slot raise blasting pattern with the burn-cut face blasting pattern. The only difference between the burn-cut face blast and slot raise blast is the direction of material movement (Liu et al. 2018). Accordingly, the idea of burn-cut face blast can be used to overcome the challenges of jamming, freezing or boulder generation from the slot raise blasting. The scientific blast design can be proposed to address these issues. Such scientific blast design will include the variations in the number and position of the reamers and blastholes, their firing sequence and the delay timings. The formulation of the scientific design will include the assessments of rock–explosive interactions and thereby designing of the drilling, charging and delay patterns. The details regarding the slot raise blasting techniques and the methodology for the slot raise blast design have been discussed in different sections of this chapter.

5.3 Drilling Pattern for Slot Raise Blasting

67

5.2 Principle of Underground Slot Raise Blasting The slot raise excavation is carried out against the limited space provided by drivages. The free face direction is also available only along the drivages. The additional free face for this purpose is made using relief/reamer holes. The diameter of relief holes is generally kept larger than that of the blastholes. The reamer holes are kept uncharged during the process of blasting. While firing of blastholes, the relief holes works to provide tension to the stress wave induced by blasting. Accordingly, the tensile breakage of rock mass takes place against the relief holes. The slot raises may be driven in a single go or in multiple rounds. In multi-round excavation process, the bottom part of the slot raise is blasted initially. The subsequent round of blasting takes place after plugging the hole from through part. If the slot raises have to be blasted in one go, then it is mandatory to ensure that the proper space is available for the excavation of muckpile generated from slot blasting. In most of the cases, the space for muckpile accumulation is provided by excavated drivages. If the slot raises of greater depth have to be excavated, then the additional space for muckpile accumulation is created. At Indian Lead–Zinc underground mine sites, the additional space is made by uphole drilling and blasting. The blastholes for slot raise excavation are drilled vertically. The holes are through from one level to the another level. The plugging of the blastholes is needed before explosive charging in such cases. The plugging works to hold the explosives and thereby prevent the possible dislodging of the explosive materials inside the open drivages. The plugging is done using woods, plastic bottles or stem plugs. The view of wooden plugging arrangement for slot raise excavation at an Indian underground mine is shown in Fig. 5.1.

5.3 Drilling Pattern for Slot Raise Blasting The drilling pattern for slot raise excavation varies with the dimension of slot to be excavated. The diameter of blastholes is 75–115 mm. The larger diameter blastholes provide ease in deeper hole drilling. It also reduces the drill deviation. However, the larger diameter blastholes increase the charge concentration in smaller volume of the excavation area, which may result into improper breakage. Accordingly, the selection of optimum diameter suitable to the respective rock type is important. The spacing between holes is generally kept 0.4–0.8 m. The diameter of relief/reamer holes is in the range of 115–165 mm. The number of blastholes and reamer holes for slot raise excavation needs to be optimized based on the strata condition. Himanshu et al. (2021) carried out study to investigate the deformation pattern in underground slot raise blasting. Authors suggested that the ratio of relief holes to blastholes in cut portion of the slot raise blasting pattern should be more than 1, to get the optimum breakage. This ratio gives more space to the cut holes for tensile breakage. The perimeter blastholes in slot raise blasting pattern are kept in square or scattered

68

5 Box-Cut Excavation and Stope Opening

Fig. 5.1 View of plugging arrangement using wooden plugs at an Indian underground mine

pattern. However, the scattered pattern of perimeter blastholes gives more controlled deformation than the square pattern. A sample slot raise blast design pattern is given in Fig. 5.2.

5.4 Charging and Initiation of Blastholes in Slot Raise Blasting Charging of blastholes for slot raise excavation may be done using slurry/emulsion explosive cartridges or site mixed emulsion (SME) explosives. The density of explosive for cut holes and perimeter holes may be varied while explosive charging using SME. The cut holes are charged with the high-density SME, whereas low-density explosives are used in perimeter holes to control damages. The initiation of blastholes is done either using long delay non-electric (NONEL) delay detonators or using electronic delay detonators. Electronic delay detonators are preferred for slot raise blasting, as it gives flexibility to the blast designers to give timings to the blastholes. The position of initiation in slot raise blasting is also important. If the delay detonators are put in the extreme bottom of the blasthole, then the confinement to the explosive energy would be less. In this case, the explosive energy utilization will reduce. So, the judicious placement of initiation point is necessary to

5.5 Delay Sequence of Blastholes for Slot Blasting Pattern

69

Fig. 5.2 Sample slot raise blast design pattern (Himanshu et al. 2021)

have optimum utilization of explosive energy. Himanshu et al. (2021) conducted the study using numerical simulation to optimize the position of initiation system during slot raise excavation. A comparison of the model output from this study, showing the deformation pattern while taking initiation point in the middle of the blasthole and at one-fourth distance from the plug in the blasthole is given in Fig. 5.3. The analysis of deformation pattern reveals that the deformation is more uniform when the detonation point has been placed at one-fourth distance of blasthole from the plug. This is due to the tensile nature of breakage of the rock strata under blast loading. The time taken by blast wave to reach near free face (plug portion) is much higher when the initiation was put in the middle of the blasthole. Hence, the magnitude of tensile breakage is lower in such case. Accordingly, it may be concluded that the detonation point at a distance of about 20–30% of the blasthole length from plug will give the maximum breakage under blast loading.

5.5 Delay Sequence of Blastholes for Slot Blasting Pattern Long delay timings between the blastholes are kept in slot blasting pattern. This is to ensure the creation of free face while blasting of subsequent blastholes. The delay timings are optimized based on the strata demand and the depth of the blasthole. In

70

5 Box-Cut Excavation and Stope Opening

Fig. 5.3 Deformation contour for the blast of cut holes while placement of initiation point at different positions (Himanshu et al. 2021)

Indian condition, the delay between holes for the cut holes of slot raise pattern is kept in the range of 100–500 ms. The maximum charge weight per delay (MCPD) in cut portion is kept optimum to maximize pull and minimize overbreak. The MCPD while blasting in perimeter portion is kept minimum to avoid overbreak. The delay sequence is arranged to ensure sequential excavation of cut holes and perimeter holes. The sample delay timings for a slot raise blasting pattern are shown in Fig. 5.2.

5.6 Special Slot Raise Blasting Pattern The best slot blasting practices have been reviewed from the literature. The literature search suggests that guided slot drilling method is one of the established slot blasting method. This concept was elaborated by M/s LKAB in Sweden. The method had been successfully applied at the Malmberget mine of M/s LKAB, where five 200 mm (8 in.) diameter holes are drilled adjacent to each other and are physically interconnected. This linear slot, sometimes referred to as Swedish slot, serves as a relief volume for blasting an inverse drop raise of up to 30 m. A view of drilled slot face using this method is shown in Fig. 5.4. The concept of such slot drilling is fairly straightforward. A first production hole is drilled in a normal fashion, either reamed or directly drilled to a desired diameter. Next, a specially designed unit with guided hammer is installed on a drill. The guide pipe or rod is inserted into the first hole before drilling the second. Thus, the guiding pipe keeps the second hole parallel and connected to the first one. After the second

5.7 Summary

71

Fig. 5.4 A view of drilling in guided slot drilling method

hole has been drilled, it serves as a guide hole for drilling of the third one, and so on and so forth. The guided slot drilling method is superior than the other slot excavation method as in this case, whether or not there is deviation, it will always provide a true linear slot composed of interconnected holes. To analyse the blasting outcomes under the existing slot blasting method at Indian mines and guided slot blasting method, the numerical simulation for two concepts under similar rock properties was carried out. The deformation pattern for slot blasting using existing pattern is shown in Fig. 5.3. The deformation pattern for cut hole blasting in guided slot raise blasting method is shown in Fig. 5.5. The comparison of output deformation for these two models shows that the guided slot raise blasting pattern gives more uniform deformation than the existing method. The deformation analysis also suggests that there is a prospect of further reducing the number of holes using this method. This method of slot blasting may be applied at the mine to get the optimum breakage. A sample blast design pattern with guided slot hole drilling is shown in Fig. 5.6.

5.7 Summary The summary of the slot raise blasting pattern discussed in this chapter is as follows: i.

The study under Indian conditions suggests that the optimum deformation occurs when the distance of initiation point from the plug is equal to about 20–30% of the length of the blasthole. ii. In drilling pattern, the uniformity in deformation has been seen when the ratio of the number of reamer holes to blastholes is greater than one.

72

5 Box-Cut Excavation and Stope Opening

Fig. 5.5 Deformation pattern due to blast of cut holes in guided slot drilling pattern

Fig. 5.6 Slot blasting pattern using guided slot drilling concept

iii. The delay sequence of the blastholes for slot raise blasting should be such that the sequential excavation is ensured. iv. The guided slot raise blasting method is one of the advanced methods used for maximizing damage while slot raise excavation.

References

73

References Allen RM (2014) An analysis of burn cut pull optimization through varying relief hole depths Baase RA, Diment WD, Petrina AJ (1982) Sublevel caving at Craigmont Mines Ltd. In: Underground mining methods handbook, pp 898–915 Chandrakar S, Paul PS, Sawmliana C (2023) Long-hole raise blasting in a single shot: assessment of void ratio and delay time based on experimental tests. Eng Struct 275:115272. https://doi. org/10.1016/j.engstruct.2022.115272 Dobran MF (1996) Comparison of techniques used to develop a slot raise. International Society of Explosives Engineers, Cleveland, OH (United States) Ferreira PH (2003) Improved technologies in longhole blast drilling, applied to drop raising and longhole stoping as well as the application of a small twin boom mechanized drillrig. J South Afr Inst Min Metall 103(04):233–240 Gertsch RE, Bullock RL (1998) Techniques in underground mining. In: Underground mining methods handbook. SME Himanshu VK, Mishra AK, Roy MP, Vishwakarma AK, Singh PK (2021) Numerical simulation based approach for assessment of blast induced deformation pattern in slot raise excavation. Int J Rock Mech Min Sci 144:104816. https://doi.org/10.1016/j.ijrmms.2021.104816 Liu Q, Tran H (2000) Techniques of inverse drop raise blasting and slot drilling. CIM Bull 93(1039):45–50 Liu K, Yang J, Li X, Hao H, Li Q, Liu Z, Wang C (2018) Study on the long-hole raising technique using one blast based on vertical crater retreat multiple deck shots. Int J Rock Mech Min Sci 109:52–67. https://doi.org/10.1016/j.ijrmms.2018.06.020 Liu K, Li X, Hao H, Li X, Sha Y, Wang W, Liu X (2019) Study on the raising technique using one blast based on the combination of long-hole presplitting and vertical crater retreat multiple-deck shots. Int J Rock Mech Min Sci 113:41–58. https://doi.org/10.1016/j.ijrmms.2018.11.012 Mann CD (1998) Sublevel stoping techniques in underground mining. In: Underground mining methods handbook, pp 223–227 Pakalnis RT, Hughes PB (2011) Sublevel stoping. In: SME mining engineering handbook, pp 1355–1363 Sharma PD (2005) Tunnel blasting—emulsion explosives and proper blast design are the prerequisite for better efficiency. Indian J Mines, Metals Fuel 10 Singh SP (1995) Mechanism of cut blasting. Trans Inst Min Metall 104(A1):134–138 Tatiya RR (2005) Surface and underground excavations: methods, techniques and equipment. CRC Press Villaescusa E (2014) Geotechnical design for sublevel open stoping. CRC Press Vishwakarma AK, Murthy VMSR, Himanshu VK, Roy MP (2022) Optimisation of operational parameters of a raise borer machine for rock excavation in an underground metaliferous mine. In: 9th Asian mining congress, pp 295–300

Chapter 6

Underground Ring Blasting

Abstract Ring blasting method is practised for the final excavation of rock mass in longhole stoping. The method is advantageous in achieving the faster production pace due to the use of large diameter drilling and blasting. However, this method has many associated operational challenges. These challenges have been overcome using the technological innovations. It has been possible to undertake ring blasting up to the depths of 50 m with the help of multi-deck firing. The advent of electronic delay detonators has given additional ease to the blast designers in planning and implementing of multi-ring blasting. The selection of optimal blast geometry, charging pattern and delay sequence is also important to obtain the desired output from a ring blast. Various empirical, numerical and statistical models have been developed by the researchers for the computation of these parameters. In this chapter, the process of ring blasting and designing of blasting parameters for a ring blasting has been discussed in detail. Some of the experimental data from the case studies at Indian underground mines have also been discussed in this chapter.

6.1 Introduction The ring blasting method is used for the final extraction of the orebody in longhole stoping. In this method, deep hole drilling and blasting are done to achieve a faster production pace. The blasting in this method is conducted against the free face provided by the excavated slot raises. The sufficient free faces in two or more directions are present for the accumulation of the blasted muck. Accordingly, this method can be treated similar to the bench blasting of openpit excavation. So, the challenges of ring blasting would also be similar to bench blasting. The major challenges are to improve rock fragmentation, muckpile movement and to reduce ground vibration. The flyrock ejection is not a concern in ring blasting, as the blasting area in underground would be devoid of any manpower during blasting. Despite the similarities in the challenges, the drilling of inclined holes is one of the major differences between the ring and bench blasting patterns. Accordingly, there is a need to devise a different method to optimize blasting parameters for the underground ring blasting.

© The Author(s), under exclusive license to Springer Nature Singapore Pte Ltd. 2023 V. K. Himanshu et al., Blasting Technology for Underground Hard Rock Mining, https://doi.org/10.1007/978-981-99-2645-9_6

75

76

6 Underground Ring Blasting

The main challenge of rock fragmentation improvement can be solved by optimizing charge factor and dimensional parameters. The optimum charge factor results in the proper breakage of rock strata with minimum nuisance and maximum productivity. The decision on the optimal charge factor and dimensional parameter in underground excavation focuses on reducing the boulder generation, accessibility of the excavators, maintaining the sequence of excavation and ensuring safety to the man and machinery. Researchers and underground mining practitioners have used different methodologies to optimize the blasting pattern including the charging parameters, dimensional parameters and delay design. The very basic empirical formulae for assessing the optimal burden was proposed by Rustan (1992). This relation correlates the burden for ring blasting with the hole diameter. This relationship has the limitation of not using rock and explosive parameters for the estimation of optimum burden. Also, the suitability of hole diameter for underground blasting is decided from the productivity and safety perspective. The major factor from the productivity perspective consists of the planned hole depth to be blasted in a round. The larger diameter of the blasthole is more suitable for the blast of deep holes, as the deeper hole tends to show comparatively more deviation. The expected deviation will be lesser while using the larger diameter of blastholes. However, the smaller diameter blasthole may be preferred to control the vibration for the safety of nearby structures (Himanshu et al. 2018a). Other different methods have been used by the researchers to address the limitations of Rustan (1992) model. The most of the models are objective based and suitable for a particular site. Wang et al. (2018) developed an algorithm to optimize the charging scheme of a blast with an objective to protect brow. This algorithm was based on Harries mathematical model and the superposition model. Duranovi´c et al. (2018) developed generalized empirical model for the estimation of optimum burden for ring blasting. The model consisted of rock and explosive parameters. Onederra (2004) developed a design evaluation model called “FRAGMENTO” for ring blasting. Onederra and Chitombo (2007) developed a systematic approach to assess the impacts of blast design in ring blasting on the expected fragmentation and potential damage to the surrounding rock mass. Some of the special techniques were also used by the blast designers to improve the blasting outputs in terms of rock mass damage, fragmentation and reduction of ground vibration (Dorssen et al. 2002; Sun et al. 2021; Zhang 2014; Zhang and Naarttijärvi 2005). Zhang and Wimmer (2018) used “dividing a single blast (DSB)” method for reduction of dilution while extraction of narrow ore deposits.

6.2 Drilling Pattern for Underground Ring Blasting The inclined blastholes are drilled along the extent of the orebody for underground ring blasting. However, sometimes the drilling is also required to be done in waste to completely exploit the ore body. The caution should be taken in such cases while charging of explosives. The inclined holes in the ring blasting are drilled from a

6.2 Drilling Pattern for Underground Ring Blasting

77

drivage opening. So, the blastholes seem to be converging at the drivages in this pattern. Accordingly, the pattern results into the varying spacing between the holes from toe to collar. So, the term “Toe Burden” and “Toe Spacing” are better used for ring drilling pattern instead of “Burden” and “Spacing” of bench blasting. Toe burden for a ring blasting pattern is defined as the distance between the ring blastholes and excavated slot, as well as the distance between the blastholes of two rings. The drilling geometry for ring blasting pattern consists of depth of blastholes, blasthole diameter, toe burden and toe spacing. The complete depth between two levels/sublevels of a longhole stope is blasted in a single go. Accordingly, the depth of the blastholes is taken as the level difference of the stope. The optimum level difference is selected on the basis of different criteria. The dip of the orebody is the main criteria among them. The optimum level difference is made to reduce the cost of crosscut development and transportation of the blasted ore upto the shaft/decline. The optimum hole diameter for ring blasting has dependency on the depth of the blastholes. It is difficult to drill deeper blastholes using small diameter drill holes. As the smaller drill holes will have more tendency towards deviation, than the larger diameter drill holes. The rock mass properties also affect the selection of optimum diameter of drill holes. It is difficult to drill in the harder rock formations using small diameter drill holes. From the perspective of improving the performance of a blast, the drilling of small diameter blastholes would be more helpful. The smaller geometry of the blast with small diameter blastholes would result into better rock fragmentation. It will also help in reducing the maximum explosive charge weight per delay for a blast and thereby reducing the induced ground vibration from the blast. The selection of optimum toe burden and toe spacing is another important parameter while designing the drill geometry. In bench blasting, the practitioners have correlated the burden and spacing with the blasthole diameter. The most popular computation method for optimum burden–spacing for ring blasting is also based on their relationship with the blasthole diameter. This relationship has been suggested by Rustan (1992). The empirical relation between blasthole diameter and burden as suggested by Rustan (1992) is shown in Eq. 6.1. Rustan has also suggested the maximum and minimum limits of the burden for this formulae. The maximum burden should be kept 50% more than the computed burden, and the minimum burden can be 35% less than the computed burden. Author has suggested taking toe spacing in the range of 1.5–2.0 times of the toe burden. Burden(in m) = 11.8 × φ 0.63

(6.1)

where φ = Blasthole diameter (in m) This suggested relationship however is very general in nature. So, there was a need for optimizing the burden–spacing by investigating the rock mass properties of the site and correlating it with the expected fragmentation from the blast. Himanshu et al. (2021) developed an algorithm for the purpose of optimizing dimensional

78

6 Underground Ring Blasting

parameters for ring blasting. The algorithm is based on the empirical Kuz-Ram model. This empirical model consists of three different equations as Kuznetsov’s equation, Uniformity index equation and Rosin–Rammler equation. Kuznetsov’s equation correlates the mean fragment size from the blasting with rock factor and explosive parameters. The rock factor in this equation may be computed using Lilly’s blastability index. The uniformity index equation correlates the dimensional parameters of blast with an index. This index further correlates the mean fragment size from the blasting with desired percentage of fragment to be retained on the screen of a defined size. Authors in the developed algorithm suggested to compute rock factor for the site of blasting using blastability index. Then to assume the expected mean fragment size and 90% passing fragment size (Assuming that only 10% oversize boulder is allowed from blasting). The optimum charge factor is back-calculated using Kuznetsov’s equation with assumed mean fragment size value. Further, the uniformity index is computed using Rosin–Rammler equation with the assumed mean fragment size and 90% passing fragment size. The computed index is used in uniformity index equation for back-calculating the optimum burden and spacing. However, many assumptions are made for computing burden–spacing using this approach. But, the approach gives an initial idea about the drilling geometry for ring blasting under different rock mass condition.

6.3 Charging of Ring Blastholes The charging of ring blastholes is performed either using cartridge emulsion/slurry explosives or using site mixed emulsion (SME)/site mixed slurry (SMS) explosives or using Ammonium Nitrate–Fuel Oil (ANFO) explosives. The explosive charging using ANFO or SME/SMS is preferred, as it is faster as compared to the manual charging of blastholes using cartridge explosives. The charging is performed either from upper level to lower level (downhole charging) or from lower level to upper level (uphole charging). Sometimes the combination of downhole and uphole charging is used for the excavation of stopes. The compressed ANFO explosives are used for charging in the uphole blastholes. The compression provides the adhesiveness to the explosives, and thereby it remains intact to the blastholes. A view of explosive charging arrangement using ANFO for an Indian underground mine is shown in Fig. 6.1. The SME/SMS explosives with high viscosity are used for uphole charging. The explosives come with high temperature for this purpose. The higher viscosity of the explosives helps it to attach to the wall of blastholes and thereby works as the resistance against free fall under gravity. A view of uphole charging using SME/SMS at an Indian underground mine is shown in Fig. 6.2. Since the spacing between the holes of a ring varies between toe to collar of the blasthole, hence the spacing between the holes near the collar is very less. The uniform charging all along the blastholes in this case will lead to over-charge concentration at some of the zone. So, the differential charging or alternate hole charging is done

6.3 Charging of Ring Blastholes

79

Fig. 6.1 Explosive charging arrangement using ANFO for uphole charging in an Indian underground mine Fig. 6.2 Uphole explosive charging using SME/SMS at an underground mine

80

6 Underground Ring Blasting

Fig. 6.3 Differential charging pattern for ring blasting (Himanshu et al. 2021)

in the zone with reduced spacing. A schematic of differential charging pattern for ring blastholes is shown in Fig. 6.3.

6.3.1 Suitability of Explosive for Different Rock Strata The breakage mechanism of rock under explosive loading includes the tensile cracking under the impact of detonation pressure and final breakage under the impact of gaseous energy. The optimum requirement of detonation pressure and gaseous energy is different for the different types of rock. The elastic nature of rock requires more detonation pressure for its state change to plasticity. The plastic (spongy) rock type needs gaseous energy-based breakage in order to get optimum blasting output with minimal hazards. The optimum detonation pressure for the respective rock type is computed by evaluation of the characteristic impedance of explosive and rock media. Characteristic impedance of explosive and rock mass represents optimization of seismic amplitude. The maximum seismic amplitude can be obtained with unit impedance coupling. This impedance coupling is the ratio of characteristic impedance of explosive and characteristic impedance of media. This can be determined by Eqs. 6.2 and 6.3. Characteristic impedance of explosive Characteristic impedance of media (Loading density of explosive charge) × (Detonation Velocity of explosive charge) Impednace = (Density of rock) × (P − wave velocity of rock) Impedance =

(6.2) (6.3)

The evaluation of suitability of explosive for the respective rock types of an Indian underground mine was done. For this purpose, the core samples of Ore, GarnetBiotite-Sillimanite-Gneiss (GBSG), Quartzite-Biotite-Gneiss (QBG) and Paste were collected from the mine management for the assessments of rock mass properties.

6.3 Charging of Ring Blastholes

81

Sometimes the development drivages at this underground mine are also made in paste backfilled area. So, the suitability of explosive for blasting in paste-filled zone was also evaluated. Density, uniaxial compressive strength (UCS), Young’s modulus, Poisson’s ratio, P-wave velocity and S-wave velocity of the core samples were tested at Rock Mechanics Laboratory of CSIR-CIMFR, Dhanbad, for this purpose. Summary of the tested rock mass properties for the core samples is shown in Table 6.1. The suitable detonation velocity for the tested rock types has been computed considering the impedance theory. The suitable correction factor for insitu p-wave velocity has been added during the computation. The computed detonation velocity for each rock type is shown in Table 6.2. The computed results show that there is requirement of very low VOD for excavations in paste. The practical availability of such low VOD is not possible. But, for the purpose of optimization, comparatively low VOD and low-density explosive should be preferred for blasting in paste. The other means of reducing damages by design alteration may be explored for the improvements in blasting outputs, while blasting in paste zone. Table 6.1 Tested rock mass properties for different rock and paste samples Rock type

Density (dry) (kg/m3 )

Uniaxial Young’s compressive modulus strength (MPa) (GPa)

Poisson’s ratio

P-wave velocity (m/s)

S-wave velocity (m/s)

Ore

3077

59

12.54

0.25

4563

2004

GBSG

2828

35

5.87

0.25

4308

1769

14.41

0.29

3327

1658

0.28

0.35

1126

826

QBG

2660

89

Paste

1391

0.74

Table 6.2 Suitable detonation velocity of explosive for different rock and paste strata Rock type

Density (dry) (kg/m3 )

P-wave velocity (m/s)

Suitable detonation velocity of explosive (m/s)

Ore

3077

4563

4500–5500

GBSG

2828

4308

4000–4500

QBG

2660

3327

3500–4000

Paste

1391

1126

< 2000 (ANFO explosive should be preferred)

82

6 Underground Ring Blasting

6.3.2 Charge Factor and Dimensional Parameters for Ring Blasting The optimum amount of explosive charge for ring blasting is to enhance rock fragmentation, improve muckpile distribution and reduce induced ground vibration. The induced ground vibration may also be reduced by distributing the explosive charge and thereby reducing the maximum explosive charge weight per delay (MCPD). The detailed discussion on this has been made in Chap. 7. From the fragmentation and distribution of muckpile perspective, the charge factor and dimensional parameters play important role. The charge factor can be defined as amount of explosive required for the breakage of one cubic meter volume of the rock mass. The optimum charge factor results in the proper breakage of rock strata with minimum nuisance and maximum productivity. Researchers and blasting practitioners have used different methods for optimizing the charge factor. The charge factor for four different classes of rock, viz. hard, medium, soft and very soft, has been defined by Dyno Nobel (2010) blasting reference manual (Table 6.3). Jimeno et al. (1995) have suggested the charge factor based on the rock mass properties of the blast face. The charge factor classification under this literature has been done for different distances between natural fractures of the rock mass, different uniaxial compressive strength values and rock density values. The suggested charge factor as per Jimeno et al. (1995) is shown in Table 6.4. Broadbent (1974) correlated the insitu P-wave velocity with the charge factor for an openpit copper mine. The correlation is shown in Eq. 6.4. Further, the approach was used by Muftuoglu et al. (1991) for overburden strata in lignite/coal mines. Adhikari (1994) reviewed some noteworthy approaches such as transfer of energy approach, based on drilling data and assessment of rock quality index, blastability index approaches for the assessment of charge factor. Himanshu et al. (2021) computed the optimum charge factor for a Lead–Zinc underground mine. The computation was made using an algorithm based on empirical Kuz-Ram model. The comparison of optimum charge factor for an Indian Lead–Zinc underground mine using different approaches is given in Table 6.5. The rock mass at this mine had severe jointing with joint dipping inside the face and having joint spacing of less than 0.1 m. The average elastic modulus of rock was 15 GPa, and average UCS was 90 MPa. The rock density was 2.84 tonne/cu-m. The P-wave velocity of the rock was 4500 m/s. K = 0.00017 × V P − 0.13,

(6.4)

where K = Charge factor (in kg/m3 ) Vp = P-wave velocity of rock (in m/s) Dimensional parameter for a blast is indirectly related to the charge factor. It includes the blast geometry, viz. burden, spacing, explosive column length, stemming length, etc. Rules of thumb used by different rock blasting practitioners have related

6.3 Charging of Ring Blastholes

83

Table 6.3 Charge factor for different rock types as per Dyno Nobel quick reference guide (Dyno Nobel quick reference guide 2010) Charge factor (kg/m3 )

Rock type Hard

0.7–0.8 0.4–0.5

Medium Soft

0.25–0.35

Very soft

0.15–0.25

Table 6.4 Charge factor classification based on the geotechnical properties of the rock strata (Jimeno et al. 1995) Charge factor Class limit (kg/m3 )

Average value (kg/m3 )

Mean distance between natural fractures in rock mass (m)

Uniaxial compressive rock strength (MPa)

Rock density (t/m3 )

0.12–0.18

0.150

< 0.10

10–30

1.40–1.80

0.18–0.27

0.225

0.10–0.25

20–45

1.75–2.35

0.27–0.38

0.320

0.20–0.50

30–65

2.25–2.55

0.38–0.52

0.450

0.45–0.75

50–90

2.50–2.80

0.52–0.68

0.600

0.70–1.00

70–120

2.75–2.90

0.68–0.88

0.780

0.95–1.25

110–160

2.85–3.00

0.88–1.10

0.990

1.20–1.50

145–205

2.95–3.20

1.10–1.37

1.235

1.45–1.70

195–250

3.15–3.40

1.37–1.68

1.525

1.65–1.90

235–300

3.35–3.60

1.68–2.03

1.855

> 1.85

> 285

> 3.55

Table 6.5 Optimum charge factor for a Lead–Zinc underground mine using different approaches Charge factor (in kg/m3 ) Dyno Nobel quick reference guide, 2010

0.5–0.6

Jimeno et al. (1995)

0.45

Broadbent (1974)

0.635

Himanshu et al. (2021), Emulsion explosive

0.61

Himanshu et al. (2021), ANFO explosive

0.73

the dimensional parameters with hole diameter. Hole diameter is, however, decided based on the bench height/stope height. The assessment seems good from the production planning perspective. However, the blasting practices with the most sophisticated modern drilling equipment have surpassed the rule of thumb. The dimensional parameters in such cases must be decided on the basis of assessment of rock mass and geo-mining conditions. The dimensional parameters such as burden and spacing for underground ring blasting have been related with hole diameter by Rustan (1992).

84

6 Underground Ring Blasting

Table 6.6 Optimum dimensional parameter for ring blasting at an Indian Lead–Zinc mine using different approaches Hole diameter (mm) 70

Toe burden (m)

Toe spacing (m)

Using Rustan (1992)

Using Himanshu et al. (2021)

Using Rustan (1992)

Using Himanshu et al. (2021)

2.2

1.8

3.3

1.9

76

2.3

2.0

3.5

2.1

89

2.6

2.3

3.9

2.4

115

3.0

3.0

4.5

3.1

This relationship however is generalized in nature and doesn’t correlate the dimensional parameters with rock parameters, explosive parameters and expected fragment size. Himanshu et al. (2021) computed the optimum dimensional parameter using an algorithm based on empirical Kuz-Ram model. The algorithm computes the parameters to achieve a desired rock fragmentation size from the blasting. The comparison of optimum dimensional parameters for a Lead–Zinc underground mine having different blasthole diameter is shown in Table 6.6.

6.4 Connection and Firing Sequence of Charged Ring Blastholes The charged ring blastholes are connected using non-electric (NONEL) delay detonators or electronic detonators. The long delay interval is used in the ring blasting. The initiation system is connected in the blastholes using cap-sensitive explosives. Cast/emulsion boosters are used as cap-sensitive explosives for the blasting faces with ANFO and SME explosives. The timing of NONEL detonators is preset. However, the timing of electronic delay detonators is programmed judiciously as per the requirement at the blast face. The shock tubes of detonators are connected on a common harness wire or detonating fuse or another cable of electronic detonator. Such wire or fuse is further connected through an electric detonator, which is fired from a safe distance from the blasting face. When the connection is made on the wire of electronic detonator, then the firing is done using logger of electronic system from a safe distance from the blasting face. A view of connection of charged uphole ring blastholes using electronic detonator is shown in Fig. 6.4.

6.4 Connection and Firing Sequence of Charged Ring Blastholes

85

Fig. 6.4 View of connection of charged blastholes of a ring blasting face using electronic detonator

6.4.1 Optimization of Delay Sequence and Delay Timing for Production Blasting The sequence of delay for stope blasting is decided to maximize the fragmentation and reduce the damages in the waste rock. In a ring blasting, the length of explosive charging in the blastholes varies depending on their position and inclination. The holes are drilled to cover all the portions of the orebody. Sometimes, the drilling has to be made in waste as well. The length of the drilled holes in the waste portion is not charged. The hole with maximum length of the explosive charge is taken as the initiation hole in the blast. This is done to create maximum relief area for the subsequent round of blasting. In the next round, the blastholes are fired along the free face created by the first hole. The blastholes of hangingwall and footwall side are fired in alternate. A sample blast design with firing sequence of blastholes for a ring blasting is shown in Fig. 6.5. The delay timings of the blastholes for the sample blast design shown in Fig. 6.5 are given in Table 6.7. The minimum delay timing between the ring holes was given 40 ms in this design. The delay timings for the firing of first holes were kept on higher side, as the free face direction is only along the excavated slot. After that, the timings between the holes from two directions were kept 40 ms. The higher delay intervals were used for the blasts of periphery blastholes. The delay interval has also been increased, if the movement of the holes has to be kept in single direction. However, for multi-hole blasting the delay timing for two holes along two extremities may be kept same, as the firing holes will not increase the intensity of vibration in the contact of wall rock in such case. The simultaneously firing holes in such cases may also result into the improvement of fragmentation due to the interaction between the blasted rocks. The optimum delay timing for the ring blasting serves two important purposes: (i) Reduction of induced ground vibration near the structures. (ii) Reduction of induced vibration and thereby damage in the wall rock.

86

6 Underground Ring Blasting

Fig. 6.5 Sample design of a ring blasting (Himanshu et al. 2018b) Table 6.7 Delay timings of the blastholes for ring blasting pattern shown in Fig. 6.5

Hole No.

Stemming column length (m)

Delay timing (ms)

4

2

260

H2

8

3

160

H3

12

4

80

H4

15

6

0

H5

14

6

120

H6

13

6.5

200

H7

12

5

300

H8

8

4

400

H9

4

4

500

H10

1

2

600

H1

Explosive charge length (m)

6.4 Connection and Firing Sequence of Charged Ring Blastholes

87

To achieve the above-mentioned purposes, the delay timings are optimized using the waveform analysis of the recorded vibration data. The nearfield ground vibration waveform gives a better idea regarding the optimum delay from the blasting. The delay sequence for a uphole ring blasting pattern using NONEL delay detonators is shown in Fig. 6.6. Two rings containing nine holes in each ring were used in this pattern. Two periphery holes in this pattern were kept uncharged. These uncharged holes work as line drilling holes and work to restrict the damage in the ore body only. This design pattern was validated at an underground Lead–Zinc mine. The nearfield ground vibration was recorded for the validation. The recorded vibration waveform from one of the blast is shown in Fig. 6.7. The charging pattern was compared with the vibration waveform. The analysis of this waveform reveals that the maximum vibration due to blast was 28.32 mm/s at 193 ms and 26.16 mm/s at 125 ms. Comparing the result with charging and initiation pattern of blastholes, it can be drawn that 28.32 mm/s vibration has resulted due to blast of two simultaneous holes with total explosive charge of 32 kg. The blast vibration of 26.16 mm/s has resulted due to blast of one hole with total explosive charge of 24 kg. This means that vibration level was almost same in both the cases even if the explosive charge in first case is much higher than second case. This is due to the movement of blastholes along free face created by blasting of first three holes. Hence, the design pattern is optimum for the blasting at the study site. Sometimes the excavation of orebody is done using both uphole and downhole rings. The drilling in such cases is done from a sublevel. This pattern gives ease to the management in drilling, explosive charging and connection from a single level.

Fig. 6.6 Delay sequence of blastholes for an uphole ring blasting face

88

6 Underground Ring Blasting

Fig. 6.7 Nearfield vibration waveform for an experimental blast conducted using blast design pattern shown in Fig. 6.6

The mucking of blasted rock in such cases is also done from a single (bottom most) level, which is an added advantage. This pattern helps in faster excavation. The delay sequence in such cases is planned such that all the holes of downhole ring are fired before the firing of first hole of uphole ring. This is done to ensure that there is no connection cut due to blasted rock from uphole ring. The firing of downhole ring initially also provides the additional free face to the uphole ring. A sample design pattern for the simultaneous excavation of uphole and downhole rings is shown in Fig. 6.8.

6.4.2 Delay Timing for Multi-Ring Blasting The multi-ring blasting pattern is used for faster excavation of rock. The optimization of delay interval between two rings is important in this case. The main issue associated with the multi-ring blasting is the intensity and substance of vibration near important underground and surface structures. The detailed discussion on this topic has been made in Chap. 7. The delay timing between rings for multi-ring blasting is generally kept such that the impact of the blasting of first ring near structure is completely dampened before the firing of the subsequent ring. A sample blast design with delay timings of holes for multi-ring blasting is shown in Fig. 6.9. The delay interval for second hole after initiation hole was 60 ms. Further, the delay timings between holes of a ring were kept 40 ms in this pattern. Larger jump in delay interval has been used in some rings for the periphery holes. However, the periphery holes of two sides were fired at same delay interval. The delay timing between first to second ring in this pattern has been kept 500 ms; assuming that the complete dampening of vibration

6.5 Summary

89

Hole No. H1 H2 H3 H4 H5 H6 H7 H8 H9 H10 H11 H12 H13 H14 H15 H16 H17 H18

Delay No. 7 6 5 3 1 2 5 6 7 13 12 11 10 8 9 11 12 13

Fig. 6.8 Sample blast design pattern for simultaneous firing of uphole and downhole rings

wave will occur below 500 ms. The larger delay timing for the last ring has been given to avoid any possibility of backbreak generation. So, the delay timing between second and third rings in this pattern has been kept 1000 ms.

6.5 Summary The ring blasting pattern is practised for rock excavation in longhole stoping. The detailed discussion on the ring blasting practices has been made in this chapter. The summary of the discussions is as follows: I.

The inclined blastholes are drilled in ring blasting pattern to cover the complete extent of the ore body. The toe burden and toe spacing are maintained during the drilling of the holes to improve fragmentation from blasting. II. The blastholes of a ring are charged using slurry/emulsion cartridges or SME/SMS/ANFO. The suitability of the explosive for strata is determined using impedance matching. III. The charged blastholes are connected using NONEL or electronic delay detonators. The firing sequence of the blastholes is optimized to utilize the advantage of tensile cracking against the free face. The delay timings between the holes

90

6 Underground Ring Blasting

Fig. 6.9 Sample delay pattern for multi-ring blasting (Roy et al., 2022)

of a ring are optimized to reduce the breakage in the wall rock and maximize the ore breakage. IV. The ring blastholes are drilled in uphole or downhole pattern based on the requirement at the site. Sometimes uphole and downhole rings are blasted simultaneously from a sublevel. Such pattern gives advantage of faster excavation and reduction in cost of production by the deployment of machineries and excavator at a single level. V. Multi-ring blasting is also preferred sometimes to achieve a faster pace of production. The delay timing between the rings is optimized in such cases to completely dampen the impact of first ring before the firing of subsequent ring.

References Adhikari GR (1994) Controlled blasting in tunnel—some issues. Tras Inst Eng (I) MN 75:56 Broadbent CD (1974) Predictable blasting with in-situ seismic surveys. Mining Eng 26:37–41 Dorssen P, Van Valicek P, Farren M, Harrison G, Joubert W, Pickering RGB, Van Rensburg HJ (2002) A long hole stoping system for mining narrow platinum reefs. J Southern African Instit Min Metall 102(3)(September 2001):151–154 Duranovi´c M, Ðoki´c N, Lapˇcevi´c V, Torbica S, Petrovi´c M, Savi´c L (2018) Optimization of ring blasting in sublevel stoping gold mine. Podzemni Radovi 33:61–68. https://doi.org/10.5937/Pod Rad1833061D

References

91

Dyno Nobel (2010) Blasting and explosives: quick reference guide. https://www.leg.mn.gov/docs/ 2015/other/150681/PFEISref_1/Dyno%20Nobel%202010.pdf Himanshu VK, Roy MP, Mishra AK, Paswan RK, Panda D, Singh PK (2018a) Multivariate statistical analysis approach for prediction of blast-induced ground vibration. Arab J Geosci 11(16):460. https://doi.org/10.1007/s12517-018-3796-8 Himanshu VK, Roy MP, Sarim Md, Singh PK, Mishra AK, (2018b) Blast design for underground long hole stope to safeguard surface industrial structures from blast induced vibration. In: 12th International symposium on rock fragmentation by blasting, Lulea, Sweden, pp 257–267 Himanshu VK, Roy MP, Shankar R, Mishra AK, Singh PK (2021) Empirical approach based estimation of charge factor and dimensional parameters in underground blasting. Mining Metall Explorat 38(2):1059–1069. https://doi.org/10.1007/s42461-020-00374-8 Jimeno CL, Jimeno EL, Carcedo FJA (1995) A review of drilling and blasting of rocks. Int J Surf Mining Reclamat Environ 10,1:V–VI. https://doi.org/10.1080/09208119608964786 Muftuoglu YV, Amehmetoglu AG, Karpuz C (1991) Correlation of powder factor with physical rock properties and rotary drill performance in Turkish surface coal mines. Int Soc Rock Mech Rock Eng 1–3 Onederra I (2004) A fragmentation modelling framework for underground ring blasting applications. Fragblast 8(3):177–200. https://doi.org/10.1080/13855140412331333262 Onederra I, Chitombo G (2007) Design methodology for underground ring blasting. Min Technol 116(4):180–195. https://doi.org/10.1179/174328607X282244 Roy MP, Himanshu VK, Kaushik AP, Singh PK (2022) Influence of ring blasting pattern on the safety of nearby underground structures. Sadhana 47:192. https://doi.org/10.1007/s12046-02201968-2 Rustan RA (1992) Burden, spacing and borehole diameter at rock blasting. Int J Surf Min Reclam Environ 6(3):141–149. https://doi.org/10.1080/09208119208944329 Sun M, Ren F, Ding H (2021) Optimization of stope structure parameters based on the mined orebody at the Meishan iron mine. Adv Civil Eng 2021:1–14. https://doi.org/10.1155/2021/805 2827 Wang M, Shi X, Zhou J (2018) Charge design scheme optimization for ring blasting based on the developed Scaled Heelan model. Int J Rock Mech Min Sci 110:199–209. https://doi.org/10. 1016/j.ijrmms.2018.08.004 Zhang ZX (2014) Effect of double-primer placement on rock fracture and ore recovery. Int J Rock Mech Min Sci 71:208–216. https://doi.org/10.1016/j.ijrmms.2014.03.020 Zhang ZX, Naarttijärvi T (2005) Reducing ground vibrations caused by underground blasts in LKAB Malmberget mine. Fragblast 9(2):61–78. https://doi.org/10.1080/13855140500140275 Zhang Z-X, Wimmer M (2018) A case study of dividing a single blast into two parts in sublevel caving. Int J Rock Mech Min Sci 104:84–93. https://doi.org/10.1016/j.ijrmms.2018.02.002

Chapter 7

Blast-Induced Hazards

Abstract Large-scale drilling and blasting are the need of the hour to maintain the production pace from the underground mining. Such large-scale blasting may be a threat to the structures. Accordingly, there is a need to devise the controlled blasting pattern to safeguard such structures. The structures to be protected from blast-induced hazards include the surface residential structures, industrial buildings and underground structures. The threshold ground vibration limits for different nature of structures would be different. So, the designing of controlled blasting pattern includes the assessment of threshold vibration limit near the structure and thereby designing the charging and delay pattern to restrict the ground vibration within the threshold limit. Over the years, different techniques have been evolved to enhance the productivity while maintaining the ground vibration within safe limits. Such techniques include the multi-deck excavation, multi-ring blasting and ring slicing. The advanced statistical analysis is used to optimize the charging pattern in order to achieve the challenge of restricting vibration within the limits. In this chapter, the detailed discussion has been made on the ground vibration propagation characteristics for underground ring blasting, blast vibration threshold for different structures and the techniques for large-scale underground blasting.

7.1 Introduction Quality of a blast is measured in terms of productivity and magnitude of associated blasting hazards. The rock blasting phenomena consists of hazards in the form of blast-induced ground vibration, air-overpressure/noise, flyrock ejections, dust and other environmental concerns. Blast design parameters are optimized to improve the productivity of the mine while minimizing the associated hazards. Some of these hazards are due to the release of huge amount of instant explosive energy; however, others are directly associated with the insitu rock mass. The hazards in the form of explosive energy release can be reduced by efficient utilization of explosive energy for the breakage of insitu rock mass. The findings from different literatures suggest that only a small fraction of the explosive energy (20–30%) is useful for rock breakage and the rest is considered as waste energy (Monjezi et al. 2011). The waste explosive © The Author(s), under exclusive license to Springer Nature Singapore Pte Ltd. 2023 V. K. Himanshu et al., Blasting Technology for Underground Hard Rock Mining, https://doi.org/10.1007/978-981-99-2645-9_7

93

94

7 Blast-Induced Hazards

energy contributes to the undesirable phenomena such as ground vibration, flyrocks, air blast and noise. The degree of impact of different undesirable phenomena of the blast will vary based on the nature of blasting work. All the types of excavation work focus to reduce the blast vibration near the houses/structures, as high levels of structural vibration caused by ground vibration from blasting can result in damage to, or failure of, structures (Singh et al. 2008). The blasting hazards in the form of flyrocks, air blast and noise are other major concerns for openpit excavation. The control over these hazards can be achieved by proper planning of blasts with optimum stemming, proper inspection of blast face before blasting, use of proper (low noise) initiation system, etc. However, the reduction of ground vibration hazards needs consideration of different controllable and uncontrollable parameters. The identification of these parameters and thereby investigations of the techniques to reduce ground vibration is the prime need for the mining industries.

7.2 Standards on Ground Vibration Limits Different standards on ground vibration limits have been fixed by researchers, organizations and government agencies. Much of this work originated in the mining industry, where vibration from blasting is a critical issue. Dowding (1985) found in the probabilistic study that no damage or cosmetic cracking in the structures takes place, when PPV is below 12 mm/s. The PPV data considered in this study was of very low frequencies (below 4 Hz). The low frequencies of the ground vibration may resonate with the frequencies of the structures. The frequency of the residential structures is generally in the range of 3–8 Hz. As per Siskind et al. (1980), no visible blast-induced cracks are generated, if the peak particle velocity is below 19 mm/s. Singh et al. (2008) and Singh and Roy (2010) conducted extensive studies to document the levels of vibrations causing cosmetic, minor and major damage to the different types of structures. The findings from their study are summarized in Table 7.1. Although the detailed investigations by Singh et al. (2008) and Singh and Roy (2010) have provided the PPV threshold for different nature of cracking in different structures, the statutory standards mostly provide the PPV threshold for different structures under different variations of associated frequency. The frequency in such cases is considered for the ground motion. It is expected that the low-frequency vibration will have greater influence on the structural damage particularly in case of multi-storied buildings. In Indian condition, the threshold vibration limits for the safety of structures have been fixed by Directorate General of Mines Safety under DGMS_ (Tech) (S&T) Circular No. 7 dated 29th August 1997. The limits (Table 7.2) under this circular are depending on the type of structures and dominant excitation frequency. The review of different standards framed for different countries revealed that PPV below 2.0 mm/s does not possess damage potential for any sensitive structures irrespective of any frequency. As the frequency of vibration increases, the threshold

7.2 Standards on Ground Vibration Limits

95

Table 7.1 Level of vibrations causing cosmetic, minor and major damages to different types of structures (Singh et al. 2008; Singh and Roy 2010) Type of structure

Floor

Mudhouse

Ground floor

Brick mudhouse

RCC Structure

Ground floor

Ground floor

First floor

Nature of cracking

PPV Level (mm/s)

Minor damage

55.0–56.1

Major damage

87.1–104.0

Cosmetic crack

51.6–56.3

Minor damage

81.0–89.7

Major damage

99.6–113.0

Cosmetic crack

68.6–71.3

Minor damage

104.0

Major damage

122.0

Cosmetic crack

71.2–72.2

Minor damage

98.3–118.0

Major damage

128.9–161

Table 7.2 Permissible peak particle velocity (PPV) in mm/sec at the foundation level of structures as per DGMS circular 7 of 1997 (DGMS Circular No. 07, 1997) Type of Structures

Dominant excitation frequency, Hz < 8 Hz

8–25 Hz

> 25 Hz

(A) Buildings/structures not belong to the owner Domestic houses/structures (Kuchha brick and cement)

5

10

15

Industrial buildings (RCC and framed structures)

10

20

25

Objects of historical importance and sensitive structures

2

5

10

(B) Building belonging to owner with limited span of life Domestic houses/structures (Kuchha brick and cement)

10

15

25

Industrial buildings (RCC and framed structures)

15

25

50

value for damage also increases. Indian standard as mentioned in Table 7.2 also recommends 2.0 mm/s of PPV for objects of historical importance and sensitive structures in a frequency range below 8 Hz. As the frequency increases above 8 Hz, the limiting PPV value is 5 and 10 mm/s for frequency above 25 Hz. If the structures are made up of insitu rock of underground mine, then the threshold vibration limits would be different from that suggested under DGMS circular 7 of 1997. Such structures may include underground haulage drivages, shaft, decline, workshop, electrical substations, repair shops, etc. The threshold vibration limits for such structures have been framed by Directorate General of Mines Safety (DGMS) Tech circular No. 06 of 2007. This circular describes the vibration limits around different structures based on rock mass rating (RMR), which is a good indicator of insitu rock condition. Although the PPV limits under this circular have been given on the basis of a scientific study conducted to study the influence of opencast blasting on

96

7 Blast-Induced Hazards

Table 7.3 Threshold values of vibration for the safety of roof in the below-ground workings for different RMR (DGMS Circular No. 06, 2007) RMR of roof rock

Threshold value of vibration in terms of peak particle velocity [mm/s]

20–30

50

30–40

50–70

40–50

70–100

50–60

100–120

60–80

120

Table 7.4 Threshold values of vibration for the safety of sidewall in the below-ground workings for different RMR (DGMS Circular No. 06, 2007) RMR of roof rock

Threshold value of vibration in terms of peak particle velocity [mm/s]

20–30

20

30–40

20–30

40–50

30–40

50–60

40–50

60–80

50

the stability of the underground structures for a coal mine, the RMR-based threshold presents its generalized nature, and it may also be used to ensure the stability of the structures of the underground metalliferous mine (Roy et al. 2022). RMR incorporates the strength of the intact rock, structural discontinuities, groundwater conditions, etc. (Bieniawski 1968). The study suggests that the magnitude of PPV at roof, floor and sidewall of the underground workings from the same blast would be different. Accordingly, this circular provides the threshold for the safety of roof and sidewall as shown in Tables 7.3 and 7.4, respectively. While measurement of ground vibration in an underground metalliferous mine working, the drill level in case of downhole ring may be considered as roof. In case of uphole ring, the roof will be at the level where the ring holes would be through. This consideration has been made on the basis that the maximum propagation of vibration would take place in continuity. The voids in draw level will tend to reduce the magnitude of PPV.

7.3 Sensitive Structures Prone to Ground Vibration Hazards Various structures, viz. surface residential structures, surface Industrial structures and underground structures, are prerequisite to be protected from blast-induced ground vibration. The threshold limits for the surface domestic structures not belonging to owner are taken as per DGMS circular 7 of 1997. The circular suggests the vibration

7.3 Sensitive Structures Prone to Ground Vibration Hazards

97

limit as 5 mm/s, 10 mm/s and 15 mm/s for the associated blast frequency of < 8 Hz, 8–25 Hz and > 25 Hz, respectively. Some of the mining industries have associated industrial buildings at the surface. These industrial buildings mainly include crusher house, mill, tailing pond, offices, etc. Since most of these industrial buildings consist of vibration generating machineries, the expected threshold to sustain vibration in these structures would be more. To understand the proneness of damage of these structures from induced ground vibration, Himanshu et al. (2018b) measured the natural frequency of such structures of an underground mine. The natural frequency is the number of times per second that the structure will vibrate back and forth when excited, and the damping of the structures is proportional to the rate at which the vibration decays with time. The measured natural frequency for some of the industrial structures of an underground mine is shown in Table 7.5. The table shows that the natural frequency of these structures is in the range of 6–13 Hz. Accordingly, the frequency of blast vibration more than 25 Hz will ensure no resonance to the surface structures. The threshold vibration limits for industrial buildings are taken from DGMS circular 7 of 1997. The circular suggests the threshold limits as 10 mm/s, 20 mm/s and 25 mm/s for the associated frequency of < 8 Hz, 8–25 Hz and > 25 Hz. The dynamic loading due to large-scale blasting in the stoping operation also influences damages in the nearby long-term underground structures, viz. shafts, declines, haulage drivages, underground workshop, electrical substations, repair shop, underground crusher house, etc. These structures are also required to be safeguard from induced ground vibration. The threshold vibration limits to protect such structures are assessed based on various damage criteria under dynamic loading or statutory norms. The scientific theories state that the damages in the underground structures occur due to the redistribution of insitu stresses. Such redistribution takes place because of the mining activity. The blast loading conditions also contribute in the redistribution of the stresses. The redistributed stresses reduce the sustenance of rock strata against failure (Yang et al. 2017). The threshold sustenance of vibration by rock strata is dependent on the strength of the rock mass to resist failure against incoming vibration Table 7.5 Natural frequency of different surface structures of an underground mine (Himanshu et al. 2018b) Structures 1. Power Station

2. Mine Office

Approximate height of the structures (in m)

Natural frequency (in Hz)

Roof

12–13 m

6.25

Second floor

8–9 m

6.13

First floor

4–5 m

6.25

Roof

11–12 m

First Floor

4–5 m

12.5 12.6

3. Canteen

Roof

4–5 m

4. S K Mill office

Roof

10–11 m

6.13

7.13

98

7 Blast-Induced Hazards

waves (Xie et al. 2016). The ratio of the calculated strength and stress on the structures to be protected (i.e. pillars, roof, drivages, etc.), gives safety factor for the stability of the particular structure. Based on this concept, the designing of controlled blasting pattern with an aim to protect the nearby underground structures is done to keep the combined effect of stress redistribution and blast loading below the strength of the rock mass. Various empirical formulae are used for the assessment of the strengths of the underground pillars and other structures. However, the estimation of the strengths of different underground structures with the acceptable accuracy is a difficult task. So, the existing statutory norms (DGMS circular 7 of 2007) are a better option for the determination of threshold vibration limits around different underground structures.

7.4 Optimisation of Ring Blasting Parameters for Structural Safety The ring blasting parameters are optimized to restrict the vibration near the structures within the threshold limits. For this purpose, the factors affecting the magnification and attenuation of ground vibration wave are identified.

7.4.1 Parameters Affecting Blast-Induced Ground Vibration There are three types of parameters responsible for amplification and attenuation of blast vibration, viz. rock mass properties, propagation media for blast vibration waves and blast design parameters including explosive parameters (Gorai et al. 2021; Gui et al. 2017; Hao et al. 2001; Himanshu et al. 2018a; Kumar et al. 2016; Resende et al. 2014; Singh et al. 2015; Wu et al. 1998). The rock mass properties and nature of propagation media are uncontrollable parameters. However, the blast design parameters are controllable parameter. The suitable designing of controllable parameters based on the assessment of uncontrollable parameters can reduce the induced blasting hazards near the structures. Researchers have considered maximum explosive charge weight per delay (MCPD) and distance of the blast face from the structures (D) as the most dominant parameter influencing blast-induced ground vibration (Ambraseys and Hendron 1968; Duvall and Fogelson 1962; Pal Roy 1991). Researchers have established the empirical relations correlating PPV with MCPD and D. Such empirical relations have been made considering that the most of the other parameters for mine would be constant. However, with the advancement in technology, the mining industries are now using variability in blasting parameters based on the availability of drilling equipment, strata condition and other associated parameters for same mine. Accordingly, the researchers have incorporated other design parameters, viz. burden, spacing, hole diameter, hole depth, etc., also in the predictor. Such predictors have been established with the help of multivariate statistical analysis and soft

7.4 Optimisation of Ring Blasting Parameters for Structural Safety

99

computing tools such as artificial neural network, random forest and K-nearest neighbour (Bahrami et al. 2011; Hasanipanah et al. 2015; Tian et al. 2019; Verma and Singh 2011; Wen et al. 2020; Zhou et al. 2021). These statistical algorithms predict PPV on the basis of analysis of data for the trial blasts conducted at the site.

7.4.2 Case Studies on Designing Controlled Blasting Pattern for Underground Ring The experimentation and modifications in blast design for controlling the induced ground vibration near structures within safe limits have been discussed in this section.

7.4.2.1

Case Study-1: Designing of Controlled Blasting Pattern for Reducing Vibration Near Surface Residential Structures

Roy et al. (2016) designed the controlled blasting pattern for the safety of surface residential structures due to blasting at Kayad underground mine. Authors used the technique to reduce the MCPD by multi-deck blasting technology. The electronic delay detonators were used at the mine for this purpose. The blasthole with maximum length was used as initiation hole. The delay sequence was given to fire the bottom deck of the blastholes initially. The fired bottom portion worked as the free face for the subsequent round of blasting. The optimum delay timing of 20–80 ms was used between the decks in this study. This delay interval worked successfully in optimizing the explosive energy utilization after deck. In this study, the comparison was made for two conditions of multiple decked ring blasting. The first condition encompasses the complete firing of all the holes at once and thereafter the firing of the decks of subsequent ring holes. In the second condition, the ring holes were divided as bottom slice and top slice. The bottom slice was fired in the first round and thereafter the firing of top slice was accomplished. The blasting sequence for these two conditions is shown in Fig. 7.1. The blasting output revealed that the fragmentation under second condition was better as compared to the first condition. Accordingly, the bottom slice–top slice technique is helpful for reducing the ground vibration near surface structures for ring blasting along with maintaining the desired fragmentation output.

7.4.2.2

Case Study-2: Designing of Controlled Blasting Pattern for Reducing Vibration Near Surface Industrial Structures

Himanshu et al. (2018b) designed the blasting parameters for blasting faces located near surface industrial structures of Sindesar Khurd underground mine. The structures in this study included crusher plant, water tank, tailing thickener, mill process

100

7 Blast-Induced Hazards

Fig. 7.1 Detonation of the ring blastholes by detonating holes one by one and by taking bottom slice first and top slice afterwards (Roy et al. 2016)

plant, control room, electrical substation, mine office, etc. The experimentation with variation in blast design was performed for this purpose. The experimental and simulation approach was used to optimize blast design parameters. The blast vibration predictors were developed based on the gathered data. The multi-deck blasting technique was used in this study to distribute the explosive charge. A view of multi-deck blasting pattern used in this study is shown in Fig. 7.2. In this study, the influence of total explosive charge on induced ground vibration was also investigated. It is presumed that the induced ground vibration from different delays may superimpose at far-field distances. Accordingly, the multivariate regression analysis was carried out correlating PPV with different blast design parameters. The predictor from the multivariate regression analysis was power in nature. It included the total explosive charge in a blasting round, number of blastholes, hole diameter and distance of blast face from monitoring point. The developed multivariate predictor under this study is given in Eq. 7.1. This equation is site specific. The total explosive charge and number of blastholes were restricted based on this equation. The technique developed under this study resulted in reducing the ground vibration near industrial structures within the limits stipulated by DGMS. PPV = 367

ϕ 0.26 Qt 0.45 N 0.45 D 1.26

(7.1)

7.4 Optimisation of Ring Blasting Parameters for Structural Safety

101

Fig. 7.2 Multi-deck blasting pattern for underground ring

Multiple coefficient of correlation = 81.78%, where ϕ = Hole diameter (mm) Qt = Total explosive charge in a blasting round (kg) N = Number of blastholes fired in a blasting round D = Distance of vibration monitoring station from blast face (m) PPV = Peak particle velocity of vibration (mm/s) 7.4.2.3

Case Study-3: Designing of Controlled Blasting Pattern for Reducing Vibration Near Underground Structures

The nearby underground structures, viz. shaft, decline, haulage drivages, underground workshops, etc., are also affected by induced ground vibrations. Under Indian conditions, the vibration near these underground structures is to be controlled within the limits shown in Table 7.3. Sometimes some of these underground structures are within 30 m distance from the blast face. The advanced statistical techniques are used in such cases to investigate the influence of different blast design parameters. One such study was carried out by Roy et al. (2022) at an Indian underground mine. In this study, the experimentations were done by variation of blast design parameters.

102

7 Blast-Induced Hazards

The PPV data was recorded along with the design parameters. The statistical analysis of recorded charging parameters for the blast was performed. The recorded data consisted of various charging parameters for the blast, distance of vibration monitoring station from the blast face (D) and recorded PPV. The charging parameters included—MCPD, total explosive charge in a ring (TCR) and total explosive charge in a blasting round (TC). The importance of different independent parameters, viz. MCPD, TCR, TC and distance (D) on dependent parameter (PPV), was estimated using artificial neural network approach. Under this study, different neural network models were developed to analyse the importance of various charging parameters and distance on the induced ground vibration from the ring blasting. The prepared ANN model was trained and tested with 70% and 30% of the data, respectively. The model performance was evaluated as coefficient of determination (R). The value of R was taken more than 0.8 in all the prediction models. Initially, the analysis has been carried out to predict the output PPV from the input parameters D, MCPD, TCR and TC. The analysis was carried out for all the recorded data. It was found that PPV shows maximum dependency on D. The order of dependency of PPV on different parameters is as—D > MCPD > TCR > TC. The result shows that the total explosive charge in a ring also has significant influence (normalized importance—47.2%) on PPV. This is due to the superposition of blast waves of two holes of a ring. To further identify the superposition characteristics of the blast wave, ANN model was prepared for PPV at different distances from the blast face. Seven different models with PPV recorded at distances of > 100 m, < 100 m, < 90 m, < 80 m, < 70 m, < 60 m and < 50 m were analysed. The analysis shows that distance is the dominant parameter influencing PPV, when recording has been done at a distance of > 60 m. But, when the distance of recording ground vibration is less than 60 m, then D and MCPD have equal influence on PPV. The role of other charging parameters also starts dominating in such case, which defines the larger impact of charging parameters than distance on PPV at nearfield locations. MCPD is the dominant charging parameter influencing PPV. However, it was observed that the normalized importance of TCR and MCPD are almost equal when recording of the ground vibration was made at a distance of > 90 m. This analysis suggests that the superposition of the blast waves of two holes of a ring has started at a distance of more than 90 m from the blast face. Since the data was measured only at nearfield distances, the superposition of the blast waves of two rings could not be found in the analysis. However, it may be interpreted from the increasing trend of importance of TC with distance that there will be impact of TC at far-field distances. The delay intervals between the rings were kept more than 500 ms during the experimentation. Accordingly, the multi-ring blast design as shown in Fig. 6.9 was adopted for productive exploitation of mineral at the study site along with controlling the ground vibration near underground structures within safe limits. With the help of the multi-ring blast design suggested using this study, the mine management has been able to conduct blast with total explosive charge of 4920 kg. The explosive charge was distributed in six different rings. The blast face was at a distance of 40 m from nearest footwall drivage. The blast was conducted successfully with the magnitude of vibration recorded near footwall drivage as 50.51 mm/s at peak

7.6 Summary

103

dominant frequency of 79.88 Hz. The recorded vibration data was well within the statutory limit as per DGMS standard.

7.5 Safety Precautions While Blasting To achieve the safe blasting at the faces in the nearby of the surface and underground structures, the following safety precautions need to be undertaken by the mine management: • Maximum charge weight per delay and total explosive charge in a blasting round should be followed considering the minimization of blast vibration within stipulated standards for safety of nearfield underground structures, far-field underground structures as well as surface residential/industrial structures. • Hole deviation measurement should be done to ensure proper toe burden as well as collapsing of two holes. As the collapsing of holes may lead to increased charge weight per delay. • Special precautions at permanent underground structures, viz. drivages, shaft pillars, crown pillar, etc., should be taken to ensure proper support after blasting. The insitu stresses are redistributed after blasting, which may lead to requirement of additional supports at these places to regain rock strength. • Loose rocks should be dressed properly before drilling/charging behind blasted face. • Stope should be scanned after blasting to get actual view of void generated after blasting. Precautions should be taken while charging where additional cavity in toe of downhole face has been observed in scan.

7.6 Summary The ground vibration induced by large-scale production blasting for underground mining works as safety and environmental threat to the surface and underground structures. The detailed discussion regarding the measurement and prediction of ground vibration has been made in this chapter. The discussions of this chapter are summarized below: I.

In Indian condition, the safe limits for the ground vibration have been framed under Directorate General of Mines Safety (DGMS) regulations. The DGMS circular 7 of 1997 defined the limits of ground vibration for the safety of surface structures. II. The safe vibration limits for underground structures, viz. underground haulage drivages, shaft, decline, workshop, electrical substations, repair shops, etc., may be taken as per Directorate General of Mines Safety (DGMS) Tech circular No. 06 of 2007. This circular describes the vibration limits around different

104

7 Blast-Induced Hazards

structures based on rock mass rating (RMR), which is a good indicator of insitu rock condition. III. Different controlled blasting pattern has been developed for underground ring blasting in order to meet the production demand along with maintaining the ground vibration within safe limits. Roy et al. (2016) use a pattern with slicing of ring into two parts. The bottom slice of the ring was given initial delay and thereafter the top slicing portion was taken. The method has resulted into the successful reduction of ground vibration near domestic structures. IV. Himashu et al. (2018b) used multi-deck blasting pattern for the reduction of ground vibration near surface industrial structures. The influence of total explosive charge on induced ground vibration was also investigated under this study. V. The multi-ring blasting technology with the use of long delay between rings was used for productive blasting with the safety of nearby underground structures. The blast design for this purpose was optimized by investigating the parameters influencing PPV.

References Ambraseys NR, Hendron AJ (1968) Dynamic behaviour of rock masses. In: Rock mechanics in engineering practices. Wiley Bahrami A, Monjezi M, Goshtasbi K, Ghazvinian A (2011) Prediction of rock fragmentation due to blasting using artificial neural network. Eng Comput 27(2):177–181. https://doi.org/10.1007/ s00366-010-0187-5 Bieniawski ZT (1968) Fracture dynamics of rock. Int J Fract Mech 4(4). https://doi.org/10.1007/ BF00186807 DGMS Circular No. 06 (2007) DGMS circular No. 07 (1997) Dowding CH (1985) Blast vibration monitoring and control. Prentice-Hall Inc., New Jersey Duvall WI, Fogelson DE (1962) Review of criteria for estimating damage to residences from blasting vibrations. Report of Investigations/United States Department of the Interior, Bureau of Mines 5968, p 19, //catalog.hathitrust.org/Record/005981586%5Cn. http://hdl.handle.net/2027/mdp. 39015078529347 Gorai AK, Himanshu VK, Santi C (2021) Development of ANN-based universal predictor for prediction of blast-induced vibration indicators and its performance comparison with existing empirical models. Mining, Metall Explorat. https://doi.org/10.1007/s42461-021-00449-0 Gui YL, Zhao ZY, Zhou HY, Goh ATC, Jayasinghe LB (2017) Numerical simulation of rock blasting induced free field vibration. Procedia Eng 191:451–457. https://doi.org/10.1016/j.proeng.2017. 05.203 Hao H, Wu Y, Ma G, Zhou Y (2001) Characteristics of surface ground motions induced by blasts in jointed rock mass. Soil Dyn Earthq Eng 21(2):85–98. https://doi.org/10.1016/S0267-726 1(00)00104-4 Hasanipanah M, Monjezi M, Shahnazar A, Jahed Armaghani D, Farazmand A (2015) Feasibility of indirect determination of blast induced ground vibration based on support vector machine. Measurement 75:289–297. https://doi.org/10.1016/j.measurement.2015.07.019

References

105

Himanshu VK, Roy MP, Mishra AK, Paswan RK, Panda D, Singh PK (2018a) Multivariate statistical analysis approach for prediction of blast-induced ground vibration. Arab J Geosci 11(16):460. https://doi.org/10.1007/s12517-018-3796-8 Himanshu VK, Roy MP, Sarim Md, Singh PK, Mishra AK (2018b) Blast design for underground long hole stope to safeguard surface industrial structures from blast induced vibration. In: 12th international symposium on rock fragmentation by blasting, Lulea, Sweden, pp 257–267 Kumar R, Choudhury D, Bhargava K (2016) Determination of blast-induced ground vibration equations for rocks using mechanical and geological properties. J Rock Mech Geotech Eng 8(3):341–349. https://doi.org/10.1016/j.jrmge.2015.10.009 Monjezi M, Ghafurikalajahi M, Bahrami A (2011) Prediction of blast-induced ground vibration using artificial neural networks. Tunn Undergr Space Technol 26(1):46–50. https://doi.org/10. 1016/j.tust.2010.05.002 Pal Roy P (1991) Vibration control in an opencast mine based on improved blast vibration predictors. Mining Sci Technol 12(2):157–165. https://doi.org/10.1016/0167-9031(91)91642-U Resende R, Lamas L, Lemos J, Calçada R (2014) Stress wave propagation test and numerical modelling of an underground complex. Int J Rock Mech Min Sci 72:26–36. https://doi.org/10. 1016/j.ijrmms.2014.08.010 Roy MP, Singh PK, Sarim M, Shekhawat LS (2016) Blast design and vibration control at an underground metal mine for the safety of surface structures. Int J Rock Mech Min Sci 83:107– 115. https://doi.org/10.1016/j.ijrmms.2016.01.003 Roy MP, Himanshu VK, Kaushik AP, Singh PK (2022) Influence of ring blasting pattern on the safety of nearby underground structures. S¯adhan¯a 47(4):192. https://doi.org/10.1007/s12046022-01968-2 Singh PK, Roy MP (2010) Damage to surface structures due to blast vibration. Int J Rock Mech Min Sci 47(6):949–961. https://doi.org/10.1016/j.ijrmms.2010.06.010 Singh PK, Roy MP, Sinha A (2008) Damage of structures due to blast induced vibrations in mining areas. 2nd Asian Mining Congress, Kolkata India, pp 385–397 Singh PK, Roy MP, Paswan RK, Dubey RK, Drebenstedt C (2015) Blast vibration effects in an underground mine caused by open-pit mining. Int J Rock Mech Min Sci 80:79–88. https://doi. org/10.1016/j.ijrmms.2015.09.009 Siskind DE, Stachura VJ, Stagg MS, Koop JW (1980) Structure response and damage produced by air-blast from surface mining. United States Bureau of Mines Tian E, Zhang J, Soltani Tehrani M, Surendar A, Ibatova AZ (2019) Development of GA-based models for simulating the ground vibration in mine blasting. Eng Comput 35(3):849–855. https:// doi.org/10.1007/s00366-018-0635-1 Verma AK, Singh TN (2011) Intelligent systems for ground vibration measurement: a comparative study. Eng Comput 27(3):225–233. https://doi.org/10.1007/s00366-010-0193-7 Wen S, Zhang C, Chang Y, Hu P (2020) Dynamic compression characteristics of layered rock mass of significant strength changes in adjacent layers. J Rock Mech Geotech Eng 12(2):353–365. https://doi.org/10.1016/j.jrmge.2019.09.003 Wu YK, Hao H, Zhou YX, Chong K (1998) Propagation characteristics of blast-induced shock waves in a jointed rock mass. Soil Dyn Earthq Eng 17(6):407–412. https://doi.org/10.1016/ S0267-7261(98)00030-X Xie LX, Lu WB, Zhang QB, Jiang QH, Wang GH, Zhao J (2016) Damage evolution mechanisms of rock in deep tunnels induced by cut blasting. Tunn Undergr Space Technol 58:257–270. https:// doi.org/10.1016/j.tust.2016.06.004 Yang JH, Yao C, Jiang QH, Lu WB, Jiang SH (2017) 2D numerical analysis of rock damage induced by dynamic in-situ stress redistribution and blast loading in underground blasting excavation. Tunn Undergr Space Technol 70:221–232. https://doi.org/10.1016/j.tust.2017.08.007 Zhou J, Li C, Koopialipoor M, Jahed Armaghani D, Thai Pham B (2021) Development of a new methodology for estimating the amount of PPV in surface mines based on prediction and probabilistic models (GEP-MC). Int J Min Reclam Environ 35(1):48–68. https://doi.org/10.1080/174 80930.2020.1734151

Chapter 8

Innovative Blasting Practices for Underground Hard Rock Mining

Abstract The state-of-the-art blasting methodologies have been developed over the years to overcome the challenges of underground mining. The major challenges associated with Indian underground mines include ore dilution and narrow vein mining. The productivity enhancement from underground is another challenge, as most of the shallow depth ore deposits are in the stage of extinction. To achieve the production pace from underground, the mining with blasting of larger number of holes in a same blasting round was the need. One special technique has been evolved for this purpose. This technique is capable of extracting 60–70% of the stope in a single blasting round. The detailed discussion regarding this technique has been made in this chapter. Additionally, the blasting methods for narrow vein mining and excavation in ore-waste combination have also been discussed in this chapter.

8.1 Introduction Production enhancement and improving grade of the excavated mineral are two main goals of a mining engineer. Optimization of blasting pattern has significant contribution in achieving these goals. The production from underground blasting can be enhanced using the blasting of larger diameter and deeper blastholes. Increment in number of blastholes and multi-ring blasting also adds to the production enhancement. The technological advancements, viz. fast charging technology using site mixed explosives, high viscosity explosives for uphole charging and electronic delay detonators for judicious decision on delay timings supports the miners to achieve faster production targets. Various advanced techniques of ring blasting are used at different underground mines like Malmberget Iron Ore mine, Kiruna Iron Ore mine, Fankou Lead–Zinc mine and Meishan Iron Ore mine (Sun et al. 2021; Wang et al. 2018; Zhang and Wimmer 2018; Zhang 2014; Zhang and Naarttijärvi 2005; Wimmer et al. 2012). Some of these techniques may be implemented with the site-specific changes at other similar mines in the global context. Another important challenge while ring blasting is to control ore dilution. Different innovative techniques are used to reduce ore dilution during blasting. The

© The Author(s), under exclusive license to Springer Nature Singapore Pte Ltd. 2023 V. K. Himanshu et al., Blasting Technology for Underground Hard Rock Mining, https://doi.org/10.1007/978-981-99-2645-9_8

107

108

8 Innovative Blasting Practices for Underground Hard Rock Mining

research articles suggest that the optimum stope blasting pattern leads to minimization of the ore dilution. Such pattern incorporates investigation of rock–explosive interaction to contain the blast-induced damages within the orebody. Various controllable and uncontrollable parameters have influence on the blast-induced rock mass damages. The investigations of uncontrollable parameters, viz. characteristics of ore body, rock mass properties, insitu stresses, etc., can give an insight to optimize the blast design parameters to contain the damages. Researchers have used different techniques for the prediction of ore dilution due to blasting. The factors responsible for dilution have also been identified in these studies. Clark (1998) has used the concept of Equivalent Linear Overbreak Slough (ELOS) for the assessment of unplanned dilution. Using this ELOS parameter, the irregular overcut orebody under exploitation is converted into average overcut depth. Henning and Mitri (2008) have discussed different methods for assessment and control of ore dilution. Stewart et al. (2007) discussed the methods for stoping in narrow veins to reduce dilution. Authors have identified three different patterns for drilling, viz. inline, dice 5 and staggered. The concept of overbreak estimation for slot raise and drivage excavation used by the researchers can also be implemented to reduce the ore dilution. Researchers have used various soft computing and numerical modelling-based tools for the assessment of rock mass damage due to blasting. Zhao and Niu (2020) have predicted unplanned ore dilution using artificial neural network. Jang et al. (2015) used neurofuzzy system for prediction of ore dilution. Some of the innovative techniques used at Indian underground mines to improve the production pace and reduce the ore dilution have been discussed in this chapter.

8.2 Technique for Simultaneous Excavation of Slot Raise and Rings in a Single Blasting Round The excavation of deeper blastholes in a single blasting round has been adopted in many mines to enhance the pace of production. The depth of the blastholes is kept as equal to the level interval in such cases. However, the level interval is decided on the basis of the dip of the ore body, to minimize the cost of the production. So, the depth of the blastholes in such cases would be limited. Accordingly, the enhancement in production pace in such a case may be achieved by increasing the number of blastholes to be fired in a blasting round. A special technique has been devised to excavate the slot raise and rings of the hangingwall side simultaneously for this case. This technique was first implemented at Rampura Agucha underground Lead–Zinc mine. The available drill hole diameter at the mine for stope excavation was 89 and 152 mm. The ore body was having a dip of 55º and width of 18–20 m. The level difference of the stope was 25 m centre to centre. The major geological constraints at the stope were plasticity of the rock strata. The plastic nature of the rock strata leads to the restricted movements of the blasted muck while excavation of slot raise. The

8.2 Technique for Simultaneous Excavation of Slot Raise and Rings …

109

Fig. 8.1 Schematic layout showing the excavation sequence of an underground stope

proper excavation of slot raise is necessary to ensure the successful excavation of subsequent ring holes. The major environmental concern at the site was to safeguard the nearby sensitive underground structures, viz. footwall drive, decline, electrical panel, etc. The operational constraint of the stope included the lack of space for accumulation of blasted rock from the slot raise and stope excavation. Initial positioning of the holes was planned on the basis of excavation sequence for the whole stope. The vertical slot holes were planned between two levels of the stoping block. The rest part of the orebody was planned to be excavated by drilling ring holes along hangingwall as well as footwall direction. The required space for accumulation of blasted muckpile of the stope was computed considering the bulking factor of the rock. Accordingly, there was need for excavation of 6 m height of the stope. The excavation of this 6 m height of the stope was planned using uphole drilling and charging. The excavation was termed as sliping. The slot raise and rings along the hangingwall direction were planned to be excavated in a single round. The rings along the footwall direction were planned to be excavated in another blasting round. The schematic layout of the excavation sequence of stope is shown in Fig. 8.1. The drilling parameters for the experimental stope were decided on the basis of analysis of deformation pattern using numerical simulation (Himanshu et al. 2021). The core samples of the rock strata of the stope were collected, and the assessment of rock mass properties was carried out. The assessed rock mass properties were used as the input in the numerical model. The explosive properties were also simulated in the model based on the properties of explosives available at the mine. The complete extraction of slot in a round was simulated in the initial stage. The deformation pattern was not satisfactory as the extent of slot along strike direction was very large. Accordingly, the slot was divided as slot and slot expansion. The different combinations of reamer and blastholes were simulated in the numerical model. The optimum drill design for the slot was finalized to maximize the deformation. The slot raise expansion holes were planned to move along the excavated slot raise. The numerical model with excavated slot raise and drilled slot expansion holes was prepared. The deformation pattern for different combinations of burden–-spacing for slot expansion holes was analysed. The optimum burden–spacing for slot expansion

110

8 Innovative Blasting Practices for Underground Hard Rock Mining

holes was decided based on the model output. The excavation of rings was planned against the excavated slot raise. The burden–spacing for excavation of rings was optimized on the basis of deformation contour analysis in the numerical model. The schematic excavation sequence for slot raise and rings is shown in Fig. 8.2. The charging quantification for blastholes of slot raise and rings was done using statistical analysis. The United State Bureau of Mines (USBM) blast vibration predictor was developed to predict the magnitude of vibration under varying Maximum Charge Weight per Delay (MCPD). The suitable MCPD was decided on

Fig. 8.2 Schematic layout showing the face movement direction while simultaneous excavation of slot raise and rings

8.3 Drilling and Blasting Methodology for Extraction of Narrow Vein Ore …

111

the basis of the distance of nearby sensitive structure. The safe limits of vibration near underground structure were taken on the basis of Directorate General of Mines Safety (DGMS) circular no. 6 of 2007. Multi-variate statistical regression analysis was carried out to study the impact of total explosive charge in slot raise/single ring/blasting round. It was observed that the complete dampening of blast vibration occurs when the slot raise and different rings are separated from each other by long delay of more than 500 ms. The delay sequence of slot raise holes was decided by analysis of deformation contour from numerical model. The slot raise holes were distributed as cut holes, first box holes and periphery holes based on the model output. The holes of first box and periphery positions were scattered to maximize the deformation. The delay timings of slot raise holes and ring holes were decided on the basis of waveform analysis and signature hole analysis. The selection of optimum delay time for slot holes was done to ensure tensile breakage of the rock mass. The cumulative face movement and void generation in different rounds of blasting were taken into consideration while selection of delay timings. The delay timing for slot raise holes from a stope blasting is shown in Fig. 8.3. The delay timings for ring holes were decided to dampen the impact of total explosive charge from multiple rings on blast vibration. The interaction between the blasted rock was also taken into consideration to maximize rock fragmentation. The delay timings for slot raise expansion and ring holes for the excavation of same stope are shown in Fig. 8.4. The rings along footwall direction are further excavated in the next blasting round. Accordingly, the complete extraction of stope using this method is accomplished in three blasting rounds. During the simultaneous excavation of slot raise and rings of hangingwall direction, about 60– 70% of the stope is excavated in a single blasting round. The major advantages of this blasting methodology are as follows: (i)

It accelerates the pace of mineral exploitation, as about 60% of the stope is excavated in a single blast round. (ii) It reduces the cost of mineral production, as this method reduces the cycles of the unit operation in stope excavation and thereby saves the cost of excavator shifting, explosive and accessories transport, manpower cost, etc. (iii) It reduces the safety risk to the miners by abolishing their multiple time exposure to the blasted face.

8.3 Drilling and Blasting Methodology for Extraction of Narrow Vein Ore Deposits The excavation of narrow vein ore deposits by drilling and blasting is a challenging task. The extent of damage while blasting in such deposit leads beyond the ore deposits. Accordingly, the wall rock gets damaged and induces dilution. Different methods have been used worldwide to excavate the narrow vein ore deposits. Stewart et al. (2007) suggested three different methods for excavation of narrow vein ore

112

8 Innovative Blasting Practices for Underground Hard Rock Mining

Fig. 8.3 Delay timings of the blastholes for excavation of slot in a stope

deposits using longhole stoping method. The suggested methods are—staggered/zigzag, dice 5 and inline method. In the staggered method, the blastholes are placed alternatively closer to the hangingwall and footwall. In dice 5 pattern, two holes are placed closer to the hangingwall and two holes are placed closer to the footwall, the remaining one hole is placed in the middle of the ore body. The pattern is repeated afterwards. In inline pattern, the blastholes are placed in the middle of the orebody. The schematic of these patterns is shown in Fig. 8.5. The blasting methodology for excavation of narrow vein ore deposits of Zawar group of mines was devised using the concept of Inline pattern. The devised methodology consisted of drilling of blastholes of diameter 64–76 mm keeping toe burden of 1.5–2.5 m. The optimum toe burden for different faces was decided on the basis of the rock mass properties of ore body and host rock. The blastholes were drilled in a single line along the strike of the ore deposit in this method. In most of the cases, the blastholes were kept in the middle of the orebody, between the hangingwall and footwall. The hangingwall is more prone to damage during the blasting. Accordingly, the blastholes were shifted towards footwall, when the width of the orebody was larger than 3 times of the toe burden. A view of the positioning of blastholes for this method is shown in Fig. 8.6.

8.3 Drilling and Blasting Methodology for Extraction of Narrow Vein Ore …

113

Fig. 8.4 Delay timings of blastholes for excavation of slot expansion and rings of hangingwall direction of a stope

Fig. 8.5 Schematic of drilling patterns for excavation of narrow vein ore deposits (Stewart et al. 2007)

The maximum explosive charge weight per delay was computed to reduce the vibration in the wall rock within the critical limit to initiate damage. The explosive charge was distributed in different decks to reduce MCPD. The suitable plugging arrangement for individual decks was done. Differential charging of the blastholes was used to utilize the combined impact of explosive energy on rock breakage. The blastholes were charged preferably by low-density ANFO or SME explosive to reduce the effective damage by shock wave.

114

8 Innovative Blasting Practices for Underground Hard Rock Mining

Fig. 8.6 Positioning of blastholes for excavation of narrow vein deposits

The sequence of blastholes was planned such that first hole was blasted along the slot raise, and the consecutive movement of blastholes has taken place along the strike of the orebody. In a particular blasthole, the bottom deck of the hole was taken first and top deck at the last in the sequence. The initiation of bottom deck of the second hole started after completion of blast of middle deck and before blast of top deck of first hole. The sequence of firing of blastholes in this methodology is shown in Fig. 8.7. The initiation of the blastholes was carried out using Electronic/NONEL detonators with suitable delay timing. The explosive decks were separated by delay interval of 20–80 ms.

8.4 Multi-Ring Blasting Pattern for Rock Excavation in Ore-Waste Combination The rock blasting in ore-waste combination need to be dealt judiciously. The major challenge in this case includes the proper breakage of the rock strata along with the reduction of the ore dilution. The conventional ring blasting practices in such case leads to improper breakage. Accordingly, a method was devised which included the differentiation of explosive charging pattern even in the rings. The differential charging pattern between different holes of a ring (as shown in Fig. 6.1) is popular for such cases. But, sometimes during the multi-ring blasting, the charge concentration at a particular zone also gets enhanced while using uniform differentiation between

8.4 Multi-Ring Blasting Pattern for Rock Excavation in Ore-Waste …

115

Fig. 8.7 Differential charging pattern and sequence of firing of blastholes in narrow vein excavation

the holes of different rings. So, the differential charging in multi-ring blasting pattern was used at one of the underground mine. In this pattern, the first blasthole of the first ring was charged upto two-third length, and rest portion was stemmed. The adjacent blastholes of the first hole were charged upto one-third length and stemmed upto two-third length of the blasthole. The pattern was repeated for all the holes of the first ring. First and second blastholes of the first ring were taken together as initiation holes in this pattern to enhance the breakage of the ore body. The blasthole of the second ring behind the first hole of the first ring was charged upto one-third length and stemmed upto two-third length. Afterwards, the pattern was repeated for the second ring similar to that of the first ring. Again the charging pattern for third ring was a copy of the charging pattern of the first ring. Since the third ring would be closer to the wall rock, the initiation of single hole at a time was used for this ring. The firing sequence and differential charging pattern used in this blasting methodology are shown in Fig. 8.8. The delay timings between the holes and between rings were optimized on the basis of waveform analysis to reduce ore dilution. The delay timing between second and third ring was kept more as compared to that between first and second ring in order to provide additional relief to the ring before blasting. The rock excavation with the reduced dilution and enhanced throw was achieved using this blasting methodology.

116

8 Innovative Blasting Practices for Underground Hard Rock Mining

Fig. 8.8 Charging pattern and firing sequence for multi-ring differential blasting pattern

8.5 Summary Three innovative blasting practices have been discussed in this chapter. These blasting techniques have been devised to address different problems of the mining industry. The summarized details of these methods are as follows: I.

The technique for simultaneous excavation of slot raise and rings is useful in enhancing the pace of production from underground. In this technique, the blasting of slot raise, slot raise expansion and rings along hangingwall contact is conducted in a single round using electronic delay detonators. The technique has advantage of saving cost of production by minimizing the machinery movement and manpower deployment. II. Various blasting methodologies are used for excavation of narrow vein ore deposits while minimizing the ore dilution. The inline drilling pattern along with differential charging of blastholes is a useful method for reducing ore dilution. III. The multi-ring differential blasting pattern may be used to minimize dilution and maximize ore breakage while blasting in ore-waste combination. The pattern consisted of the differential charging of ring blastholes of different rings.

References

117

References Clark LM (1998) Minimizing dilution in open stope mining with a focus on stope design and narrow vein longhole blasting. University of British Columbia Henning JG, Mitri HS (2008) Assessment and control of ore dilution in long hole mining: case studies. Geotech Geol Eng 26(4):349–366. https://doi.org/10.1007/s10706-008-9172-9 Himanshu VK, Mishra AK, Roy MP, Vishwakarma AK, Singh PK (2021) Numerical simulation based approach for assessment of blast induced deformation pattern in slot raise excavation. Int J Rock Mech Min Sci 144:104816. https://doi.org/10.1016/j.ijrmms.2021.104816 Jang H, Topal E, Kawamura Y (2015) Decision support system of unplanned dilution and ore-loss in underground stoping operations using a neuro-fuzzy system. Appl Soft Comput 32:1–12. https://doi.org/10.1016/j.asoc.2015.03.043 Stewart P, Trueman R, Lyman G (2007) Development of benchmark stoping widths for longhole narrow-vein stoping. Min Technol 116(4):167–175. https://doi.org/10.1179/174328608 X318270 Sun M, Ren F, Ding H (2021) Optimization of stope structure parameters based on the mined orebody at the Meishan iron mine. Adv Civil Eng 2021:1–14. https://doi.org/10.1155/2021/805 2827 Wang M, Shi X, Zhou J (2018) Charge design scheme optimization for ring blasting based on the developed Scaled Heelan model. Int J Rock Mech Min Sci 110:199–209. https://doi.org/10. 1016/j.ijrmms.2018.08.004 Wimmer M, Nordqvist A, Ouchterlony F, Nyberg U, Furtney JK (2012) Burden movement in confined drift wall blasting tests studied at the LKAB Kiruna SLC mine. In: The 10th international symposium on rock fragmentation by blasting, pp 373–383 Zhang ZX (2014) Effect of double-primer placement on rock fracture and ore recovery. Int J Rock Mech Min Sci 71:208–216. https://doi.org/10.1016/j.ijrmms.2014.03.020 Zhang ZX, Naarttijärvi T (2005) Reducing ground vibrations caused by underground blasts in LKAB Malmberget mine. Fragblast 9(2):61–78. https://doi.org/10.1080/13855140500140275 Zhang Z-X, Wimmer M (2018) A case study of dividing a single blast into two parts in sublevel caving. Int J Rock Mech Min Sci 104:84–93. https://doi.org/10.1016/j.ijrmms.2018.02.002 Zhao X, Niu J (2020) Method of predicting ore dilution based on a neural network and its application. Sustainability 12(4):1550. https://doi.org/10.3390/su12041550

Chapter 9

Challenges and Way Forward

Abstract Underground blasting for hard rock mining consists of three main stages as drivage development, slot raise excavation and production ring excavation. The nature of challenges in these stages is different. These challenges are dealt with the technological advancements in operations, mechanization and instrumentation. The advent of advanced prediction techniques has also provided ease to the blast designers. However, the underground mining at greater depth of cover will come with the additional opportunities for the technological advancements and scientific solutions. The developments of low VOD explosives, plasma blasting techniques or chemical-based rock breakage technology are some of the solutions for the future challenges. The advanced assessments of rock mass properties and thereby accurate prediction of blasting outcomes is also necessary to achieve safe and productive underground excavation. This chapter presents some of the future challenges and opportunities in underground blasting.

9.1 Overview Underground mining for exploitation of mineral deposits are becoming dominant with the extinction of shallow depth ore deposits. Maintaining the production pace from underground is a major challenge to the mining practitioners. The longhole stoping with large-scale drilling and blasting can be a suitable alternative to overcome these challenges. The planning and implementation of mass blasting for longhole stoping require thorough scientific investigations of the expected outcomes. Such investigations need to be supported by the state of the art instrumentation, advanced predictive models and adequate case histories. This book presented the technological advancements in underground blasting in India over the years. The mining operation using longhole stoping method comprises of three main steps as—drivage development, slot raise excavation and ring blasting. The challenges in these steps are different. The enhancement of pull and reduction of overbreak is the main aim of a blast designer in drivage blasting (Himanshu et al. 2022). The optimum blast design to achieve these objectives is specific to the rock mass condition of the site. Accordingly, the assessment of rock mass condition is done © The Author(s), under exclusive license to Springer Nature Singapore Pte Ltd. 2023 V. K. Himanshu et al., Blasting Technology for Underground Hard Rock Mining, https://doi.org/10.1007/978-981-99-2645-9_9

119

120

9 Challenges and Way Forward

to devise the blasting pattern for drivage blasting. The cut portion of the burn-cut face blasting design, is mainly responsible for the pull from the drivage blasting. The optimum number and diameter of blastholes, relief holes are selected to maximize the pull (Murthy and Dey 2002; Verma et al. 2018). Additionally, the delay timings between the holes are also chosen judiciously to optimize the damage in the cut blasting round. The objective of overbreak reduction is achieved by optimizing the resultant detonation pressure along the periphery holes. This is done using the waveform analysis of the nearfield vibration (Mandal and Singh 2009; Vishwakarma et al. 2020; Himanshu et al. 2022). The slot raise blasting pattern is somehow similar to the burn-cut face blasting pattern, as the breakage in both the cases takes place along the tensile relief provided by reamer holes (Himanshu et al. 2021a). However, the slot raise pattern shows dissimilarity in the direction of movement of blasted rock. Accordingly, the suitable placement of initiation system is important in such case. The slot raise blasting comes with the challenges of muckpile movement, face jamming, improper breakage, etc. The ratio of reamer hole to the blastholes is optimized to get optimum breakage in this case. Additionally, the delay timing and sequence of firing of blastholes are also optimized based on the waveform analysis (Himanshu et al. 2021a). The production blasting in longhole stoping method includes the drilling of holes in ring pattern. The ring blasting takes place against the free face created by the excavated slot raise. The optimum burden and spacing are selected for the ring blasting to get the proper fragmentation. Uphole and downhole charging are used in the ring blasting based on the need. The ring blasting operation has the challenges to optimize the rock fragmentation and reduce the induced ground vibration near surface and underground structures (Himanshu et al. 2021b; Roy et al. 2022). The drilling parameter, charging pattern and delay timings are suitably chosen to enhance the rock breakage with reduced ground vibration. The specialized pattern in ring blasting has been devised to enhance the pace of production. Such pattern includes the complete extraction of slot raise and rings of hangingwall direction in a single blasting round. The ring blasting pattern for narrow vein mining and controlling ore dilution has also been devised over the years.

9.2 Future Challenges of Underground Blasting and Way Forward The discussions made in this book are based on the technological advancements adopted in Indian underground mine for exploitation of minerals. The scientific approaches used for devising the blasting methods have been discussed in different chapters. However, with the increasing challenges of mining specifically with the increment in the depth of mining and extinction of easily mineable ores, there would be requirement of more sophisticated techniques for prediction of blasting outcomes and implementation of advanced blasting technology in future. Such predictions

9.2 Future Challenges of Underground Blasting and Way Forward

121

would be required to be backed up by advanced instrumentations, sophisticated prediction models and practical rock testing data. Some of the specific challenges in these areas have been discussed in this section. Rock mass properties are the basis of prediction and designing of the blasting parameters for a site. The accurate assessment of the rock mass properties is a prominent need for blast design. Under most of the situations, the rock mass properties are assessed to carry out static numerical simulation. The additional properties, viz. strain rate dependency, pressure dependency, etc., need to be assessed for the dynamic numerical simulation, and thereby for the prediction of blasting outputs. Split Hopkinson Pressure (SHPB) can serve this purpose. The compressive and tensile behaviour of rock mass can be tested under different strain rate loading conditions using SHPB (Gong et al. 2019). In most of the research work carried out for prediction of blasting outcomes using numerical simulation, RHT concrete and Johnson–Holmquist constitutive models were used (Borrvall and Riedel 2011; Wang et al. 2018; Baranowski et al. 2020). These models have been devised for predicting damages in homogeneous concrete media under dynamic loading. The rock, however, has different nature of heterogeneity due to the presence of joints, shear zone, faults, folds, etc. Different failure models for static loading conditions, viz. Hoek & Brown rock failure criterion, Sheorey rock failure criterion, etc., are available (Jaeger et al. 2014). But, the constitutive model for the prediction of rock failure under dynamic loading is not known till date. Accordingly, there is a need to develop the constitutive model comprising of insitu rock properties, viz. rock mass rating (RMR), rock quality designation (RQD) or Q system along with the dynamic rock mass properties. The rock mass to be excavated from the greater depth of cover would also be under the influence of insitu stresses. There is a need to design the blasting parameters considering the insitu stresses in such cases. Apart from the numerical model-based damage prediction, the empirical models also have important role in predicting the blasting outputs and thereby designing the blast for a practical problem. The production blasting specifically is designed to get optimum fragmentation. The charge factor and dimensional parameters for the production blasting are computed based on the available rock fragmentation predictors, viz. Kuz-Ram model, Swebrec function, etc. (Cunningham 1983; Kuznetsov 1973; Rosin and Rammler 1933; Kansake et al. 2016; Adebola et al. 2016). These prediction models, however, are for the openpit bench blasting. The underground ring blasting has different set of circumstances due to inclination of the hole. The rock fragmentation prediction model in such case would be different from the bench blasting. Accordingly, there is a need to develop a separate rock fragmentation model for underground blasting. In blasting operation, the challenges are there in controlling ore dilution, conducting blast in very close proximity to the structures and excavating very soft rocks using drilling and blasting. In future, the availability of high-grade mineral deposits would be limited. While extraction of lean-grade ores, the dilution of even smaller amount will lead to the economic loss to the mining industries (Jang et al. 2015; Zhou and Niu 2020). In such condition, the detailed guidelines on the blasting

122

9 Challenges and Way Forward

methodology for controlling ore dilution would be beneficial to the mining community. The rising demands of the raw material, will lead to a challenge to recover the ores lying in very close proximity to the structures. The development of nonexplosive-based rock breaking methodologies would be helpful in such cases. The available technologies such as Plasma Blasting, Autostem and rock breaking chemicals (Zhou et al. 2018) would be required to meet such challenges. The development of explosive with low detonation velocity would also be helpful in such cases. The advancements in the monitoring of blasting outputs will also be required in future. The existing methods of ground vibration monitoring are based on placement of seismograph at a particular point. However, with the increment in sensitivity, the need for monitoring at continuous points would be required. The accessibility of such monitoring results from surface will also be required. The Internet of Things (IoT) enabled monitors or fibre optic-based sensors would be helpful in such cases. The continuous fragmentation analysis based on images from the respective rounds of muckpile removal will also be required. The advanced fragmentation analysis system needs to be developed for this purpose. The analysis of the recorded monitoring data and identification of trend of their variation using machine learning algorithms is also important. The latest soft computing tools need to be adopted for the identification of the importance of controllable and uncontrollable parameters influencing blasting outputs.

9.3 Summary The challenges in underground blasting for mineral exploitation would increase with the extinction of easily mineable ore deposits. These challenges need to be overcome by adopting the technological advancements in instrumentation, numerical modelling, statistical analysis, etc. The operational developments also need to be made to cope up with the demands of the industry.

References Adebola JM, Ajayi OD, Elijah OP (2016) Rock fragmentation prediction using Kuz-Ram model. J Environ Earth Sci 06(05):110–115 Baranowski P, Mazurkiewicz Ł, Małachowski J, Pytlik M (2020) Experimental testing and numerical simulations of blast-induced fracture of dolomite rock. Meccanica 55(12):2337–2352. https:// doi.org/10.1007/s11012-020-01223-0 Borrvall T, Riedel W (2011) The RHT cconcrete model in LS-DYNA. In: 8th European LS-DYNA users conference Cunningham CVB (1983) The Kuz–Ram model for prediction of fragmentation from blasting. In: 1st international symposium on rock fragmentation by blasting, pp 439–453 Gong F-Q, Si X-F, Li X-B, Wang S-Y (2019) Dynamic triaxial compression tests on sandstone at high strain rates and low confining pressures with split Hopkinson pressure bar. Int J Rock Mech Min Sci 113:211–219. https://doi.org/10.1016/j.ijrmms.2018.12.005

References

123

Himanshu VK, Mishra AK, Roy MP, Vishwakarma AK, Singh PK (2021a) Numerical simulation based approach for assessment of blast induced deformation pattern in slot raise excavation. Int J Rock Mech Min Sci 144:104816. https://doi.org/10.1016/j.ijrmms.2021.104816 Himanshu VK, Roy MP, Shankar R, Mishra AK, Singh PK (2021b) Empirical approach based estimation of charge factor and dimensional parameters in underground blasting. Mining, Metall Explorat 38(2):1059–1069. https://doi.org/10.1007/s42461-020-00374-8 Himanshu VK, Mishra AK, Vishwakarma AK, Roy MP, Singh PK (2022) Explicit dynamics based numerical simulation approach for assessment of impact of relief hole on blast induced deformation pattern in an underground face blast. Geomech Geophys Geo-Energy and Geo-Resour 8(1):19. https://doi.org/10.1007/s40948-021-00327-5 Jaeger JC, Cook NGW, Zimmerman RW (2014) Fundamentals of rock mechanics (4th Edition). Wiley Blackwell Jang H, Topal E, Kawamura Y (2015) Decision support system of unplanned dilution and ore-loss in underground stoping operations using a neuro-fuzzy system. Appl Soft Comput 32:1–12 Kansake BA, Temeng VA, Afum BO (2016) Comparative analysis of rock fragmentation models—a case study. In: 4th UMT biennial international mining and mineral conference, pp 1–11 Kuznetsov VM (1973) The mean diameter of the fragments formed by blasting rock. Sov Min Sci 9(2):144–148. https://doi.org/10.1007/BF02506177 Mandal SK, Singh MM (2009) Evaluating extent and causes of overbreak in tunnels. Tunn Undergr Space Technol 24(1):22–36. https://doi.org/10.1016/j.tust.2008.01.007 Murthy VMSR, Dey K (2002) Development of predictive models for blast-induced rock damage assessment (BIRD) in tunnels. MHRD Project No: MHRD (27)/99-00/111/ME, pp 12–43 Rosin R, Rammler E (1933) Laws governing the fineness of powdered coal. J Instit Fuel 7:29–36 Roy MP, Himanshu VK, Kaushik AP, Singh PK (2022) Influence of ring blasting pattern on the safety of nearby underground structures. Sadhana 47:192. https://doi.org/10.1007/s12046-02201968-2 Verma HK, Samadhiya NK, Singh M, Goel RK, Singh PK (2018) Blast induced rock mass damage around tunnels. Tunn Undergr Space Technol 71:149–158. https://doi.org/10.1016/j.tust.2017. 08.019 Vishwakarma AK, Himanshu VK, Kumar S, Roy MP (2020) Overbreak control in development face blasting of underground metal mine—a case study. In: Proceedings of national conference on advances in mining (AIM-2020), pp 473–482 Wang J, Yin Y, Luo C (2018) Johnson-Holmquist-II(JH-2) constitutive model for rock materials: Parameter determination and application in tunnel smooth blasting. Appl Sci (Switzerland) 8(9). https://doi.org/10.3390/app8091675 Zhou X, Niu J (2020) Method of predicting ore dilution based on a neural network and its application. Sustainability 12:1550 Zhou H, Xie X, Feng Y (2018) Rock breaking method to replace blasting. IOP Conf Series: Mater Sci Eng 322, 022014