The Effect of Seam Dip on the Application of the Longwall Top Coal Caving Method for Inclined Thick Seams

The effect of seam dip on the application of the Longwall Top Coal Caving method for inclined thick seams

By Dao Hong Quang

B.E. (Hons) in Mining Engineering, University of Mining and Geology, Vietnam M.E in Mining Engineering, University of New South Wales, Australia

A thesis submitted in fulfilment of the requirements for the degree of Doctor of Philosophy

School of Mining Engineering The University of New South Wales Sydney, Australia

March 2010 ORIGINALITY STATEMENT

I hereby declare that this submission is my own work and to the best of my knowledge it contains no materials previously published or written by another person, or substantial proportions of material which have been accepted for the award of any other degree or diploma at UNSW or any other educational institu- tion, except where due acknowledgement is made in the thesis. Any contribution made to the research by others, with whom I have worked at UNSW or elsewhere, is explicitly acknowledged in the thesis. I also declare that the intellectual content of this thesis is the product of my own work, except to the extent that assistance from others in the project’s design and conception or in style, presentation and linguistic expression is acknowledged.

Signed ...... Date ......

i Abstract

This thesis presents the results of research into the potential of underground mining methods applicable to inclined thick seams in the Quangninh coalfield, Vietnam. For this research, an inclined thick seam is defined as a coal seam with a thickness of greater than 3.5 metres, and a seam dip ranging from 15 to 35 degrees. This research further investigated the feasibility of applying the most appropriate underground methods for the extraction of inclined thick seams in the Quangninh coalfield, in order to meet the increasing demand for coal products and standards of safety management. The primary objectives of this research are: 1. to investigate the most suitable underground methods applicable to inclined thick seams in the Quangninh coalfield; 2. improve understanding of the operational and geotechnical issues associated with the application of chosen methods in thick seams with steeply dipping conditions. From a risk and operational assessment, the Longwall Top Coal Caving (LTCC) method is considered the most appropriate method for the Quangninh mining situation. The LTCC method, original referred to as the “soutirage” method, offers great potential for the efficient extraction of thick seams by caving coal from the upper section during the mining of the lower section. This significantly reduces the development cost per tonne of coal product. Compared to High reach Single Pass Longwalling, the LTCC method offers a lower face height, which results in smaller and less expensive equipment, and better face conditions. The output of coal per metre of gate road development from the LTCC method is much more than the multi-slice longwal method, and the entry layout of the LTCC method

ii iii is much simpler than that of the multi-slice longwall method. Results from this study indicated that for the extraction of an inclined thick seam, the face retreating along the strike has better operational advantages and has better cavability than the face retreating updip or downdip of the seam. The major operational issues of the LTCC method when extracting inclined seams are: the stability of the support, the difficulties of installation and transport in steeply dipping conditions, and the difficulties in roof control at the transition between face ends and the gateroads. In terms of geotechnical issues, better cavability of the top coal is anticipated for flat coal seams compared to inclined seams. In addition, the chain pillar for flat coal seams is subjected to a higher vertical stress distribution than that of inclined ones. An improved cavability assessment (CPI) method for the categorisation of the cavability of the top coal with four categories, ranging from 1 (excellent cavability) to 4 (very poor cavability), was suggested to assist underground mining design for predicting the cavability of the top coal of an inclined LTCC face. The cavability assessment method was conducted by numerical analysis combined with back analysis from the database obtained from past LTCC practices. Acknowledgements

I would like to extend sincere thanks to following persons and organizations for providing assistance and guidance during the process of research and writing of this thesis: Professor Bruce Hebblewhite, my supervisor, for his assistance and constructive criticism during my research and his support and encouragement which made the completion of this dissertation possible. Dr. Yuejun Cai and Dr. Rudrajit Mitra, my co-supervisors, for assistance in the numerical modelling, providing valuable comments and suggestions on my work. Dr. Serkan Saydam, the postgraduate co-ordinator of the School of Mining Engineering, for his organisational assistance during the time of my research at the School. Ms Bronwen Phillips and Mr Ed Malone for editing this thesis and providing comments during the write-up. Dr. Abouze Vakili for his comments and discussion of the results for my cavability criterion development and modelling. Associate Professor David Laurence for his help, encouragement during my study. Dr. Nguyen Anh Tuan and staff members at the Institute of Mining Science and Technology (IMSAT), Vietnam, for their assistance with compiling geotechnical data of the Quangninh coalfield, and for their constructive criticism. The Ministry of Education and Training, Vietnam, through MOET scholarship, the University of New South Wales through Top-Up scholarship, and the School of Mining Engineering through the grant of Professor Bruce Hebblewhite. Staff members of the School of Mining Engineering for their individual help,

iv v encouragement and providing both resources and friendship during my time with the school. Canh Quang, Danh, Dat, Kha and Cuong for their support and friendship throughout my stay in Australia. Finally, to my family for their constant love and support, and for encouraging me to pursue higher studies. The deepest appreciation is reserved for my mother Du Thi Nguyen, my wife Ly Thi Ninh, and my daughter Trang Ninh Dao whose love, understanding and caring sustained me throughout my study. Table of Contents

Abstract ii

Acknowledgements iv

Table of Contents vi

List of Figures xii

List of Tables xix

1 Introduction 1 1.1 Background ...... 1 1.2 Research objectives ...... 2 1.3 Thesis structure ...... 3 1.4 The contributions of the research ...... 5 1.5 Published papers ...... 6

2 Thick Seam Mining Challenges in the Quangninh Coalfield 8 2.1 Introduction ...... 8 2.2 General geology of the Quangninh coalfield ...... 9 2.2.1 Geological structure of the Quangninh coalfield ...... 10 2.2.2 Structure of tectonic setting ...... 14 2.2.3 Hydrology setting ...... 15 2.2.4 Coal quality ...... 15 2.2.5 Analysis of thick seams reserves in Quangninh ...... 16

vi vii

2.2.6 Assessment of geological conditions in the Quangninh coal- ﬁeld ...... 21 2.3 Thick seam mining challenges in Quangninh ...... 22 2.3.1 Status of thick seam mining in Quangninh ...... 23 2.3.2 Demands for coal production and the availability of coal reserves in Vietnam ...... 29 2.3.3 Discussion ...... 30 2.4 Conclusions ...... 32

3 Potential Underground Thick Seam Mining in the Quangninh Coalﬁeld 35 3.1 Introduction ...... 35 3.2 Literature review ...... 36 3.2.1 The Room and Pillar method ...... 37 3.2.2 High reach Single Pass Longwall method ...... 41 3.2.3 Multi-slice Longwall method ...... 43 3.2.4 Hydraulic mining method ...... 47 3.2.5 Longwall Top Coal Caving method ...... 49 3.3 Potential thick seam mining methods applicable to Quangninh’s conditions ...... 53 3.3.1 Room and Pillar and Blasting Gallery methods ...... 55 3.3.2 High reach Single Pass Longwall method ...... 59 3.3.3 Multi-slice Longwall method ...... 64 3.3.4 Hydraulic mining ...... 66 3.3.5 Longwall Top Coal Caving ...... 69 3.4 Conclusions ...... 72

4 Operational issues for the extraction of inclined thick seams by the LTCC method 75 4.1 Introduction ...... 75 4.2 Face directions in inclined seams ...... 76 4.3 LTCC face equipment ...... 80 viii

4.4 LTCC supports for steeply dipping seams ...... 91 4.4.1 Support tilting ...... 91 4.4.2 Slippage of equipment at an inclined longwall face . . . . . 94 4.4.3 Support transport ...... 98 4.5 Transition between the face ends and the gateroads ...... 99 4.6 Conclusions ...... 102

5 Eﬀect of Seam Dip on the Cavability of Top Coal in the LTCC method 105 5.1 Introduction ...... 105 5.2 Literature review ...... 106 5.2.1 Coal strength ...... 109 5.2.2 Discontinuities ...... 111 5.2.3 The cover depth of the seam ...... 115 5.2.4 Overlying strata ...... 118 5.2.5 Top coal thickness ...... 121 5.2.6 Pre-mining stress regime ...... 122 5.2.7 Moisture sensitivity ...... 124 5.2.8 Discussion ...... 125 5.3 Numerical analyses for underground mining ...... 125 5.3.1 Types of numerical methods ...... 127 5.3.2 The numerical methods used in the current study . . . . . 127 5.4 The eﬀect of face orientation on the cavability of inclined thick seams128 5.4.1 Scope of the modelling ...... 128 5.4.2 Mining geometry ...... 129 5.4.3 Material properties and constitutive models ...... 130 5.4.4 Simulation of face extraction and top-coal caving . . . . . 132 5.4.5 Discussion of numerical modelling results ...... 134 5.5 Conclusions ...... 144 ix

6 Caving mechanism of the Top Coal in the LTCC method for Inclined Thick Seams 146 6.1 Introduction ...... 146 6.2 Mining geometry ...... 147 6.3 Material properties and constitutive models ...... 149 6.4 Simulation of face extraction and top coal caving ...... 150 6.5 Discussion of the numerical modelling results ...... 151 6.5.1 Effect of the seam dip ...... 152 6.5.2 Effect of the strength of coal seam ...... 156 6.5.3 Effect of discontinuity spacing ...... 156 6.5.4 Effect of the thickness of the top coal ...... 159 6.5.5 Effect of the depth of cover ...... 161 6.6 Conclusions ...... 161

7 The eﬀect of seam dip on stress distribution in an LTCC panel 163 7.1 Introduction ...... 163 7.2 Literature review ...... 164 7.2.1 Distribution of the stress in the chain pillar ...... 166 7.2.2 Discussion ...... 172 7.3 The eﬀect of seam dip on stress distribution in the chain pillar . . 172 7.3.1 Mining geometry ...... 173 7.3.2 Material properties and constitutive models ...... 173 7.3.3 Simulation of face extraction and top coal caving . . . . . 175 7.3.4 Simulation of the gateroad in the rib pillar ...... 177 7.3.5 Discussion of the numerical modelling results ...... 178 7.4 Conclusions ...... 188

8 Development of a cavability assessment criterion for the LTCC method in inclined thick seams 190 8.1 Introduction ...... 190 8.2 Literature review ...... 191 8.2.1 Method of vertical displacement in top coal ...... 192 x

8.2.2 The fuzzy cluster evaluation method ...... 193 8.2.3 Empirical cavability assessment method ...... 196 8.2.4 CSIRO’s cavability criterion ...... 198 8.2.5 Numerical cavability assessment method ...... 200 8.3 Discussion ...... 201 8.4 Cavability assessment criterion for inclined thick seams ...... 202 8.5 Other case studies ...... 209 8.5.1 No.8 seam Tay Vangdanh, Vangdanh Coal Company . . . 210 8.5.2 No.6b seam, Lotri, Thongnhat Coal Company ...... 215 8.5.3 No.7 seam, Thanthung, Nammau Coal Mine ...... 218 8.6 Conclusions ...... 222

9 Conclusions and recommendations 223 9.1 Summary and conclusions ...... 223 9.1.1 Introduction ...... 223 9.1.2 Potential mining methods in the Quangninh coalﬁeld . . . 225 9.1.3 Operational issues associated with the application of the LTCC method in steeply dipping conditions ...... 227 9.1.4 Geotechnical issues associated with the application of the LTCC method in steeply dipping conditions ...... 230 9.1.5 Cavability assessment method for extracting inclined thick seams by the LTCC ...... 232 9.2 Recommendations for further research ...... 233

References 235

A Fish functions used in numerical modelling analysis 250 A.1 Fish functions to generate the support elements in 3DEC modelling 250 A.2 Fish functions to generate the top coal caving in FLAC modelling 257

B Model results from FLAC analysis 265 xi

C Multiple regression analysis of the CPI method 270 C.1 Raw results of 3DEC modelling used in the multiple regression analysis ...... 270 C.2 Multiple regression analysis ...... 272 List of Figures

2.1 Map of Vietnam showing the location of the Quangninh coalfield . 9 2.2 Geologic map showing coal regions in the Quangninh coalfield . . 11 2.3 The classification of coal reserves according to thickness ...... 18 2.4 Thick seam classification according to the dip ...... 19 2.5 Thick seam classification according to the depth of cover . . . . . 19 2.6 The classification of thick seams according to the length along the strike of the coal block (L) ...... 20 2.7 The classification of thick seam according to their gas content . . 21 2.8 Vietnamese coal production between 1995 and 2005 ...... 23

3.1 Typical layout of the Room and Pillar method ...... 38 3.2 Typical cross sections of roadway extraction at Westfalen colliery 38 3.3 Room and Pillar mining in Aumance ...... 40 3.4 Diagram of the blasting gallery method ...... 41 3.5 Perspective and close up view of a High reach Single Pass Longwall face ...... 42 3.6 Principle of the Multi Slice Longwall method in three slices . . . . 44 3.7 The multi-slice longwall method separated between slices by a layer of coal or stone parting ...... 45 3.8 Descending multi-slice longwalling with backﬁll method ...... 46 3.9 Ascending multi-slice longwalling with backﬁll method ...... 47 3.10 Concept of Hydraulic Mining ...... 48 3.11 Conceptual Model of LTCC System ...... 50 3.12 The sub-level caving method ...... 52

xii xiii

3.13 The sub-level caving method, separate slices by flexible wire mesh applied Kuzbass coalfield ...... 53 3.14 The LTCC method used in Velenje Mine ...... 54 3.15 The BG method used in Quangninh for steep coal seams . . . . . 58 3.16 Common types of coal loss when using the BG method in the Quangninh coalfield ...... 59 3.17 The effect of extraction height on stability of support ...... 61

4.1 Different directions of the LTCC face for extraction of an inclined thick seam ...... 77 4.2 The potential slippage of supports for the face retreating up-dip in an inclined seam ...... 78 4.3 The longwall face supported by semi-mechanized shields . . . . . 80 4.4 Longwall face supported by hydraulic single props ...... 81 4.5 Operational sequence of semi-mechanized shields used at LTCC faces in the Quangninh coalfield ...... 82 4.6 Operational problems when using the semi-mechanized shields in the Quangninh coalfield ...... 83 4.7 Model ZH1600/16/24Z semi-mechanized support ...... 84 4.8 Mechanized support model VINAALTA - 2.0/3.15 trialed in Vang- danh coal mine ...... 85 4.9 An example of the European designed shield ...... 87 4.10 Operating Range of the European designed support with steeply dipping conditions ...... 87 4.11 An example of Chinese style shield ...... 88 4.12 Model ZFSB 2200/16/24A light duty LTCC Support ...... 89 4.13 Model ZFSB 2200/16/24B light duty LTCC Support ...... 90 4.14 Model ZFQ 2000/16/24 light duty LTCC Support ...... 90 4.15 Analysis of support tilting ...... 92 4.16 Support tilting in an inclined face ...... 93 4.17 Factors leading to support slippage ...... 95 xiv

4.18 A method to improve the stability of the supports in a dipping face 96 4.19 Aligning shield in steeply dipping face ...... 97 4.20 Tentional in face anchorage ...... 97 4.21 Transport support by monorail system ...... 98 4.22 Transition between the face end and maingate ...... 100 4.23 Transition between the face end and the tailgate ...... 101 4.24 Another method of transition between the face end and the tailgate101

5.1 Roof strata in conventional longwall and LTCC methods . . . . . 107 5.2 Three zones in the overburden strata due to longwall mining . . . 108 5.3 Summary of rock mass characteristics, testing methods and theoretical consideration ...... 110 5.4 Schematic illustration of coal cleat geometries ...... 112 5.5 Effect of surface roughness and normal stress on the friction angle of the discontinuity surface ...... 113 5.6 Relationships between shear and normal stress on sliding surface for different discontinuity geological conditions ...... 114 5.7 Vertical stress re-distribution in an across section at the middle of the longwall face ...... 116 5.8 Area involved in Load balance ...... 117 5.9 Fracture zones in top coal ...... 118 5.10 Loading forces acting on top coal ...... 122 5.11 Roof instability due to horizontal stress ...... 124 5.12 Diagram shows the location of the UDEC analysis ...... 129 5.13 Overall context of UDEC model ...... 130 5.14 Monitoring tools used for the investigation of caving mechanism . 133 5.15 Differences in the cavability of top coal between two face directions at different face retreat distances (L) ...... 135 5.16 The cavability of top coal for a face advancing down the dip at different face retreat distances (L) ...... 136 xv

5.17 Cavability of the top coal at the retreat distance of 100m with different seam dips ...... 138 5.18 Effect of horizontal stress on the cavability of the top coal with different seam dips (@60m of retreat distance) ...... 140 5.19 Diagram shows the location of the UDEC analysis for the face retreating along strike ...... 141 5.20 Cavability between flat and inclined seams ...... 142 5.21 Differences in vertical stress distribution around the longwall face between two face directions (at 40m of retreat distance) ...... 143

6.1 Typical mining geometry in the 3DEC analysis ...... 148 6.2 Typical mining geometry in the 3DEC analysis ...... 149 6.3 Comparison of the average vertical displacement in different seam dips ...... 152 6.4 Status of top coal blocks with different seam dips (20m of retreat distance) ...... 154 6.5 Average vertical displacement of top coal along the longwall panel in different seam dips ...... 155 6.6 Caving perforamnce with different seam dips ...... 155 6.7 Coal strength vs. cavability of top coal ...... 156 6.8 Effect of the joints/cleats spacing on the cavability of top coal . . 157 6.9 Status of top coal blocks with different bedding spacing (at 20m of retreat distance) ...... 158 6.10 Effect of the bedding (horizontal) spacing on the cavability of top coal ...... 159 6.11 Caving situation between different top coal thickness (face advance distance 14m) ...... 160 6.12 Effect of the seam thickness on the caving mechanism of the top coal160 6.13 Caving performance vs depth of cover ...... 161

7.1 Typical longwall layout in the Quangninh coalﬁeld ...... 164 7.2 Redistribution of vertical stress around a single longwall . . . . . 167 xvi

7.3 Distribution of the stress in the goaf area ...... 168 7.4 Load balance in the ribside pillar ...... 168 7.5 Relationship between rib pillar width and the vertical closure in gate roadways ...... 170 7.6 Relation between the pillar width and gateroad performance . . . 171 7.7 Diagram shows the location of FLAC analysis ...... 173 7.8 Relationship between the applied stress and the volumetric strain used in FLAC analysis ...... 176 7.9 Vertical stress distribution around a longwall panel in a flat seam 178 7.10 Vertical stresses at the centre of top coal in flat seams at different panel widths ...... 179 7.11 Vertical stress distribution in the chain pillar, updip side of the longwall panel in different seam dips (α) ...... 180 7.12 Vertical stress distribution in the chain pillar, downdip side of the longwall panel in different seam dips (α) ...... 181 7.13 Vertical stress at centre of coal seam, downdip side of the longwall panel in different seam dips ...... 181 7.14 Values of maximum axial force for different locations of gateroad between flat and inclined seams ...... 182 7.15 Yield zones above the roof of longwall face in different seam dips . 183 7.16 Coal strength vs depth of peak stress distribution at the chain pillar, downdip side of a longwall panel ...... 184 7.17 Vertical stress distribution in the chain pillar beside the maingate in a flat coal seam with different vertical/horizontal stress ratios (k)185 7.18 Vertical stress at the centre of a flat coal seam in different vertical/horizontal stress ratios (k) ...... 186 7.19 Vertical stress distribution in the chain pillar, downdip of a longwall panel for inclined seam under different vertical/horizontal stress ratios (k) ...... 187 7.20 Vertical stress in the chain pillar, downdip side of an inclined (300) longwall panel under different vertical/horizontal stress ratios (k) 187 xvii

8.1 The face distance of positions where the vertical displacement reaches 100mm with diﬀerent UCS values ...... 193 8.2 Comparision between Jia’s classiﬁcation and the CPI method . . . 207 8.3 The relationship of the CPI vs the recovery rate of top coal in Chinese LTCC faces ...... 207 8.4 Some lithological sections of No.8 seam, Tay Vangdanh ...... 210 8.5 Face conditions in No.8 seam Tay Vangdanh ...... 212 8.6 Status of caved coal behind the longwall face ...... 213 8.7 Typical lithological section of No.8 seam, Tay Vangdanh . . . . . 214 8.8 Geological section of some boreholes in No.6b seam, Lotri, Thongn- hat Coal Company ...... 216 8.9 Typical lithological section of 3 boreholes in No.7 seam, Thanthung, Nam Mau Coal Company ...... 218 8.10 Face condition at No.7 seam, Thanthung ...... 219 8.11 Caved coal size behind the longwall face ...... 221

B.1 Contours of vertical stress and the axial forces of liner elements for flat coal seam, pillar width = 5m ...... 265 B.2 Contours of vertical stress and the axial forces of liner elements for flat coal seam, pillar width = 10m ...... 266 B.3 Contours of vertical stress and the axial forces of liner elements for flat coal seam, pillar width = 15m ...... 266 B.4 Contours of vertical stress and the axial forces of liner elements for flat coal seam, pillar width = 20m ...... 267 B.5 Contours of vertical stress and the axial forces of liner elements for 300 dipping seam, pillar width = 5m ...... 267 B.6 Contours of vertical stress and the axial forces of liner elements for 300 dipping seam, pillar width = 10m ...... 268 B.7 Contours of vertical stress and the axial forces of liner elements for 300 dipping seam, pillar width = 15m ...... 268 xviii

B.8 Contours of vertical stress and the axial forces of liner elements for 300 dipping seam, pillar width = 20m ...... 269 List of Tables

4.1 Speciﬁcations of 11 types of light duty LTCC supports ...... 90

5.1 Materials and joint properties of different layers used in the models 132 5.2 CD and TAVD for different directions of longwall ...... 137 5.3 CD and TAVD for different influenced parameters ...... 141

6.1 Parameters studied in the 3DEC analysis ...... 151

7.1 Properties of coal and rockmass used in FLAC models ...... 174 7.2 Properties of liner element used in the FLAC analysis ...... 177

8.1 Cavability classiﬁcation of some Chinese LTCC faces developed by Jia ...... 195 8.2 The eﬀect of individual parameters on the cavability of top coal in Jia’s criterion ...... 196

8.3 Ratio of cover depth to coal UCS versus µ1 ...... 197

8.4 Category of immediate roof versus µ2 ...... 197

8.5 Category of main roof versus µ3 ...... 197

8.6 Ratio of cut to cave versus µ4 ...... 197

8.7 Cleat spacing versus µ5 ...... 197

8.8 Band thickness versus µ6 ...... 197

8.9 The UCS of band layer versus µ7 ...... 198 8.10 Weightings of diﬀerent factors ...... 198 8.11 Cavability classiﬁcation developed by Shandong Coal Mining Ad- ministration ...... 198

xix xx

8.12 CSIRO’s cavability classification ...... 199 8.13 Cavability classification developed by Vakili ...... 201 8.14 Regression Statistics ...... 203 8.15 Comparision of CPI and different cavability assessment methods . 206 8.16 Comparison between the database of Chinese LTCC mines and the CPI values ...... 208 8.17 Cavability classification of the top coal based on the CPI values . 208 8.18 Categorisation of individual parameters in the CPI criterion . . . 209 8.19 Weighting of individual parameters in CPI criterion ...... 209 8.20 Cavability assessment of No.8 seam Tay Vangdanh by the CPI method ...... 215 8.21 Cavability assessment of No.6 seam, Lotri area, Thongnhat Coal Company by the CPI method ...... 217 8.22 Cavability assessment of No.7 seam Thanthung by the CPI method 220

C.1 Raw results from 3DEC modelling ...... 271 C.2 Regression Statistics ...... 273 C.3 ANOVA table ...... 273 C.4 Results for multi-regression analysis ...... 274 Chapter 1

Introduction

1.1 Background

This thesis presents the results of research into a potential underground mining method applicable to inclined thick seams in the Quangninh coalfield, Vietnam. For the purpose of this research, an inclined thick seam has been defined as a coal seam having a thickness greater than 3.5 m and dip ranges from 150 to 350. As outlined later in this thesis, there is a large proportion of inclined thick seam reserves which are suitable for extraction by underground methods in the Quangninh coalfield, Vietnam. A number of research investigations conducted on behalf of the Vietnamese mining industry have been made over the past twenty years to improve coal production and safety from these resources with the maximum level of coal recovery. However, at this point in time, the output of mining operations in the extraction of inclined thick coal seams has achieved limited results in both production and safety management. One of the main reasons is that most mining methods applied to inclined thick seams are undertaken by utilising methods which are applied for flat seams from overseas or other coal mines, with little or no understanding of the operational and geotechnical issues associated with these methods when extracting dipping coal seams. Another reason is that most of the underground mines in Quangninh operate with low technology and simple equipment that require a large amount of manual work. In the plan of the development of the mining industry in Vietnam, mining pro-

1 2 duction from open pit mining will be scaled down and closed by 2014 to protect the environment in the area, especially to maintain the natural landscape of Ha- long bay nearby. Coal reserves in the Quangninh coalfield will only be available for extraction by underground methods. In order to meet the rapidly increasing demand of coal for the Vietnamese economy, improvements in underground mining technology are necessary. Therefore, it is believed that productivity and safety standards could be much improved if suitable mining methods, coupled with appropriate equipment for extracting the inclined thick coal seams can be defined to suit these particular geotechnical conditions. In parallel with this, it is also necessary to improve the understanding of both the operational and geotechnical issues associated with these methods in the steeply dipping conditions in the Quangninh coalfield. This is the primary goal of this research.

1.2 Research objectives

The purpose of this research contains the following key components:

1. Evaluation of the underground thick seam mining methods with the greatest potential that are applicable to the inclined thick seams under the current mining conditions in the Quangninh coalﬁeld.

2. Investigation of the operational issues associated with the application of the chosen method(s) for inclined thick seams.

3. Investigation of the geotechnical issues associated with the application of the chosen method(s) for inclined thick coal seams.

Site visits were made to two coal mines in the Quangninh coalfield: Vangdanh and ThongNhat. The purpose of the site visits was to gain an experience of operating conditions and practices of the coal mines, and to carry out a detailed investigation of the geotechnical aspects of the mine design for back analysis of the suggested cavability assessment method. The research began with an updated investigation of coal reserves and geo- mining conditions in the Quangninh coalfield, in order to provide a focus for the 3 issues that need to be addressed. This investigation also provided a means of evaluating the underground method most suitable for these geo-mining conditions. Through a literature review of underground thick seam mining methods currently applied around the world, the study focussed on evaluating five main thick seam mining methods, mostly well known, which are applicable to Quangninh’s geo-mining conditions. They are:

• Bord and Pillar method, and modiﬁcations of this method;

• Increase extraction height of single pass longwall;

• Multi-slicing longwall;

• Hydraulic mining, and;

• Longwall Top Coal Caving.

Arising from risk assessment, two mining methods: Hydraulic mining and Longwall Top Coal Caving appeared as the most suitable methods applicable to Quangninh’s geological conditions. However, greater detailed assessment showed that hydraulic mining has limitations in its application to the Quangninh coalﬁeld under the current technology, particularly from the operational aspect. The Longwall Top Coal Caving method, which was initially applied in Euro- pean countries where it was known as the “soutirage”method, and then significantly developed in China, was identiﬁed as a top priority thick seam method compared to the other options. As a result, the major component of this research then focussed on improving the knowledge base associated with applying LTCC to inclined thick seams.

1.3 Thesis structure

The progress of this research is divided into three parts and addressed in nine chapters of this thesis. The first part outlines the reasons why the Long- wall Top Coal caving method is at the forefront of underground thick 4 seam mining in the Quangninh coalfield. Chapter 2 provides background information explaining why the extraction of inclined thick coal seams is an issue to be addressed in Quangninh. A detailed investigation of the coal reserves was undertaken in some underground coal mines which contain reliable geological information, and these resources will be the main target of coal production in the next decades. The reserve analysis focussed on evaluating parameters that affect the selection of potential mining methods, such as seam thickness, seam dip, cover depth of the coal seam, and the gas content in the seam, to provide a clear view on the quality and quantity of thick seam reserves in the Quangninh coalfield. In this chapter, demand for coal production, the availability of coal reserves and the level of mechanization of the mining industry in Vietnam will also be analyzed. Chapter 3 evaluates the most suitable mining methods applicable to inclined thick seams in the Quangninh coalfield. This evaluation was undertaken by investigating the advantages and disadvantages of the mining methods which have been currently applied in coal mines around the world to extract thick coal seams, in parallel with an evaluation of the geological conditions in the Quangninh coalfield, where seam thickness and seam dip are the main factors of consideration. In addition, the current underground mining operations and the limitation of the underground mining methods which are being currently applied to the Vietnam mining industry will also be discussed in order to find the most suitable method applicable to inclined thick seams in Quangninh under current mining technology. This chapter concludes with acknowledgement that the Longwall Top Coal Caving was identified as the top priority method to be applied to inclined thick seams, in comparison with other options. The second part of the thesis concentrates on the operational and geotechnical issues associated with the application of the LTCC method to inclined thick coal seams in the Quangninh coalfield. Chapter 4 studies the operational issues associated with the application of the LTCC method for the extraction of inclined thick seams in Quangninh. The research presented in Chapter 4 focuses on the equipment required in the Quangninh’s geo-mining conditions, and the stability of the mining equipment when applied to an inclined 5 face. In addition, the difficulties of roof control at the intersection between the face ends and the gate roadways are also discussed. In terms of geotechnical issues, this study pays attention to the effect of seam dip on the caving mechanism of top coal when using the LTCC method for the extraction of inclined thick seams. It contains:

• research on the eﬀect of seam dip on the caving mechanism of top coal when using the LTCC method for extracting inclined thick seams, described in Chapter 5 and Chapter 6.

• the investigation of the eﬀect of seam dip on stress distribution in the chain pillar, presented in Chapter 7.

The last part of the thesis, which is presented in Chapter 8, is a desciption of the development of a method to predict the cavability of the top coal when extracting inclined thick coal seams using the LTCC method. The development of the method was undertaken by numerical analysis based on the six most signiﬁcant parameters of the stratiﬁed rockmass. The results from the numerical analysis are back analyzed with observations obtained from past experience. Chapter 9 presents the conclusions drawn from this work. Recommendations for future works arising from this project are also illustrated in this chapter.

1.4 The contributions of the research

This research has contributed to the advancement of knowledge in the following ways:

• The information of coal quality and quantity as well as the geomechanical conditions of the Quangninh coalﬁeld have been analysed. In addition, the underground mining methods with the most potential and application to inclined thick seams in the Quangninh coalﬁeld have been determined. 6

• Understanding of the operational issues associated with the application of the LTCC method for extraction of inclined thick coal seams has been improved.

• Eﬀect of seam dip on the caving mechanism of the top coal when using the LTCC method for the extraction of inclined thick coal seams has been investigated.

• Models have been developed to explain the diﬀerences of stress distributions around the longwall face when extracting thick coal seams by the LTCC method under ﬂat or inclined conditions.

• An assessment method has been developed for predicting the cavability of the top coal when using the LTCC method for extraction of inclined thick seams. This criterion will assist mine planners to evaluate the feasibility of the application of the LTCC method into certain geo-mining conditions.

1.5 Published papers

The following papers containing extracts of this thesis were published during the period of research:

• Quang, D.H., Cai, Y., and B. Hebblewhite, 2008. Prospects of underground thick seam mining potential in the Quangninh coalﬁeld, Vietnam. Proceed- ings of the International Conference in Advances of Mining and Tunnelling. Hanoi, Vietnam, Publishing House for Science and technology. pp.70-83.

• Quang, D.H., Cai, Y., and B. Hebblewhite, 2008. Numerical Analysis of Some Geotechnical Factors Inﬂuencing the Application of the Longwall Top Coal Caving. Proceedings of the 27th International Conference on Ground Control in Mining. S. Peng, M. Christopher, G. Finﬁnger, S. Tadolini, A. Khair, K. Heasley, and Y. Luo, Eds. Morgantown, West Virginia, USA. pp.222-228. 7

• Quang, D.H., Mitra, R., and B. Hebblewhite, 2009. Eﬀect of Seam Dip on Face Orientation of Longwall Top Coal Caving, Proceedings of the 43rd U.S. Rock Mechanics Symposium ARMA09-110 (CD-ROM). American Rock Me- chanics Association, USA. paper number 110 (pp.1-9). Chapter 2

Thick Seam Mining Challenges in the Quangninh Coalﬁeld

2.1 Introduction

The Vietnamese mining industry has long been faced with challenges and difficulties associated with the underground extraction of thick coal seams. These are not only related to operational, technical, safety and economic issues, but also to the pressure of improving coal production and resource recovery in order to meet the increasing demand of the domestic economy. The extraction of thick seams in the Quangninh coalfield becomes more complicated as many coal seams are located in areas with difficult geological conditions, including those with structural disturbance, complicated surrounding strata, and especially, steeply dipping conditions. The demand for coal production in Vietnam will increase rapidly in the next decades. This demand, coupled with the requirement for improvement in the recovery rate of the coal resources, makes it necessary to improve underground mining production, especially in thick seam mining. This chapter will review, update and analyze the results of data on the potential of thick seams in the Quangninh coalfield, and explain the reasons why underground thick seam mining is a challenge here.

8 9

2.2 General geology of the Quangninh coalﬁeld

The Quangninh coalﬁeld is located in Halong city, which is approximately 150km North-East of Hanoi, the capital of Vietnam. Travel time to Halong city is roughly three to four hours by road from Hanoi. The location of the Quangninh coalﬁeld is illustrated in Figure 2.1.

Figure 2.1: Map of Vietnam showing the location of the Quangninh coalﬁeld (GotoVietnam, 2009)

The Quangninh carboniferous basin is part of the South China offshore, which has complicated geological conditions with many faults or dykes. The sediment basin that contains the coal deposit is 130km in length and 10 - 30km in width, and covers an area of 1400km2 (Thanh and Dac, 1995; Industry Investment Con- sulting Company, 2005; Dac and Tuan, 2006). According to Dac and Tuan (2006), the Quangninh coalfield was investigated by 2,000,000 m3 of exploration trenches with approximately 5000 bore holes which generated nearly 300 reports. However, the details of these investigations vary regarding the depth of the carboniferous 10 strata. The density of exploration holes to the depth of -501, is 10 - 30 bore holes/km2, while the average density of bore holes to a depth of -150 is 14 bore holes/km2, and to a depth of -300 is 3 - 5 bore holes/km2. The exploration of the coal reserves below the depth of -300 is limited, with only 59 exploration bore holes being drilled. The depth of those holes mainly range from 450 - 600m, and only 10 of those bore holes have a depth of 950 - 1200m (Dac and Tuan, 2006). Therefore, coal resources in the Quangninh coalfield in the range of a cover depth of -300 or deeper, compared to sea level, is a raw estimate that needs to be further investigated. Coal resources in the Quangninh coalfield to a depth of -300 below sea level (equivalent to a cover depth of 320-700m below the ground surface) are estimated to be 4.0 billion tonnes, while coal resources at a greater depth of -300 below the sea level are estimated at 6.9 billion tonnes (Industry Investment Consulting Company, 2005). The thickness of coal seams in the Quanhninh coalfield varies largely between the different coal areas. In some locations, the seam thickness changes rapidly over a short distance within the same coal district area. Seam thickness ranges from thin (less than 1.2m), medium (from 1.2 to 3.5m) to thick (greater than 3.5m). Some coal seams have a thickness of up to 92.5m (Ngan, 2004). It is estimated that 66.3% of coal reserves in the Quangninh coalfield belong to thick seams, which have a thickness of greater than 3.5m (Thanh and Dac, 1995). The detailed geo-mining conditions of the Quangninh coalfield are described in following sections.

2.2.1 Geological structure of the Quangninh coalﬁeld

The stratigraphic sequence in the Quangninh coalﬁeld is at the middle of Hongai strata that incorporates the outcrop areas of Yentu and Hongai anticlines. The thickness of the carboniferous strata in the Quangninh coalﬁeld varies from 200m to 2900m, and can be divided into three main regions: Uongbi, Hongai, and Campha as illustrated in Figure 2.2.

1The depth referred here is the depth in metre as compared to the sea level 11 map showing coal regions in the Quangninh coalﬁeld (modiﬁed after Industry Investment Consulting Company, 2005) Figure 2.2: Geologic 12

The characteristics of the carboniferous strata to a depth of 300m below the sea level in those areas are described as follows (Ngan, 2004; Industry Investment Consulting Company, 2005):

• Uongbi region: The carboniferous strata in this area has an approximate thickness of 2900m and is separated by a large fault named Trungluong fault. Coal resources in this area, to a depth of -300 below sea level, are estimated at 1.35 billion tonnes. The carboniferous strata in the Southern area of the Trungluong fault contains 54 to 64 coal seams which have a simple structure with gentle warping.

The carboniferous strata in the northern area of the Trungluong fault is more complicated than that of the southern area. This carboniferous strata has a thickness of 150 - 700m and contains 6 - 20 coal seams. The coal seams in this area are crumpled by a series of synclines and anticlines both along the strike and down dip. In some areas, the seam dip changes rapidly over a short distance. In addition, the coal seams in this area are separated into small blocks because of many small faults.

• Hongai region: This area has a thickness of 200 - 1500m and contains 2 to 17 coal seams. Coal resources in this area (to a depth of -300 below sea level) are estimated to be 740.4 million tonnes. The distance between the seams ranges from 40 - 60m, although in some areas, this distance decreases to 15m. The variation of seam thickness and seam dip in this coalﬁeld area changes rapidly from very thick to medium or even thin coal over short distances along the strike or along the dip, resulting in complications in choosing the proper mining method for the extraction of that resource. In addition, the appearance of various faults in the area has separated the coal seams into small blocks.

• Campha region: This carboniferous strata contains 2 to 36 coal seams and has a thickness of 450 - 1350m. Coal resources in this area (to a depth of -300 below sea level) are estimated to be 1.96 billion tonnes. The coal 13

seams in this area are mainly thick seams. The thickness of the seam varies from thin to very thick, up to 92.5m in the thickest. Coal seams in this area have low variations in seam thickness and seam dip.

The carboniferous strata in the Quangninh coalfield consists of various layers: coal, coal-bearing claystone, shale, mudstone, sandstone and conglomerate. The layers of sequence lie conformably, according to the processes of sedimentation. According to Hoan (1990), the properties of coal and rockmass vary rapidly in different parts of the Quangninh coalfield. Generally, the characteristics of each type of rockmass in Quangninh are as follows:

• Shale or coal-bearing claystone is usually located near the coal seams. This type of rockmass consists of a number of thin beddings with a thickness of 2 - 5cm. It can absorb water and becomes loose and easily separated from the other layers when in contact with water. The uniaxial compressive strength (UCS) of this rockmass ranges from 100 - 190kg/cm2.

• Thin-laminated mudstone is usually located above shale layers or the coal- bearing claystone. The thin-laminated mudstone mostly consists of a number of layers with a thickness of 10 - 35cm. In some areas, there are other types of weaker rockmass laid between the mudstone layers, such as shale or coal-bearing claystones, that separate them easily from each other. The UCS of this rockmass ranges from 90 - 480kg/cm2.

• Thick-laminated mudstone is usually located above the thin-laminated mudstone, and is considered the bridging rockmass between the thin-laminated mustone and the sandstone layers. The UCS of this rockmass ranges from 2100 - 1200kg/cm2.

• The sandstone layer consists of numerous beddings with a thickness varying from 10 to 150cm. The UCS of this rockmass ranges from 380 - 2380kg/cm2.

• Conglomerate consists of numerous bedding layers with a thickness of 100 - 300cm. In some locations the bedding layers of the rockmass is very strong 14

and diﬃcult to separate. The UCS of this rockmass ranges from 760 - 2770kg/cm2.

Research on the characteristics of the joints in rockmass undertaken by Hoan (1990) has shown that the rockmass in the Quangninh area has four joint systems, where two joint systems are at right angles to the bedding layers, 1 joint system is parallel to the bedding layers, and 1 is oblique to the bedding layers. The locations of these jointing system are closely connected to the direction of the rockmass, and they change in parallel to the change of the bedding layers. Due to the distribution of the jointing systems, the bedding layers in the rockmass have the characteristic of being separated from each other.

2.2.2 Structure of tectonic setting

The Quangninh carboniferous basin is comprised of a series of faults, dykes or synclines and anticlines in directions both along the strike and along the dip of the coal seams. Faults in the Quangninh coalfield mainly have large damage zones and dip-slip throw, which generally range from 20 - 40m, but in some cases, reach 150m. In addition, the rockmasses which are located near the faults are mainly weak, as a result of the effect of damage zones in the faults. As a result, it is extremely difficult for the longwall face to pass through the faults which have a large offset and damage zone in terms of both technical and economical aspects. An investigation into the geological conditions of the Quangninh coalfield, undertaken by Thanh and Dac (1995), pointed out that large faults have separated 60.84% of the coal seams into small blocks with a length of less than 600m along the strike or downdip. The faults also divide 27% of coal seams in the Quangninh coalfield into blocks with a triangular shape. As a result, the presence of faults or dykes that divide the seam into small blocks would certainly influence the efficiency of the mining operations, producing more complications in the Quangninh carboniferous strata. 15

2.2.3 Hydrology setting

At present, many underground coal mines are operating under large depth cover. For instance, the Maokhe Coal Company is operating at a depth of 150m below sea level (which ranges from 200m to 350m below the ground surface in this area). The Vangdanh Coal Company is extracting coal reserves at a depth of 250m to 450m from the hilly ground surface, while in the Mongduong Coal Company, coal reserves are extracted at a depth of 97.5m below sea level (130 - 200m below the hilly ground surface) (Dac and Tuan, 2006). It is estimated that after 2010 all underground coal mining companies will have plans for extracting coal resources up to a cover depth of -300m below sea level. Therefore, underground mining operations will be considerably aﬀected by groundwater when mining operations are undertaken at a large cover depth. In addition, the volume of groundwater varies according to seasons because it has a close relationship with the surface water. In summer, the volume of groundwater is 15 to 20 times as much as that of the ground water in winter. The maximum volume of groundwater in some underground mining areas in summer sometimes can reach up to 5000 - 6000 m3/h (Du, 1999).

2.2.4 Coal quality

Coal in the Quangninh area is mainly anthracite and semi-anthracite, with high specific energy and a low sulphur content. Its calorific value ranges from 6000 to 8000 kcal/kg and the sulphur content is lower than 1%. Most of the seams in the Quangninh coal basin have a low content of methane, which accounts for less than 1m3 of methane per tonne. The methane content in some seams in the Maokhe coal mine, however, accounts for 5.9m3/tonne (Dac, Ba and Chinh, 2005). Therefore, the potential risk of spontaneous combustion or gas explosions is considered low in the Quangninh coalfield. 16

2.2.5 Analysis of thick seams reserves in Quangninh

It is known that a large proportion of coal reserves in the Quangninh coalfield belong to thick seams. Analyses of coal reserves in the Quangninh coalfield have been undertaken by numerous researchers (Thanh and Dac, 1995; Du, 1999; Du, 2003; Tuan, Dac, Du, Ngan and Thang, 2004). These studies have provided much information about the conditions in the Quangninh coalfield. However, these studies investigate coal reserves with different ranges of seam thickness, from thin to thick, and do not focus specially on thick seams. In addition, some technical parameters of thick seams that seriously affect the evaluation of suitable mining methods, such as the cover depth, and seam dip, have not been investigated in detail. This section will review, update and analyze recent information on the potential of thick seams in Vietnam and explain the results of data analysis in more detail, in order to achieve better understanding. Based on the definition in the Regulations of Technique and Safety in Viet- namese underground mining operations issued in 2006 (Vietnamese Ministry of Industry, 2006a), a seam having a thickness of greater than 3.5 metres has been defined as a thick seam under the current study. The analysis of this thesis was undertaken by using information of both previous studies and geological reports of exploration companies. The collection of coal seam databases, conducted by Tuan et al. (2004), has been used as a main source for this assessment. Parameters that may affect the evaluation of potential mining methods were analyzed. They are: cover depth, seam thickness, seam dip and coal properties. The databases used in the analysis include:

• Data on 69 coal seams. Each coal seam consists of its characteristics such as physical, and geomechanical properties and quality information,

• The reports of geological exploration results in preliminary and detailed exploration periods and general reports of coal reserves and geological information of underground coal mines in Quangninh.

• Previous studies, technical reports (Tuan et al., 2004; Industry Investment 17

Consulting Company, 2005), and other information sources.

• The status of mining areas, including mined out areas and remaining coal reserves.

• The plan of mining development in coal seams including: the area in operation, mines in the development stage, and other information provided by underground coal companies.

• The temporary classiﬁcation of coal mines according to gas content by the Minister of Industry, Vietnam (Vietnamese Ministry of Industry, 2006b).

The investigation in this study focuses on analyzing seam databases from some underground mine sites which have reliable geological information with total coal reserves of 560,845,000 tonnes to a depth of -150 (150m below sea level). They are: Maokhe, Vangdanh, Nammau, Duonghuy, Ngahai and Thongnhat. Due to a confidentiality agreement, the detailed databases of coal reserves in mine sites were not be disclosed. The first step in the data analysis was to clarify the proportion of coal reserves with different seam thicknesses. The results from the analysis show that 3.4% of coal reserves have a thickness of less than 1.2 metres. The coal reserves with a thickness between 1.2 and 3.5 metres constitute 39.8% of the total investigated coal reserves. 56.7% of coal reserves are reported as having a thickness greater than 3.5 metres, where coal reserves having a thickness greater than 3.5 and less than 5.0 metres account for 23.5% of the total investigated, coal seams having a thickness of greater than 5.0 metres and less or equal to 9.0 metres constitute 26.8%, and 6.4% of investigated coal reserves have a thickness of more than 9.0 metres. A detailed classification of coal reserves according to seam thickness is presented in Figure 2.3. The second significant criterion in the data analysis, affecting the evaluation of the potential mining methods for the extraction of thick seams, is the seam dip. The dip of a seam is classified according to the criterion defined by the Vietnamese Ministry of Industry (2006a) as follows: 18

Figure 2.3: The classiﬁcation of coal reserves according to thickness

• A seam having a dip of less than 150 is deﬁned as a ﬂat seam.

• Seams having a dip of between 150 and 350 are considered to be inclined seams.

• Seams having a dip ranging from greater than 350 and less than or equal to 550 are considered to be steep seams, and

• Seams having a dip greater than 550 are deﬁned as highly steep seams.

Based on this information, it can be observed from the analysis that most of the seams are inclined. The result indicates that 10.8% of thick coal reserves have a dip of less than 150. 58.2% of thick coal reserves are inclined seams, which range from 150 to 350. The percentage of thick coal reserves which have a seam dip ranging from 350 to 550 constitutes 20.8%, and the rest of the coal reserves (10.3%) are reported as highly steep seams (the seam dip is more than 550). Details of thick seam classification according to their dip are illustrated in Figure 2.4. Another critical parameter affecting the selection of a mining method is the average depth of cover of the seams. According to this classification, 21.3% of thick seams are classified as shallow seams with an average depth of cover less than or equal to 200m. 48.0% of the seams are at moderate depth (between 200m 19

Figure 2.4: Thick seam classiﬁcation according to their dip (α, degree) and 400m), and the rest of the thick seams (30.7%) are considered to be deep seams (greater than 400m). Figure 2.5 illustrates the results of the cover depth of the seams analyzed in the database.

Figure 2.5: Thick seam classiﬁcation according to the depth of cover

The configuration of the blocks of coal seams is also one of the parameters that determines the efficiency of the application of mining methods. The dimensions of the coal blocks affect the time of the installation of mining equipment, mining operations and moving faces, and consequently influence the efficiency of mining operations. Investigations from the study show that due to the appearance of many faults and dykes in the Quangninh coalfield, the coal seams are divided into small blocks along the strike. The main challenge is that those faults mostly have 20 a large damage zone so it is difficult for the longwall face to pass through, both technically and economically. Figure 2.6 presents the classification of the length of coal blocks along the strike in the investigated thick coal seams. According to the classification, 7.1% of thick coal seam reserves have a block length along the strike of less than 500m. The thick coal seam reserves, having a block length along the strike between 500m and 1000m, account for 40.4% of the total investigated. 30.4% of thick seam reserves have a block length along the strike of greater than 1000m and less than 1500m, and 22% of thick seams reserves have a block length along the strike greater than 1500m.

Figure 2.6: The classiﬁcation of thick seams according to the length along the strike of the coal block (L)

The properties of the coal seam also have a significant effect on the selection of mining methods and design processes for any mining operation. Therefore, the gas content of the coal seams is also investigated in this study. Figure 2.7 shows the classification of the gas content of thick seams in the Quangninh coalfield. The gas content of the seams is classified according to the Regulations of Technique and Safety in Vietnamese underground mining operations (Vietnamese Ministry of Industry, 2006a) and by interpolation from the test results obtained directly from the current working faces. According to the results of this analysis, 85.7% of thick seams contain a low gas content (less than 5 m3 gas per tonne of coal). 2.7% of thick coal seams have a moderate gas content, between 5m3 and 10m3 gas 21 per tonne of coal. There are no thick coal seams containing a high gas content in the investigated coal reserves (a gas content of greater than 10m3 and equal or less than 15m3 gas per tonne of coal). The rest of the seams (11.6%) were classified as having a very high gas content (which is more than 15m3 gas per tonne of coal).

Figure 2.7: The classiﬁcation of thick seam according to their gas content

2.2.6 Assessment of geological conditions in the Quangn- inh coalﬁeld

Results from the analysis of coal reserves in the Quangninh coalfield clearly show that a large proportion of its reserves belong to thick seams. Nevertheless, coal reserves in the Quangninh coalfield are mostly located in areas which have difficult geological conditions, for example, structural disturbance, dipping seams and complicated surrounding strata that may restrain mining productions. They are:

• Geological conditions: the geological conditions in the Quangninh coalﬁeld are complicated by many folds and faults. Coal seams are mainly steep- ing, causing instability in roof support in the longwall face operations and diﬃculty in mining operations.

• Small coal seam blocks: coal seams are separated by faults with the length along the strike ranging from 500 to 1500m. This makes the application 22

of heavy equipment diﬃcult, due to high costs and time consumed in the longwall face movement.

• Steep seams: a large proportion of coal reserves in the Quangninh coalﬁeld consist of inclined thick seams. The high seam dip may cause diﬃculties in mining operations.

• High variation in seam dip and seam thickness: The variation of seam dip and seam thickness is high, causing diﬃculty in the application of the multi-slice longwall method.

2.3 Thick seam mining challenges in Quangninh

Underground mining has been operating in Quangninh for many years. Coal reserves in the Quangninh coalfield were extracted by French colonialist companies in the early to mid 20th century. When the Vietnamese government took control of the country, the state owned coal companies started renovating the old mining collieries acquired from the French and constructed new mines with the aim of increasing the production of coal in Quangninh to 10 million tonnes in 1980. However, due to both the complications of geo-mining conditions and poor mining equipment, coal production remained at 4 to 6 million tonnes per year for many years (Du, 2003). Since 1990, however, much improvement in both technology and management has been seen in coal mining operations. This has resulted in a dramatic increase in coal production. For instance, in 1994, coal mining companies only exploited 9.4 million tonnes, but this figure grew to 34.7 million tonnes in 2005, 95% of which came from the Quangninh coalfield (Tuan, 2005; Dac and Tuan, 2006; Industry Investment Consulting Company, 2005). The output from the longwall face increased from 40,000 - 60,000 tonnes per year (the face supported by timber props) to approximately 150,000 tonnes per year. Details of coal production over the period between 1995 and 2005 are presented in Figure 2.8. At present, coal production from underground mining in the Quangninh coalfield is undertaken mainly by 14 Coal Companies. According to Tuan (2005), 23

Figure 2.8: Vietnamese coal production between 1995 and 2005 (Tuan, 2005; Dac and Tuan, 2006)

52.3% of the total of 12.5 million tonnes of coal product from underground mining in 2005 was from thick seam mining. The thick seams in Quangninh were mainly extracted by three methods: Longwall Top Coal Caving (LTCC), Multi-slice longwall without septum (MSL), and multi-slice longwall with septum (MSLS). A detailed description of these mining methods is provided in Chapter 3. In parallel with the development of the Vietnamese economy, underground mining also underwent many stages of improvements, especially in the last two decades. The following sections will only discuss the current status of underground thick seam mining in the Quangninh coalﬁeld.

2.3.1 Status of thick seam mining in Quangninh

For a long time, underground mining in Vietnam faced many challenges in the extraction of thick seams, as the maximum cutting height for the longwall faces of the thick seam was limited to 2.2 - 2.5m. Initially, the flat and inclined thick seams in the Quangninh coalfield were all extracted by a single pass longwall method. In this method of extraction, most of the longwall faces were supported by wooden props, while coal was extracted from the face by the drilling and blasting method. Due to being supported by 24 wooden props, the height of the longwall face was limited to approximately 2.0 - 2.2m, at which average high workers could reach the roof and attach the beam, as well as control the potential risks of a roof fall. As a result, the major operational problem of thick seam extraction was that coal is either left in the roof or, more commonly, in the floor of the mining panel. In an attempt to increase coal production and improve the rate of coal recovery, the multi-slice longwall method was applied for the extraction of thick seams with a thickness greater than 4.5m. This mining method was applied in the late 1980s in some underground coal mines in Quangninh, such as Vangdanh and Halam, with the aim of improving coal recovery in thick coal seam mining (Du, 2003). In this method, a thick coal seam was extracted by two or more slices which were taken parallel to the floor in descending order. The slices were extracted one after another at independent faces from top to bottom either simultaneously or non-simultaneously, with an extraction height in each slice of 2.0 - 2.2m. To extract lower slices, an artificial roof was created by leaving a layer of coal seam with a thickness of 0.4 to 1.0m between slices (Cu, 1990; Du, 2003). Although the application of the multi-slice longwall method improved the rate of coal recovery rate compared to the single pass longwall, the productivity of this mining method was low, due to some difficulties. One of the operational difficulties for the multi-slicing longwall method was the extraction of the lower slices (Cu, 1990). The main problem of extracting the lower slices was controlling the stability of the roof. The artificial coal layer left between the slices was easily fractured by the weight of the broken rockmass from the goaf of the upper slices. As a result, if the artificial coal layer is too thin, more work is required for stabilizing and controlling the potential risks of roof fall in the lower longwall face, especially where the face of the lower slices was operating under the goaf area of the upper slices. In addition, the instability of the roof may cause a potential risk to safety. Leaving a thick artificial coal layer between slices can prevent the potential risk of the broken rockmass falling from the goaf area of the upper slices, but it causes a larger coal loss. The high variation in seam thickness in the Quangninh coalfield was another 25 factor affecting the efficiency of the multi-slice longwall method. For the coal seams which are just thick enough for two slices, the high variation in seam thickness can cause difficulties in the extraction of the lower slices. For example, this problem has been reported in the Vangdanh coal mine (Cu, 1990), where the multi-slice longwall method was used to extract thick seams with an average thickness of 5.0m, but thickness in the seam ranged from 3.5 to 7.0m. For this case, the upper slice normally operates with an extraction height at 2.2m. During the extraction of the lower slice, however, in some areas where the seam is thinner than the average thickness, the thickness of the remaining bottom coal section was not thick enough for the height of the longwall face. As a result, the longwall face presented only two options: (1) cutting one part into the immediate floor to create a reasonable face height for passing through or (2) leaving the remaining coal of the lower section of the seam unmined and moving the face to an area where it is thick enough for the installation and extraction of the lower section of the thick seams. Both of these options can cause difficulties in longwall operations. Due to operational difficulties and the high cost of extracting the immediate floor, the option of leaving the remaining coal in thick seams unmined was usually chosen by the coal miners in Quangninh. Statistics indicate that due to the high variation in seam thickness, coal loss in the multi-slicing longwall method accounted for 48.6% in the Vangdanh Company (Cu, 1990). In order to remove the shortcomings of the multi-slice longwall, especially the high variation in seam thickness, the multi-slice longwall method with septum was undertaken in an experimental trial at the Vangdanh coal mine in 1987, and then in the Halam and Tanlap coal mines. In this method, a thick seam was extracted by two slices. The upper slice operated as the conventional longwall method, and wire-mesh was laid during the mining process in order to prevent the dilution caused by goaf from the upper slice mixing with caved coal in the lower slice. The lower slice extracted the lower section of the thick seam and drew a coal layer between the two slices during extraction. Although improvements in coal recovery were achieved by applying the multi- slice longwall with septum method, its overall productivity was limited. The 26 major operational difficulty of this method was the time consumed in laying wire mesh on the floor of the upper slice. Laying of wire mesh requires much time and labour because its procedures are undertaken manually. Therefore, the advance rate of the longwall face was slow, resulting in higher stress upon the supports, and more labour cost was needed to reinforce the supports for stabilization of the working place. In addition, because there was a long delay in the extraction between the two slices, the wire-mesh could not carry a high loading and was easily torn after such a long period of time, causing problems in preventing the dilution of the goaf from the upper slice into the coal layer of the lower slice (Cu, 1990; Du, 1999). Moreover, the multi-slice longwall method was only suitable for very thick coal seams. The extraction of the seams with a thickness ranging from 3.5m to 4.5m remained a challenge in the Quangninh coalfield. The Longwall Top Coal Caving (LTCC) method was first trialed in the Quangn- inh coalfield in the late 1990s and was based on mining experience from China. The LTCC method brought many benefits compared to the multi-slice longwall method, such as reducing both the operating costs due to more coal retrieved from the top coal section and per metre gateroad development. However, due to poor support equipment, which was mainly wooden props and cribs, mining operations faced difficulties in supporting workplaces under high ground pressure conditions. The inappropriate supports caused problems for the support of workplaces, a low advance rate and more frequency of face spall, resulting in low production and poor safety management. In addition, low coal recovery in total of a seam thickness is also a crucial issue for the LTCC method. For example, coal loss during the application of the LTCC method in the Halam coal mine was 43%, and in the Vangdanh coal mine was 43.4-52% (Du, 1996; Du, 1999). In an attempt to improve coal production and the recovery rate for the extraction of thick seams in the Quangninh coalfield, since 1997, hydraulic single props have gradually replaced the wooden props and cribs in the longwall faces (Du, 2003; Dac and Tuan, 2006). Since 2000, semi-mechanized shields have supported the faces in the LTCC method to strengthen support capacity. A detailed description of those supports is provided in Chapter 4. 27

The application of new mining equipment has improved the overall technical- economical parameters. Coal production in the longwall faces supported by hydraulic props or semi-mechanized shields has increased rapidly to 100,000 - 150,000 tonnes per year as compared with the production of 50,000 - 70,000 tonnes per year when the face was supported by wooden props and cribs (Du and Quang, 2005; Dac and Tuan, 2006). Coal productivity in the hydraulic single props increased to 2.5 - 5.37 tonnes per man-shift, average of 3.0 - 3.5 tonnes per man-shift. More importantly, the application of the hydraulic single props has reduced the consumption of wood in coal extraction. For example, after the application of hydraulic single props or semi-mechanized shields, the expenditure of wood per 1000 tonnes of coal production reduced from approximately 40m3 to 10.4 - 19.8m3 for the hydraulic single props and 2.0-8.3m3 for semi-mechanized shields (Du and Quang, 2005). Due to the improvement in coal production, productivity, and the standards of mine safety, wooden props and cribs were rapidly replaced by hydraulic supports as the main supports for the longwall faces. In 1998, for instance, there were only three longwall faces supported by hydraulic supports with a total coal product of 83,196 tonnes, accounting for 1.99% of overall underground coal production. Until 2005, 106 longwall faces supported by hydraulic supports with the total coal production of 8,369,000 tonnes accounting for 67.06% in total of coal production from underground mining operations. However, the output of the longwall faces remains low. The statistics of coal production from longwall faces in Vietnam gathered by Tuan (2005) show that the maximum output of a longwall face supported by semi-mechanized shields (model XDY-1T2/LY) was approximately 150,000 tonnes per year. This output is smaller than the average output of both the conventional longwall and the LTCC methods currently applied in many countries. In an attempt to improve the stability of the supports, as well as to increase the production and productivity of the LTCC method, in 2007, other types of semi- mechanized shields (model ZH1600/16/24Z and model GK/1600/16/24/HTD) made in China were trialed at the LTCC face of some underground coal mine 28 companies. The main difference of these types of support compared to the previous models is that their canopy is wide enough to prevent broken rockmass from falling into the face, and there is consequently no need to lay wire-mech during mining operations. As a result, much manual work has been saved. In addition, there is a system of roof beams attached to the shield canopies so that the problem of the supports tilting is reduced. Results from initial applications show that the production of an LTCC face supported by these types of support range from 200,000 to 250,000 tonnes per year (Du, 2009b). However, disadvantages still exist in this kind of support. The first is that it is only suitable for extracting coal from the face by drilling and blasting because there is no base for aiding the double-acting rams, and pushing the conveyor advance forward is very difficult. In addition, the advancement of the supports needs a large amount of time because each shield requires the digging of holes for the toe of the props. As a result, production is still limited. Moreover, there is no cover on the rear of the supports, so preventing broken rock in the goaf area from falling into the working face is a challenge. In order to further investigate the effectiveness of using mechanized mining equipment for the extraction of thick coal seams, in November 2007, another LTCC face, supported by fully mechanized equipment (mechanized shield mod- eled VINAALTA- 2.0/3.15, shearer, and chain conveyor), was also trialed to extract No.8 seam, Vangdanh Coal Company, which has an average thickness of 7.2m and a seam dip of 140 (Tuan, Du, Tuan and Ngan, 2008). The production of this longwall face reported approximately 450,000 tonnes per year (Du, 2009a). The initial results from the trial are encouraging, and they further confirm the importance of improving technology in underground mining in the Quangninh coalfield. However, this longwall face is limited by flat and gently dipping coal seams, and there is a need to study mining equipment suitable for longwall faces with steeply dipping conditions. A detailed description of the above mentioned supports is provided in Chapter 4. 29

2.3.2 Demands for coal production and the availability of coal reserves in Vietnam

The analysis of Industry Investment Consulting Company (2005), based on the forecast of the rate of economic development in Vietnam to the year 2025, estimates the demand of domestic coal production to the year 2010 at 29 - 32 million tonnes; the demand in 2015 will be 47 - 51 million, in 2020, 71.5 - 75.5 million, and in 2025 112 - 118 million tonnes. This means that the demand for coal production for domestic use during the period from 2011 to 2020 will increase by an average of 8 - 9% per year, and from 2021 to 2025, by an average of 9.0 - 10.5%. The aim of the Vietnamese mining industry is to increase its coal production in order to meet the demand of the domestic economy. According to the long term development plan of the Vietnamese mining industry, coal production will increase to 46 - 50 million tonnes in 2010, 50 - 55 million tonnes in 2015, 57 - 63 million tonnes in 2020, and up to 59 - 66 million tonnes in 2025 (Industry Investment Consulting Company, 2005). The statistics of coal resources in different locations in Vietnam indicate that coal is mainly located in two coalfields: Quangninh and Binhminh-KhoaiChau. According to Industry Investment Consulting Company (2005), from the total of 6.0 billion tonnes of coal reserves investigated to a depth of -300m compared to sea level (which is the main target of the mining operations of the Vietnamese mining industry in the near future), 4.0 billion tonnes come from the Quangninh coalfield, 1.6 billion tonnes come from the Binhminh-Khoaichau coalfield, and 0.4 billion tonnes come from the other locations. Therefore, coal resources in the Quangninh and Binhminh-KhoaiChau coalfields will be the main target in the next decades. As 95% of the current coal production in Vietnam comes from the Quangn- inh coalfield, it is estimated that coal output will mainly come from that area in the coming year. It is planned, by the Vietnam Coal and Mineral Resource Group, that from now to 2015, the most suitable mining methods will be chosen and planned for application in the Binhminh-Khoaichau coalfield so that the first 30 tonne of coal product will be available in 2015. In the year 2020, coal production in Binhminh-Khoaichau will increase to 5 - 10 million tonnes per year. In order to achieve this target, coal production, in 2020, from the Quangninh coalfield will have to increase by 54 - 56 million tonnes per year (Industry Investment Con- sulting Company, 2005). As a result, technological improvement in underground mining in the Quangninh coalfield is essential.

2.3.3 Discussion

The fundamental nature of mining technology in Vietnam, generally, is still at a low level of mechanization. All processes of mining, such as coal extracting, supporting, and transporting are mainly undertaken manually. Coal at the faces is drilled by handle drillers. After blasting, coal is raked into the chain conveyor by the workers, and then it is hauled along the longwall face, transferred to the wagons or belt conveyor and transported to the processing plants. As a result, coal production from underground mining operations still remains low. Analysis of the current status of underground mining operation and geological conditions in the Quangninh coalfield shows that thick seam mining is mainly carried out by manual methods using low technology. Although there has been some improvement in applying semi-mechanized mining equipment to support the longwall face, the essence of the mining operations is still technologically limited. Coal extraction from the face is usually undertaken by the drilling and blasting method, and broken coal is uploaded to the chain-conveyor manually, requiring a large workforce. In addition, the use of low mechanized support equipment, such as single hydraulic props or simple hydraulic shield also requires a large workforce and is time consuming. Statistics from Tuan (2005), and Dac and Tuan (2006) show that during the period from 2002 to 2005, coal production from underground mining increased from 6.1 million tonnes to 12.5 million tonnes, equivalent to an average increase rate of 26.2% per year. Yet, in order to achieve that rate of development, the underground coal mining companies in Quangninh have also increased their longwall faces from 134 in 2001 to 185 in 2005. This indicates 31 that underground coal production increased not only because of improvements in mining equipment, but also by increasing the number of working faces. The most recent improvement in mining equipment for the longwall face has achieved impressive results, but it is limited to flat and gently dipping seams. As a result, these improvements in mining equipment in Quangninh generally have a limited effect in increasing coal production and productivity. In the coming years, the demand for coal in the domestic market in Vietnam will increase rapidly, requiring a similar increase in coal production. Nevertheless, under the pressure of environmental protection in the area, especially maintaining the natural landscape of Halong bay nearby, open pit mining operations will have to be gradually scaled down and will end by 2014 (Anh, 2008). As a result, improvements in underground mining are crucial in order to both increase coal production and improve the standard of safety management in underground mining in the Quangninh coalfield. It is obvious that improving the technology for thick seam mining is essential in Quangninh. In the last two decades, many attempts have been made to suggest improvements in underground thick seam mining. Du (2003) researched ways of reducing coal loss in the sub-level mining and the LTCC methods. The study first discussed reasons for coal loss in thick seam mining, and then gave suggestions for reducing coal loss, including the reduction of the coal pillar between longwall panels, and the creation of an artificial roof for preventing coal dilution by laying wire mesh on the floor of the upper slice during mining operations in the sub-level caving method. In his study, Du (2003) also investigated the optimum coal seam thickness at which coal recovery, in the LTCC method, is maximal. While this research comprehensively discussed the status of coal loss in thick seam mining in the Quangninh coalfield, it only concentrated on current mining operations in Quangninh, which are based on simple mining equipment and technology. It did not investigate the applicability of mechanized mining equipment and advanced technology for extracting thick coal seams. In addition, the effect of seam dip and the properties of rockmass/coal on the application of mining methods were also not addressed. 32

In order to meet the increasing demand of coal products for the Vietnamese economy by underground methods, the planners and researchers in Vietnam (Industry Investment Consulting Company, 2005; Dac and Tuan, 2006) all agree that underground mining must be improved through choosing the most suitable mining methods, coupled with the appropriate mechanized mining equipment to match Quangninh’s geotechnical conditions. In this area, research on applying mechanized mining equipment for all processes of mining production, including coal extraction, transportation, and coal preparation, adaptable to the current overall infrastructure in Quangninh, is the most important factor in increasing coal production by underground mining. Currently, the Vietnamese mining industry is focusing on researching the application of mechanized mining equipment for all processes of mining production, including coal extraction, transportation, and coal preparation. One fully mechanized LTCC face has been trialed to extract a thick coal seam in the Vangdanh Coal Company, but it is limited to flat and gently dipping conditions. Therefore, the mining challenge of inclined thick seams in Quangninh remains for the future. For that reason, underground mining methods, coupled with appropriate mining equipment that enable the economic extraction of thick seams must be developed. These future methods must provide a higher coal production, productivity, and a safe working environment for all employees. In addition, they must increase the level of coal recovery in the mining panel and reduce the operational difficulties seen at the mining face. Therefore, it is necessary to improve the understanding of both operational and geomechanical issues associated with the extraction of thick seams in the Quangninh coalfield.

2.4 Conclusions

The Quangninh coalﬁeld, located approximately 150km North-Eastern of Hanoi, Vietnam, holds a large proportion of the coal resources in Vietnam. Coal resources in the Quangninh coalﬁeld to a depth of -300m (compared to sea level) are estimated to be 4.0 billion tonnes, and at a depth greater than the level of 33

-300m (compared to sea level) are estimated at 6.9 billion tonnes. Coal resources in Quangninh, which will be the main target of coal production in the next decade, are considered to have complicated geo-mining conditions. Many coal seams are thick (greater than 3.5m) and are steeply dipping. Detailed analysis of coal reserves indicates that a large proportion of coal reserves (56.7%) are thick seams, and 58% of those thick seams have a seam dip range from 15 to 350. 78.7% of thick coal seams are located under a moderate or deep cover depth (greater than or equal to 200m), and most thick coal seams (88.4%) have low or moderate gas contents (less than 10m3 gas per tonne of coal). The thick coal seams have high variations in thickness. In addition, a proportion of thick seam reserves has a short length along the strike (less than 1000m) due to the appearances of many faults in the coalfield which may influence the application of mechanized mining equipment. Over the last two decades, many attempts have been made to extract the the inclined thick seams in the Quangninh coalfield. Since 1990, underground mining methods, such as the multi-slice longwall (both with and without septum), and the LTCC method have been trialed in coal mines with the aim of increasing coal production, productivity, and the recovery rate. The application of those methods, combined with improvements in both technology and management has resulted in a dramatical increase in coal production from 9.4 million tonnes in 1994 to nearly 35 million in 2005. Although there is some improvement in coal production, productivity and coal recovery, the outcome of these methods is still limited in both capacity and safety management. The increase in total coal production is mainly based on the increase in the number of longwall faces. In addition, the maximum output of the longwall face is low when compared to the longwall faces in developed countries. The main reasons for the limits in coal production and safety management are inappropriate mining equipment and a lack in the understanding of the geotechnical and operational issues associated with the extraction of coal seams with steeply dipping conditions. In the coming year, the demand for coal products in the Vietnamese economy will increase rapidly, which requires an improvement in the mining industry. It is 34 estimated that the demand for coal products for the domestic economy in 2010 will be 29 - 32 million tonnes, but this will increase dramatically to 112 - 118 million tonnes in 2025. Due to the pressure of the environmental protection of Halong bay and the community nearby, open-pit mining will be gradually scaled down and will end in 2014. As a result, there is a need to improve the underground mining sector, especially in underground thick seam mining. In order to ultimately overcome the operational problems and realise the potential revenue from the thick seams in Quangninh, underground mining methods, combined with appropriate equipment that enables economical, high production, with a safe working environment for the extraction of thick seams, must be developed. These future methods must also increase the level of coal recovery in the mining panel and, most importantly, must be adaptable to the current infrastructure of the Quangninh mining industry. In parallel with these goals, it is necessary to improve the understanding of both operational and geomechanical issues associated with the extraction of thick seams in the Quangninh coalfield. This is the primary goal of this study. Chapter 3

Potential Underground Thick Seam Mining in the Quangninh Coalﬁeld

3.1 Introduction

As discussed in Chapter 2, thick coal seam extraction by underground mining methods has long been one of the challenging topics for mining planners and researchers in Vietnam. Over the years, a number of studies investigated underground thick seam mining to identify suitable methods of extracting the inclined thick seams of the Quangninh coalﬁeld, Vietnam, for the maximum level of coal recovery. These include Cu (1990) Du (2003), Tuan et al. (2004), and Du (2009a). However, at this point in time, the problems associated with thick seam mining have not been solved to a satisfactory level. The demand for greater coal production, the recovery rate, and a higher standard of safety management require improvements in the technology used for underground mining, especially for thick seam mining. As a result, there is a need to examine and review the advantages and disadvantages of thick seam mining technologies which have been, or are being, applied around the world, and to ﬁnd the most reasonable way to solve the problems related to thick seam mining in Quangninh.

35 36

This chapter is aimed at investigating the most suitable underground thick seam methods applicable in the Quangninh coalfield by evaluating the advantages and disadvantages, coupled with the applicable geological/geotechnical conditions of the methods which were, or are currently being, used to extract thick coal seams around the world. This study focuses on coal seams with a thickness greater than 3.5 m, since this is the lower limit for the definition of thick seams under the current research study, and seam dips ranging from 150 - 350. To achieve these objectives, a brief literature review of thick seam mining methods which have been, or currently are, used in coal mines around the world will be undertaken. Then, based on the reviewed mining methods, coupled with the geological/geotechnical conditions in the Quangninh coalfield and the current status of the mining situation in Vietnam, the most applicable mining method for the extraction of inclined thick seams in the Quangninh coalfield will be proposed to satisfy the Vietnamese situation.

3.2 Literature review

Numerous research projects carried out over the past two decades make it evident that thick seam mining operations are a challenge in various geo-mining conditions around the world. Seedsman (1990), and Hebblewhite, Roberts and Oldroyd (1997) researched the potential application and related issues of the High reach Single Pass Longwall technique in the extraction of thick coal seams in Australia, and Sarkar and Chatterjee (1992) did this for Indian mining conditions. The feasibility of the application of the Longwall Top Coal Caving method was studied by Hebblewhite, Simonis and Cai (2002), and Xu (2004). A literature review of international thick seam mining practice undertaken by Ghose (1984) and Cai (1997) identified that at different times, the following methods were mainly used to extract flat and inclined thick seams:

1. Room and Pillar method (or modiﬁcation of this method);

2. High reach Single Pass Longwall; 37

3. The multi-slicing Longwall Method;

4. Hydraulic Mining, and;

5. The Longwall Top Coal Caving method.

The above mining methods have been continuously optimized in order to be suitable under diﬀerent geo-mining conditions. These improvements are in con- junction with developing mining equipment and technology required for those mining methods to increase production capacity, productivity and safety management, as well as reducing coal loss. The following sections brieﬂy describe the above mentioned methods and the applicable conditions of each. Details of the advantages and disadvantages of these mining methods will be discussed in Section 3.3.

3.2.1 The Room and Pillar method

The room and pillar method, also known as the bord and pillar method, is a non- longwall method. This method is developed by driving the drifts from the main roadways to the block of a coal seam forming the rooms. Pillars are created by leaving the coal unmined between the rooms. Depending on the characteristics of the coal and the rockmass above the seam, some proportion of coal from the pillars will be extracted, and the remaining coal pillars will be left unmined for roof support and controlling strata movement. A diagram of the room and pillar method is presented in Figure 3.1. This method is widely applied in areas where the surface needs to be protected, such as coal seams located under residential areas, rivers or roadways. For example, this method was used to extract a 7 metre thick seam with a maximum working thickness of 6.1 metres in Westfalen colliery, Redbank via Ipswich in South Queensland, Australia (Edger, 1976). According to Edger (1976), the room and pillar method was chosen due to the constraints on capital development and for the prevention of surface subsidence. Because the mining area was located under a residential area, coal recovery was limited to a 40% extraction, and it 38 was decided to form pillars and not extract them. The method was developed by driving the roadways to create the rooms at the top section of the coal seam, the bottom section of the thick coal seams was then mined out by machine miners, as illustrated in Figure 3.2.

Figure 3.1: Typical layout of the Room and Pillar method (after Stefanko, 1983)

Figure 3.2: Typical cross sections of roadway extraction at Westfalen colliery (after Edger, 1976)

The room and pillar method was also applied to extract the Greta seam, which has a thickness of 7 - 10 m, and a dip of 1 in 12, in the central area of the South 39

Maitland coalfield, New South Wales, Australia (Harrison, 1976). The roadways of the room and pillar method were driven in the lower section of the seam with an extraction height up to 3m. The top coal was mined either in a secondary retreat operation or concurrently with pillar extraction. According to Harrison (1976), in some areas, the first workings were driven to seam partings as high as 5.5m from floor, but rarely removed the very top portion of the seam as a result of mining from the floor. The room and pillar method is also applied in areas where longwalling methods face problems due to the difficulty of cavability of the top coal and/or the overlying strata. In India, for example, only 3.8 million tonnes in the total production of 60 - 70 million tonnes from underground was from longwalling, and the balance was from conventional room and pillar mining (Dhar and Singh, 1992). In the past, coal seams in India, with a thickness of up to 8.5m, were extracted in one lift, but this had to be abandoned due to the difficulty of roof support (Singh, 1992). Multi-lift room and pillar mining in ascending order, with hydraulic sand stowing, prevailed in the extraction of No.14 seam with a thickness of 8.4m in the Tisco collieries of Jharia coalfield, India (Trehan, 1992). In this method of extraction, a seam was divided into two lifts. Pillars were split by driving 6m wide split galleries, 2 level and 2 dip splits. The extraction height at the bottom and top sections was limited to 4.8m and 3.6m, respectively. The mined out area was stowed with hydraulic sand. Multi-lift room and pillar mining with hydraulic sand stowing was also practiced to extract the King seam, Singareni Collieries Company Limited (Vyas and Benjamin, 1992). According to Vyas and Benjamin (1992), this method enables the extraction of up to 60 - 65% of the coal seam. The room and pillar method was also practiced to extract the 4 - 4.5m thick seam with a cover depth of 150 - 200m at the Aumance mine in France, with a recovery rate of 45% (Gouiloux, 1980). Coal extraction included driving rooms with 5m wide spacing at 10m intervals in the 75m x 15m pillars. When the room had been driven, the pillar was then reduced. The final shape of the pillar is illustrated in Figure 3.3. 40

Figure 3.3: Room and Pillar mining in Aumance mine (after Gouiloux, 1980)

Another method, named the blasting gallery (BG) method which is considered a modified room and pillar method, has been practiced in French and Indian coal mines. The BG method operates with roadway development along the floor of the coal seam on the room and pillar pattern, followed by the final extraction of the seam in one pass, by driving drill holes from the developed galleries in ring patterns. The coal then will be broken by the blasting method. Blasted coal is then removed by load haul dump machines (LHD) (Saha, Jian and Misra, 1992). The difference between the two methods is that coal extraction from the pillar in the BG method is undertaken by the drill and blasting method, while in the conventional room and pillar method, roadway development and coal extraction are carried out by the continuous miners. As a result, the BG method can extract very thick seams by one pass, which is usually divided into slices when extracted by the continuous miners. A diagram of the BG method is presented in Figure 3.4. The BG method was applied for the extraction of thick (7 to 12m) and gently dipping seams in India. It was used to extract the thick seam at Chora 10 Pit Colliery (Samanta, 1997). The seam had a thickness of 7.5 - 8.0m, a seam dip of 1 in 12 (equivalent to 50), and a cover depth of 80 - 90m. The mining equipment used included load haul dumpers, jumbo drills and hydraulic props. The target production of this method was 800 tonnes per day (Samanta, 1997). The BG method was also introduced to extract a 7.5m thick seam at Katras Colliery with an average production of 354 - 391 tonnes per day, and a recovery rate of 55 - 41

65% (Jha and Seam, 1992).

a, Blasting Gallery district plan b, Diagram of loading operations

c, Support principles of gateroads d, Drilling pattern

Figure 3.4: Diagram of the blasting gallery method (after Vyas and Benjamin, 1992)

3.2.2 High reach Single Pass Longwall method

The High reach Single Pass Longwall (HSPL) method operates similarly to the conventional longwall method, the diﬀerence being the use of higher loader capacity supports, combined with a shearer, chain conveyor and auxiliary equipment (Cai, 1997). In principle, the longwall face is created by driving a gateroad that connects the headings driven in parallel (known as maingate and tailgate), with the longwall face extending from the maingate to the tailgate. The block of in-situ coal formed by the parallel headings, the main roadways, and the longwall face 42 is known as a longwall panel. The length of the longwall panel is the same as the length of the maingate and tailgate headings. The longwall face is supported by hydraulic shields, with the function of providing a safe working space for personnel and equipment during the processes of coal extraction. The rockmass layers above the coal seam behind the supports are allowed to collapse or stowed by other materials. Figure 3.5 shows a typical arrangement of a modern longwall face.

Figure 3.5: Perspective and close up view of a High reach Single Pass Longwall face (after Hebblewhite et al., 2002)

The HSPL method used for extracting thick coal seams has been applied by many coal producers around the world. Currently in Australia, for example, there are a number of mines operating longwall faces using the HSPL method with an extracted height of around 4.5 - 5m, including West Wallsend, Dartbrook, New- stan, North Goonyella, Oaky North and Moranbah North (Hebblewhite, 1999; Hebblewhite et al., 2002). At the West Wallsend colliery in the Newcastle coal- ﬁelds, for instance, the HSPL method has been applied to extract the Borehole seam, which was a ﬂat, up to 6m thick coal seam with an extraction height of up to 5m (Hamilton, 1999). This mining method replaced the traditional longwall method with an extract height of 2.2 - 2.4m. There were no reports that detailed output, but according to Reid (1997), the average production of longwall faces in Australia for extracting coal seams with a thickness of more than 4m was 3,572 43 million tonnes per year, with a maximum production of 5,072 million tonnes per year. In the USA, the HSPL method has been applied to extract the Fruitland No.8 seam in New Mexico and the upper Hiawatha seam in the Utah mine with a maximum thickness up to 156 inches (3.96m) (Fiscor, 2006). At Westfalen colliery in Germany, the HSPL method has been applied to extract a double seam prasident/helene with several band layers in the coal seam. The extraction height varies between 3.8 and 5.0m, the seam dip at 30 - 50, the longwall face length is approximately 250m, and the average output is 3500 - 4000 tonnes per day (Bussmann and Schroth, 1992). At the Novokuznetskaya mine in Russia, a coal seam 5m thick and dipping at 250 has also been extracted by the HSPL method (Reid, 1997). The application of the HSPL method has developed rapidly in the last decades in China. This method for extracting thick coal seams was used as early as the 1980s. For instance, A HSPL face with a maximum extraction height of 4.5m at No.2 coal seam, Dongpan mine in Xiangtai Coal Mining Bureau, was reported in 1985 to have a maximum output of 120,000 tonnes per month (Xu, 2004).

3.2.3 Multi-slice Longwall method

The multi-slice longwall (MSL) method, also known as “multi-lift” or “multi- pass” longwall mining (Ghose, 1984), divides a thick coal seam into two or more thinner slices; each slice is extracted by the conventional longwall method with a working height of 2 - 3m. The top slice of the coal seam can be extracted ﬁrst, then the lower slice (descending slicing), or alternatively, coal extraction can start by mining the bottom slice of a seam and then moving to the top (ascending slicing). Slices can be mined in sequence where there is a time delay between the extraction of slices, or the coal seam can be extracted in a number of descending, simultaneous slices. When the seam is gently inclined, the slice is taken parallel to the ﬂoor, but with more steeply dipping seams or in seams with an irregular outline, horizontal slices are taken (Ghose, 1984). The mined 44 out space behind the face in the slices can be allowed to collapse or be stowed by other materials. Figure 3.6 illustrates how the MSL method is applied to a thick seam, in this instance the seam is divided into three slices.

Figure 3.6: Principle of the Multi Slice Longwall method in three slices (after Hebblewhite et al., 2002)

When the coal seam is extracted in a descending order, the floor can be in-situ coal or rockmass, and the roof of the lower slices can be naturally consolidated goaf or stowing materials. In the multi-slice longwall method with descending sequences, the roof is commonly controlled by naturally consolidated goaf. This method of roof control was the most popular method used in China for the extraction of more than 5m thick seams in the 1980s and early 1990s, due to its well established mining technology and the high performance of equipment used in medium thick seams (Cai, 1992). Over 20 coal mining administrations in China introduced mechanized longwall faces using the multi-slice method with descending order. The slices were extracted at 2.5 - 3.5m in height. Figure 3.7 illustrates the multi-slice longwall method in descending order with the roof controlled by naturally consolidated goaf. The naturally consolidated goaf can be used as the roof for the lower slices when the goaf of the upper slice can self strengthen after a period of time. If the goaf of the upper slice is not consolidated, an artificial roof for the lower slice must always be constructed in the following ways: (1) the laying of wire mesh, bamboo mats, or wood timber, or (2) the goaf may be strengthened by the addition of 45 cement materials, or a combination of (1) and (2), which are normally constructed on the floor during upper slice mining operations. The advantage of creating an artificial roof is that it improves the stability of the lower slice working space and reduces the dilution of coal quality. However, this process is time consuming and also increases the operating cost by the need for both additional materials (wire mesh, bamboo, timber or cement) and labour.

Figure 3.7: The multi-slice longwall method separated between slices by a layer of coal or stone parting (after Dorling, 1980)

An easier and cheaper way of creating the roof over the lower slice in the multi-slice longwall method is to leave a layer of coal or a stone parting between the slices. This method of the extraction has been deployed in the Kostenko mine, in the former Soviet Union, to extract a thick seam varying from 7m to 7.5m in thickness, and dipping at 80 (Dorling, 1980). The seam was divided into 2 slices, each slice has an extraction height of 3.2m. A layer of coal, with a thickness of 0.5 - 0.8 m, was left between the two slices to form the roof of the lower slice (Figure 3.7). This method can eliminate the cost of creating an artificial roof for the lower slice, but it reduces coal recovery and can potentially cause spontaneous combustion due to a large amount of coal left in the goaf. In addition, if the coal layer is not strong enough, it may fracture and collapse, causing the broken rockmass in the goaf area of the upper slice to fall into the working face of the lower slice. These problems certainly affect the advance rate 46 and productivity of mining operations in the lower slice as well as potentially cause injury to personnel. Moreover, if the roof of the lower slice collapses, this could cause a hazardous cavity to form and extend over the top slice and possibly even extend forwards as well as upwards. In descending slicing sequences with backfill, the roof slice is mined out first and the mined out space is stowed with materials known as backfill. The backfill materials, containing binding agents, create an artificial roof for the lower slice, and the mining of the next slice is carried out under that artificial roof. The backfill materials in this case must be strong enough to make the mining under them possible (Palarski, 1999). The principle of the descending slicing method with backfill is illustrated in Figure 3.8.

Figure 3.8: Descending multi-slice longwalling with backﬁll method (after Palarski, 1999)

Ascending slicing sequence is another way of multi-slice longwalling. In this method of extraction, the floor of the upper slices is the backfill materials stowed in the mined out lower slice. The roof is the in-situ coal or rockmass, as illustrated in Figure 3.9. This method is practiced widely in Poland to extract thick coal seams which are located below industrial and residential areas (Palarski, 1999). The thick seam is divided into slices, each slice is 2 - 3m thick, and the direction of the extraction is up-dip (Figure 3.9a) or down-dip (Firgure 3.9b). The goaf area is grouted with hydraulic backfills. 47

Figure 3.9: Ascending multi-slice longwalling with backﬁll method (after Palarski, 1999)

Stowing backfills in the mined area has some advantages. The first is that it restricts the surface subsidence. The second is that the space where coal was mined out is replaced by stowing materials, so it reduces the fracture of the rockmass above the coal seam, resulting in a decrease in surface subsidence and the potential danger of hazardous rock bursts. In addition, when the mined area is filled by stowing materials, the interaction between the air and the faces of coal will decrease, resulting in the prevention of potential fire or spontaneous combustion. However, the major disadvantage of this method is that the cost of the mining operation increases due to the extra expense of making and transporting stowing materials, as well as the maintenance of fills and drainages (if the stowing material is hydraulic backfills). In addition, it is difficult for heavy mining equipment to operate on the backfill materials when extracting slices in ascending sequences, compared with the in situ coal or rockmass.

3.2.4 Hydraulic mining method

Hydraulic mining is the use of high pressure water to break and transport coal from the seams. In this mining system, coal at the face is broken by hydraulic means. The broken coal, mixed with water, is transported hydraulically in ﬂumes of suitable dip by gravity. After wet-screening, the coarse coal is carried out by conveyors or locomotives while the ﬁne coal is pumped to the surface as a slurry (Jha and Seam, 1992). High pressure water is created by pumps, carried through 48 a network of pipes and applied to break out the coal through a monitor. A schematic diagram of hydraulic mining is presented in Figure 3.10.

Figure 3.10: Concept of Hydraulic Mining (after Nishioka, 2000)

Hydraulic mining has been deployed in different countries to extract thick coal seams in complicated geo-mining conditions, such as, when the seams have a high gradient or when the seams are separated into pieces by many faults where mechanized methods have little or no application. For example, this method was operated to extract a 3.7 to 4.1m thick seam, dipping at 8 - 100, at Yubilenaya Mine in Kuznetsk, Russia. The output from hydraulic mining reached 1 - 1.5 million tonnes per year, with an average daily output of 4,886 tonnes and a monthly labour productivity of 2,240 tonnes per worker (Mills, 1978). Experience from the application of the hydraulic method in the Kuzbass coalfield, where the geological conditions on the site were complicated, shows that the efficiency was 3.5 to 4 times better than other mining methods under similar conditions (Kolesnikov, 1980). A trial application of hydraulic mining was also carried out at Gopalichak colliery, Jharia coalfield, India, for extracting the X seam with a thickness of 6.0 - 7.8m, and dipping at 7 - 80 (Singh, 1992). In this system of extraction, the coal was cut by a 100 bar pressure jet and was flushed out at 20 bar pressure. The slurry was transported to the slurry pumping station. After screening, +4mm or greater sized coal was transported in tubs and the slurry containing -4mm sized coal was pumped to surface. At the surface, the slurry was dewatered in a 49 clearing pond. Hydraulic mining was applied to extract the 50 foot thick Balmer seam with an average dip of 350 (ranging from 250 to 550) at Sparwood, British Columbia, Canada (Patching, 1979). According to Patching (1979), while the geo-mining conditions at Sparwood were not suitable for the conventional mining methods, the hydraulic mining method exhibited many attractive features. Hydraulic mining achieved productivity of 18 - 20 tonnes per man-shift for all underground and surface crews, and a recovery rate of 60 - 70%. Patching (1979) therefore concluded that hydraulic mining was better than other underground mining methods in Sparwood’s geo-mining conditions. The hydraulic mining method has also been trialed in some parts of China. According to Tian, Chen and Liu (1979), one monitor produced 347 tonnes per hour in Yanzhuang Mine, its monthly output was 30,000 - 60,000 tonnes, and its annual capacity was between 400,000 and 500,000 tonnes. In 1977, the maximum production capacity of hydraulic mining reached 1,170,000 tonnes at Lujiatuo Mine (Tian et al., 1979).

3.2.5 Longwall Top Coal Caving method

The Longwall Top Coal Caving (LTCC) method is basically a combination of the traditional longwall method and the caving method of top coal during the process of mining operations. The terminology of “Longwall Top Coal Caving” is confusing in diﬀerent countries due to the development and improvement of this mining method at diﬀerent times. The original name of the LTCC method was the “Soutirage” method (Nath, 1979; Proust, 1979). The other names of this mining method were “Single Sub-level caving” (Callier, 1972), “the basic sub-level caving” (Bewick, 1983), or the “sub-level caving method” (Adam, 1976; Hams, 1976). In the LTCC method, the panel is created by a set of headings (gateroads), the same as in the conventional longwall panel (usually retreating longwall method), but the roadways are driven in the lower section of the coal seam. The shields 50 with caving function, coupled with other equipment are then installed to create the longwall face. During mining operations, the lower section of the coal seam is extracted by the conventional longwall method. The remaining upper section of the coal seam will be drawn through the window of the shields after caving behind the supports in parallel with the process of longwall operations. Figure 3.11 illustrates the principle of the LTCC method.

Figure 3.11: Conceptual Model of LTCC System (after Xu, 2001)

The LTCC method was first deployed approximately 40 years ago in France and Yugoslavia, and then applied in other countries, such as Poland, Turkey, and China, to extract thick coal seams. This method was initially trialed at Blanzy coal mine, Centre Midi coalfields, France in the early 1960s to extract irregular coal seams with a thickness of up to 20m and a gradient up to 300 (Nath, 1979). Proust (1979) stated that the average daily output, in 1977, to extract an average 6.5m thick coal seam under very difficult seam conditions (seams dip exceeding 300 and up to 450 locally) was 1350 tonnes/day. The rate of recovery was 75%, and productivity was 13.6 tonnes per man-shift. The LTCC method has also been applied in the Velenje mine, Yugoslavia since 1953, and was referred to as the “Velenje mining method”. It has been trialed widely at other collieries for extracting thick (up to 20m or more) and steeply dipping (300 - 600) seams. In 1977, this mining method accounted for approximately 53% of total underground coal production in Yugoslavia (Ahcan, 1979). During the 1970s, it was also used in other countries, such as Hungary, 51

Romania and the former Czechoslovakia (Tien, 1998). The LTCC method was first introduced in China in 1982, and then developed rapidly over the next two decades. The first successful LTCC face was achieved at face No.8603 of Yangquan coal mine with an output of 140,000 tonnes per month in 1990, and a recovery ratio at the working face of over 80% (Jian, Xianrui and Yaodong, 1999). Currently, production from the LTCC method accounts for nearly 10% of China’s underground production (Tien, 1998). By improving technology, mining equipment and operation management, significant development and an impressive performance has been achieved in China with the application of this mining method. After 5 years of applying the LTCC method to extract the 5.6 - 6.5m thick coal seams with a gradient of 3 - 80 in Dontan mine, Yazhou coalfield, the maximum monthly output increased from 151,786 tonnes in 1994 to 501,068 tonnes in 1999, and annual productivity increased from 2,821 tonnes to 14,306 tonnes per man. The maximum annual production of a face has reached up to 5.1 million tonnes per year (Yingdi, Tianzhi, Guishan and Weiqing, 1999). Now the LTCC method is being studied for deployment in more difficult conditions. Currently, the LTCC method has been trialed in China to extract thick coal seams, dipping up to 450. Output from one mining face was 97,100 tonnes per month, productivity was 45 tonnes per man-shift, and the recovery rate was 82.27% (Xie, Gao and Shangguan, 2005). For the extraction of very thick seams or thick seams which have a strong, competent roof, the LTCC method can be modified to a technique named “sub- level caving”. In this method, the thick seam is divided into two slices. The top slice is extracted by the conventional longwall method, and the bottom slice is extracted by the conventional longwall method combined with caving the remaining coal layer between the two slices. Another variation of the LTCC method in very thick coal seams is that both slices are extracted by the longwall method with the caving of the top coal. If a seam is too thick to be divided into two slices, three slices are used. In this case, the top slice can be extracted by the traditional longwall method, and the middle and the bottom slices can be mined by the longwall method combined with winning caving coal between slices. Another 52 principle of the LTCC method is illustrated in Figure 3.12.

Figure 3.12: The sub-level caving method (after Bewick, 1983)

The sub-level caving method has been widely practiced in the former Soviet Union, the former Yugoslavia, France, and China. An early application of the sub-level caving method was in the Kuzbass coalfield of the former Soviet Union to extract the flat and 6 - 12m thick coal seams with the help of flexible steel mats (Vorobjev, 1962). In this mining system, the thick seam was divided into two slices. A 1.5 - 1.8m thick slice was extracted by longwalling from the top of the coal seam. Flexible steel mats, which consist of interlacing steel strips, were laid on the floor of the top slice during mining operations. The bottom slice was extracted to a height of 2.5m, supported by KTU shields, advancing down the dip. The coal layer between the two slices was fractured by blasting and won through the windows of the shields. The sub-level caving method, as applied in Kuzbass coalfield, is illustrated in Figure 3.13. The sub-level caving method has also been deployed at Blanzy coal mine, France, to exploit its 9.9 - 14.8m thick coal seam, dipping at up to 260, with a maximum monthly output, in 1975, of 69,456 tonnes, and a productivity per man-shift of 15.7 tonnes (Adam, 1976). The sub-level caving method was also practiced widely in China to extract the flat (dip ranging from 00 to 250) and medium to steep (dip ranges from 260 to 450) thick seams during the 1970s (Tian et al., 1979). 53

Figure 3.13: The sub-level caving method, separate slices by ﬂexible wire mesh applied in Kuzbass coalﬁeld (after Vorobjev, 1962)

At Velenje mine in the former Yugoslavia, to extract the very thick coal seams, the sub-level caving method was further modiﬁed (Ahcan, 1979). In this case, the upper slice was extracted from the middle section of the thick seam instead of at the top of the seam. The function of this slice is to achieve preliminary fracturing of the top of the coal seam. All broken coal was won by caving at the bottom slice. According to Ahcan (1979), mining losses were largely reduced by using this method. A diagram of the sub-level caving method deployed at the Velenje mine is illustrated in Figure 3.14.

3.3 Potential thick seam mining methods applicable to Quangninh’s conditions

Historically, the extraction of thick seams in the Quangninh coalfield has been a matter of meeting the demand planned by the central government, while the other aspects of mining operations, such as profitability, and the coal recovery rate have not been paid reasonable consideration. As a result, a large amount of coal was left unmined, causing a high coal loss in thick seam mining operations. As the Vietnamese economy shifts to a market economy, beside operational and technical aspects, coal companies are now forced to think about how to operate mining activities, not only with high production but also with profitability. As minable 54

Figure 3.14: The LTCC method used in Velenje Mine (after Hebblewhite et al., 2002) deposits at shallow depth are depleted, coupled with the increasing demand for coal production and the pressure of scaling down open-pit mining production for environmental protection, there is a need for improvement in underground mining. This improvement should satisfy the high and sustainable levels of safety standards, production, productivity and recovery rate required in Vietnam. For this reason, the following sections will discuss in detail the suitability of each underground thick seam method described in Section 3.2, for the current situation in the Quangninh coalﬁeld. The selection criteria was based mainly on the evaluation of the suitability of each mining method for the geo-mining conditions of the mine sites, the current situation of the infrastructure, the level of mechanization and the other systems of mining production in the Quangninh coalﬁeld. In addition, equipment requirements, capital investment, the recovery rate, strata control, and overall economic considerations were also considered in the evaluation processes. 55

3.3.1 Room and Pillar and Blasting Gallery methods

The room and pillar and blasting gallery methods are considered to be suitable for extracting the thick coal seams in complicated geological conditions, as they have the following advantages:

• Low capital investment compared with longwall mining methods,

• High productivity per man-shift,

• More ﬂexibility in the application of mining equipment compared to the longwalling methods.

Therefore, these two methods are suitable for complicated geo-mining conditions. However, the room and pillar method was considered to have a limited application in the Quangninh coalﬁeld for the following reasons:

• The main challenge of this mining method is the low coal recovery rate, due to a large amount of coal left unmined for pillar creation. Operational experience has shown that coal recovery from this mining method is generally less than 50%. For example, the coal recovery rate to extract a 4 - 4.5m thick seam by this method at Aumance mine in France was 45% (Gouiloux, 1980), and 49% recovery was achieved from the production in the Bharat coal area in India (Singh, 1992). Coal extraction by the room and pillar method at Ipswich in Queensland was limited to approximately 40% (Edger, 1976). Experience in extracting the thick, No.14 seam in the Tisco collieries of Jharia coalﬁeld, India, by the room and pillar method yielded a recovery of 42% as against a theoretical recovery rate of 67% (Trehan, 1992).

• The other challenge of applying the room and pillar method in the Quangn- inh coalﬁeld is the current technology in roadway development which is being used in Quangninh. In the room and pillar method, a large number of gateroads need to be driven, and coal extracted from roadway development accounts for a large proportion of the total coal product. In order to get 56

a high output from the room and pillar method, the gateroads should be driven by mechanized mining equipment in order to achieve a high rate of advancement. However, the limitation in the Quangninh coalfield is that almost all gateroads in the underground coal mines in Quangninh are driven by the drilling and blasting method, coal is loaded on the chain conveyor or locomotive cars by manual means, and the face is supported by steel arches or timber beams. With that technology, the rate of advance from driving the gateroads in coal seams ranges from 120 - 150m per month (Du and Quang, 2005). Since 2004, with the application of roadheaders for driving the gateroads, the advance rate increased to 200 - 300m per month (2.6 - 4.0m/shift). This advance rate is still much lower when compared to the average advance rate in many countries (the average advance rate in Australia is 20m per shift (Roberts, 1998)). The main reason for this low advance rate is the time and labour consumed in transporting and erecting support materials, because all the gateroads are supported by arch steels or steel/wooden bars. In the case where the coal seams are so steeply dipping that the continuous miners can be used, manually driving the gateroads and transporting support materials will certainly increase the costs and time of roadway development. As a result, the low advance rate of roadway development will limit the efficiency and prevent the increase of coal production if the room and pillar method is applied in the Quangninh coalfield.

• The high depth of cover is another challenge to using the room and pillar method in the Quangninh coalfield. When the cover depth increases, the stress load upon the pillar will increase and cause difficulties and danger in mining operations. Experience of room and pillar extraction at some collieries in the South Maitland coalfield, New South Wales, Australia has shown that, down to the cover depth of 150m and provided their size is adequate, the pillar can be extracted by conventional methods. When the cover depth exceeds 150m, it has generally been found that if complete extraction with wide panels is attempted, as the goaf develops, loading on the 57

face becomes excessive, resulting in the driving of the first two pockets, in the pocket and fender system adopted for pillar and top extraction system, becoming extremely difficult. The high loading on coal pillars results in excessive pillar bumping, spalling and ribside roof guttering in the pocket. In addition, the closing of the roadways due to the collapse of top coal, floor heave and the failure of coal pillars, has been seen as a result of excessive pressure from the superincumbent strata with a further increase in the extraction of the pillar in the field (Harrison, 1976). As a result, higher coal loss will be anticipated. For that reason, the room and pillar method is not suitable for the extraction of coal seams at a high cover depth. As discussed in section 2.2.5, a large proportion of coal reserves (78.7%) are located at a cover depth of greater than 200m in total of the investigated coal reserves in the Quangninh coalfield. In addition, steeply dipping seams are also a constraint on the application of the room and pillar method for the extraction of the inclined thick coal seams in the Quangninh coalfield, because pillars in inclined seams are less stable than in flat seams. This means that coal recovery would decrease if the room and pillar method was applied in the Quangninh coalfield.

The blasting gallery (BG) method, modiﬁed from the room and pillar method, is also considered limited for extracting the inclined thick seams in Quangninh because of following reasons:

• The ﬁrst reason, as mentioned previously, is the large rate of roadway development per tonne of coal production of this method, and many gateroads are driven with steeply dipping conditions. In addition, with the same pattern of roadway development, the BG method is more eﬃcient in thick seams than in thinner ones. The challenge when applying this method in Quangninh is that its thick coal seams mainly range from 3.5m to less than 9m. Therefore, the rate of roadway development per tonne of coal production will be high. As a result, a low rate of roadway development would prevent an increase in coal production from this method. 58

• The second reason is the fact that most thick seams in Quangninh have a weak to moderately strong immediate roof, that may cause high coal loss from the BG method. In the past, the BG method has been applied in the Quangninh coalﬁeld to extract thick coal seams with a seam dip of more than 350 (Ngan, 2004). In this method of extraction, a gateroad is ﬁrst driven along the strike, then from that gateroad, inclined drifts are created with a length of 13 - 23m, and the distance between the drifts of 10 - 12m, forms the pillars, as illustrated in Figure 3.15. To extract the coal from the pillars, rings of blasting holes were drilled and blasted at the drifts. After the blasting operation, the broken coal was drawn by gravity with the help of mine workers.

Figure 3.15: The BG method used in Quangninh for steep coal seams (after Ngan, 2004)

Results from practice have shown that this method has a reputation for high productivity per man-shift. However, low coal recovery and a high rate of accidents are the constraints of this mining method. Statistics in many coal mines applying this method show that coal recovery is less than 60%. The reasons for this is that when extracting the coal seams with a weak to moderate strong immediate roof, a large amount of coal is blocked 59

by broken rock collapsed from the roof, or by props which were used for supporting the drifts before blasting, as illustrated in Figure 3.16. In some situations, a large volume of coal was still hanging in the goaf area. This problem may encourage mine workers to break safety management rules. In reality, many cases of death have been reported in the Quangninh coalﬁeld when the miners broke safety regulations and entered into the goaf area to drill one or more small blast holes in order to get some of the coal still hanging.

Figure 3.16: Common types of coal loss when using the BG method in the Quangninh coalﬁeld

For the reasons discussed above, this study recommends that the room and pillar and the BG methods should only be applied to extract thick coal seams which are under areas where the surface needs to be protected or for coal seams which have a strong rockmass that is diﬃcult to mine by other mining methods at the current stage of the Vietnamese mining industry.

3.3.2 High reach Single Pass Longwall method

It was revealed in literature review that, with the development of technology, the height of the shearer and the supports of the High reach Single Pass Longwall (HSPL) method have already gradually increased to 5m and above. Some of the advantages of the HSPL method are: 60

• A simple operation and ventilation system, high production and productivity, and high coal recovery rate.

• The low rate of gate road development per tonne of coal production. In the West Wallsend colliery, for instance, after the deployment of the HSPL method to replace the traditional longwall method, with an extraction height of 2.4m, progressing from an extraction height of 2.4m to maximum 5m in West Wallsend, led to a 40% reduction in roadway development (Hamilton, 1999).

Besides the obviously advantage of the HSPL method, it is estimated that this mining method would face some disadvantages when applied to the Quangninh coalﬁeld. They are:

• The short length of coal blocks in the seams. In the Quangninh coalfield, due to the appearance of many faults, the coal seams are separated into small coal blocks. Results from the reserve analysis of thick coal seams in Section 2.2.5 identify that a large amount of coal reserves in the Quangninh coalfield are in short length blocks (47.5% of thick seams have a block length of less than 1000m). These results indicate that frequent movements of the longwall face would be needed when applying longwall extraction to the Quangninh coalfield. The problem with the frequent movement of the longwall face in the HSPL method, is that the supports and the shearer in the HSPL method are larger and heavier than in conventional longwall equipment. Ryan and Fowler (1992) stated that the higher reach of the HSPL method required more support density to maintain support stability. In addition, to guarantee stability, the centre of gravity of the support must be maintained as low as possible. As a result, the width of the supports should be wider. Hence, the supports for the extraction of thick seams are heavier and more complex than those for medium and thin seams. Due to the greater height of extraction, the processes of moving the mining equipment from the face end of one panel to another would cost more time and labour than the face move of traditional longwalling. The high capital 61

investment cost of the HSPL method requires a high output in order to be cost effective. Hebblewhite (1999), based on experience in the application of the HSPL method in Australian mining conditions, stated that production from that method needs to be high in order to meet the requirements of capital investment. As a result, the face moves need to be minimized. Therefore, the frequency of face moves applied in the Quangninh coalfield could affect the cost effectiveness of the HSPL method.

• The stability issues of the supports due to the dipping factor. As noted by Ryan and Fowler (1992), to guarantee stability, the centre of gravity of the support should be maintained as low as possible. For support in the dipping seams, part of the force acting through the centre of gravity is a downdip slipping force and the other part is a restraining force at right angles to the seam ﬂoor. For that reason, in the case when the extraction heights are the same, the higher the seam dip, the more instable the supports, as illustrated in Figure 3.17. Therefore, for extracting inclined seams, the longwall with a high extraction face is certainly not suitable in terms of maintaining the stability of the supports.

Figure 3.17: The eﬀect of extraction height on stability of support

• The complications in transportation of the heavy roof supports and other equipment in the mine. It should be noted that this problem could easily be solved by a modern mining industry where the roadways are wide and high enough for transport of heavy equipment. However, in operating mines like 62

Quangninh, where the infrastructure has been developed before the consideration of deploying heavy and large mining equipment, the transportation of heavy equipment would be a big issue. Difficulties in the transportation of heavy and large sized mining equipment on narrow roadways, which were driven before the decision to purchase thick seam longwall equipment in the West Wallsend colliery, Australia, is an example. Hamilton (1999) reported the difficulties in hauling the supports and shearer when using the HSPL method at the West Wallsend colliery after the roadways had been driven. The roadways at the West Wallsend colliery were driven with a planned minimum height of 2.5m. For that clearance height, the roof supports, which have a transport height of 2.2m, can be easily transported. However, in some parts of the roadways where the minimum clearance decreases to 1.9m, the roof supports had to be laid on their side, and dragged on slides through the low sections before being stood upright and transported normally using an Eimco trailer. Roadway developments for underground coal mines in Quangninh were mostly driven by the drilling and blasting method, and supported by steel arches with a small face area, which ranged from 6 - 10m2. The transportation of coal and mining equipment in mines is mainly undertaken by wagons with a capacity of 3 - 5 tonnes, pulled by locomotives. As a result, It would be very difficult to transport the large, heavy supports and shearer on those narrow roadways in Quangninh. In addition, the installation of large, heavy supports in the inclined face when moving the face also constrains the use of the HSPL method in the inclined thick seams in the Quangninh coalfield.

• The diﬃculties in supporting the intersection between the face-end and gateroads. Gateroads usually are driven to meet the requirements of transportation and ventilation in the mine. Depending on the types of vehicles used in the mine, the width and height of some roadways may be diﬀer- ent from others. Gateroads in longwall panels in many underground coal mines are generally driven at a height of 2 - 3.5m. With a cutting height 63

on the longwall face of up to 5 or 6m, there will be a sudden step between the gateroads and the longwall face. The free face of the step that exceeds the height of the gateroads could cause spall on the longwall face, resulting in unsafe conditions for equipment and personnel working at the face or the roadways. In addition, the big step between the longwall face and the gateroads may cause potential dead spots in which ventilation is diﬃcult.

• The difficulties in controlling the potential danger of face spalling. Spalling, especially of high longwall faces, may cause many difficulties in mining operations, such as overloading the armoured face conveyor (AFC) or causing danger to the personnel working along the longwall face. Face spalling is caused by many reasons. The first is the effect of the abutment stress on the longwall face. The other cause of spalling is the inadequate resistance of the roof supports. If the roof supports are not strong enough to hold the roof above the longwall face, spalling will occur. Roof supports are fitted with flippers, which play an important role in resisting face spall and protecting the working area from spalling. However, even with support from flippers, spalling has been reported to frequently occur in some coal mines, and this problem obviously affects the productivity and safety management of mining operations (Hamilton, 1999).

Currently, the extraction height of the longwall faces in the Quangninh coalfield is approximately 2 - 2.5m. Therefore, the HSPL method is not considered a suitable method for the extraction of inclined thick seams under current situations in the Quangninh coalfield. It is recommended that the high reach of the single pass longwall method should be first focused on for seams having a thickness range from 2.2m to 3.5m. If the efficiency of mining production and capital investment is verified, then a further study will be undertaken for seams with a thicknesses of greater than 3.5m. 64

3.3.3 Multi-slice Longwall method

In the multi-slice Longwall (MSL) method, by dividing a thick seam into slices with a medium extraction height, mining equipment for medium thick coal seams can be used. As a result, reasonable capital investment and simple mining operations are the advantages of this method. In addition, in multi-slicing with descending sequences, the top slice can operate as the stress control slice, and the lower slices operate in shadow stress conditions. Therefore, rock bursts or a sudden collapse of the rock mass above the roof can be prevented. As a result, this method is suitable for those thick seams which need assistance in fracturing the immediate roof. However, the main disadvantage of applying the multi-slicing longwall method for extracting thick coal seams in Quangninh’s conditions is that many thick seams have a high variation in thickness. Analysis of the variation of seam thickness in the Quangninh coalﬁeld, undertaken by Thanh and Dac (1995), clearly shows that the thickness of coal seams in Quangninh vary widely. Their study investigated the variation of seam thickness using a variation ratio index which was created by the VNIMI institute in Russia. The variation index, Vm, is determined by equation 3.1, as follows:

σm Vm = · 100% (3.1) mtb where:

σm= the standard deviation, which is determined by equation 3.2,

s P (m − m )2 σ = i tb (3.2) m n − 1

mi = seam thickness at the investigated point, m,

mtb = the average thickness of the coal seams, m, n = number of investigated points.

The results from that study show that the variation ratio of coal seams in Quangninh ranges from 39 - 64%. One challenge of applying the MSL method to 65 extract thick seams with high variation in thickness is that it may either cause the loss of one or more slices when the seam thickness in some areas is thinner than the average thickness, or can cause coal loss in cases when the seam thickness is greater than the average thickness. Applying the MSL method to extract a thick seam with descending order is an example. As the upper slice operates with a normal extraction height, the lower slice may mine into the goaf area of the upper slices or into the immediate floor, if the coal seams get thinner. This problem can interrupt mining operations because either mining equipment has to be relocated to another place where the remaining of a thick seam is thick enough to set up a longwall face, or the face of the lower slice has to cut into the rockmass on the floor to retain the height of longwall face in that slice. Either of these situations would seriously affect the efficiency of mining operations. This problem has been occuring in some coal mines in Quangninh where the MSL method was used. In contrast, if the seam thickness is thicker than average, coal reserves of the portion which is thicker than average will be lost due to the fixed height of the longwall face in each slice. Another challenge when applying the MSL method in thick coal seams is the arrangement and maintenance of entry systems (the gate roads in different slices). In general, a set of gate roads (the maingate and the tailgate) are needed for each slice. In a descending order, for instance, once the top slice has been mined out, the gate roads of the lower slice must be located either beneath the goaf area or in the in-situ coal seam. When these roads are located beneath the goaf area of the upper slice and in shadow stress conditions, the roof of the gate roadways of the lower slices may already be fractured as a result of stress distribution during the process of mining the upper slice. Therefore, driving and protecting those gateroads would be difficult and costly. In contrast, if the gate roads are located in the in-situ coal seam, they may be directly beneath the abutment stress location. As a result, the protection of these gate roads would also be difficult and costly. For these reasons, the MSL method is only recommended for application to the thick coal seams in Quangninh, where the immediate roof is strong and needs assistance in fracturing prior to the extraction of the lower slices in the coal seams. 66

3.3.4 Hydraulic mining

Hydraulic mining is considered to have signiﬁcant potential in complicated geological conditions, where mechanized methods would encounter operational diﬃcul- ties or be economically unfeasible. This is because it has the following advantages (Hebblewhite et al., 2002; Hebblewhite, 2005):

• The level of mechanical complexity in hydraulic mining is significantly reduced compared to conventional mechanised mining, even though the mining layouts in both mining systems are similar. This advantage makes hydraulic mining operationally flexible in the extraction of coal from working mine areas where mechanised methods would otherwise encounter operational difficulties or would be economically unfeasible. In addition, hydraulic mining operates successfully in steeply dipping seams which have challenged traditional mechanised techniques because the broken coal from the face is mixed with water and transported by gravity flow. Therefore, the profitability of a hydraulic mining operation is less affected by these constraints. As a result, this method is suitable to the geological conditions in the Quangninh coalfield.

• Hydraulic mining is much safer than traditional underground coal mining methods due to the removal of personnel from the face area, the ability to automate equipment operations, the reduction of coal dust and the elimi- nation of frictional ignition sources from the production process.

• Hydraulic mining oﬀers signiﬁcantly lower capital and operating cost structures than conventional mining, particularly if the mining faces are located above the drainage level.

• Typically, the tonnages obtained from hydraulic mining operations are less than those obtained from mechanised methods. However, due to reduced personel requirements, hydraulic mining is still highly productive (high tonnes per man-year rating) 67

However, the application of hydraulic mining would be limited under the current operational situation in Quangninh. This is due to ﬁve challanges:

• The first challenge when applying hydraulic mining in the Quangninh coalfield is the high rate of development per tonne of coal production. In general, the layout in hydraulic mining is the same as in the room and pillar method. The difference is that coal extraction in hydraulic mining is carried by high pressure hydraulic jet. As discussed in Section 3.3.1, the high rate of development per tonne of coal production would cause difficulties in increasing production, results in increasing operation costs due to the low advance rate of roadway development in the Quangninh coalfield.

• The second challenge is the complexity of hydraulic transportation and storage under the current operational status of Quangninh. At present, coal processing is undertaken by the central coal preparation plant. This means that one coal preparation plant in Quangninh is responsible for the processing of coal extracted by several coal mines in the region, with the distance from the mine sites to the coal processing station ranging from 4 - 10km. Therefore, if hydraulic mining is applied in the Quangninh coal- ﬁeld, the transportation of the mixture of coal and water from mine sites to the preparation station would be complicated and costly. In addition, transportation of recycled water back to the mine sites is also complicated. In order to apply the hydraulic mining method to underground coal mines in Quangninh, a new dewatering station needs to be built near the mine sites to reduce the cost of coal transportation. This could increase the capital and operating costs in coal production as well as the complexity of operational management.

• The third challenge of the application of hydraulic mining in Quangninh is the inappropriate technology in coal preparation being used by many plant stations. Coal preparation in Quangninh is mainly undertaken by the suspension method with limitations in the separation of ﬁne coal (Industry Investment Consulting Company, 2005). Due to this method’s diﬃculties 68

in fine coal beneficiation, only a coal product with a size of 10mm (some plants have a size of 15mm) or greater will be beneficiated. Coal product with a size less than 10mm will only be classified into products according to their ash content, and will not be processed. As a result, there will be a big difference in the coal price between the different sizes of coal product. For instance, due to being beneficiated, the domestic price of anthracite coal product with a size of equal or greater than 10mm and ash content of 9% is 2,500,000 Vietnam dong (VND) per tonne, while the fine coal (not being beneficiated) with a size of less than 10mm and ash content of 36 - 42% has a market value of 510,000 VND per tonne (VINACOMIN, 2008). In addition to this, the fine coal, as a result of run-off mine product from suspension coal separation, is dried by discharging it into the reservoir (tailing dam), for gradual natural drying. The fine coal needs a long time to dry and it still retains a high percentage of moisture and a high ash content. Hence, its market value is low.

• Another challenge in applying hydraulic mining to the Quangninh coalfield is the treatment of spillage and contaminated runoff water created by this method. Currently in Quangninh, this treatment is undertaken with poor technology, mostly by natural methods. Water from the mine sites is normally discharged directly into the streams or river nearby, causing environmental pollution in the region. For this reason, management and treatment of the very large amount of spillage and contaminated run-off water created by hydraulic mining is another disadvantage to apply hydraulic mining in the Quangninh coalfield.

• In addition, at the present, coal reserves in the Quangninh coalﬁeld located above the drainage level are almost mined out, and many underground coal mines have started to develop and extract coal reserves at a lower elevation compared to the level of the infrastructure stations. Therefore, the cost of transportation of coal won by hydraulic mining could account for a large portion of operating costs. This is also another diﬃculty when 69

using hydraulic mining in the Quangninh coalﬁeld.

In summary, hydraulic mining has the potential to be highly productive under steep and complicated geological conditions, but has limited potential at the present time due to the current operational situation in the Quangninh coalﬁeld. Before consideration of its application in the Quangninh coalﬁeld, the following improvements in mining technology would need to be addressed:

• Improvements in roadway development, especially in applying rockbolts or cable supports in order to increase the advance rate of roadway development, to reduce time and operational costs.

• Research and experimental application of improved methods in coal dewatering and coal preparation, especially for ﬁne coal.

• Means of improving waste water management.

3.3.5 Longwall Top Coal Caving

The LTCC method is proven to have potential in thick seam mining as it overcomes the disadvantages of both the HSPL and the MSL methods. The main advantages of this method have been pointed out by many researchers (Jian et al., 1999; Hebblewhite and Cai, 2004; Xu, 2004; Hebblewhite, 2005; Vakili, Cai and Hebblewhite, 2007). Those are:

• Reduction of operating costs: The output of coal per metre of gate road development from the LTCC method is much more than the conventional longwall method due to coal extraction from caving, which reduces the development costs. Jian et al. (1999) stated that the output of the LTCC method can be 1 to 3 times more, while the cost is 30 - 50% less than the MSL method under the same mining conditions. In addition, the entry layout of the LTCC method is much simpler than that of the MSL method, so some development work can be saved. 70

• Compared to the HSPL method, the LTCC method oﬀers a lower face height, resulting in smaller and less expensive equipment, and better face conditions (Xu, 2004; Hebblewhite, 2005).

• Resource recovery and the mine ﬁnancial performance: The LTCC can extract up to 80 - 85% of seam in the 5 - 9m thickness range. This mining method is suitable for extracting coal seams with irregular thickness which face problems in extracting in the last slice in the MSL method.

However, the LTCC method is not suitable for all thick coal seams. There are some constraints that need to be considered in relation to the application of the LTCC method. They are:

• The LTCC method is not suitable for coal seams which have a strong, competent roof overlying strata. If the roof above the coal seam is too strong or competent and does not collapse behind the roof support, a large gap between the canopies of the roof supports and the immediate roof will appear after the top coal of the longwall face is drawn. If the immediate roof still hangs as the longwall face advances, it could suddenly collapse when it exceeds a critical limit. The sudden collapse of the roof over a large area would cause windblast and, potentially, serious injuries to miners and equipment. More details of the geological factors inﬂuencing the success of the LTCC will be discussed in chapter 5.

• The properties of top coal in the seam also play an important role in the successful application of the LTCC method. If the top coal is too hard and it does not cave or caves in large blocks that cannot be drawn through the shield’s windows, this method will be inefficient because of the low recovery rate from top coal. In contrast, if the top coal is too soft, preventing roof cavity during mining operations, (the roof in this case is the top coal) would be a challenge. This problem may slow the advance rate of the longwall face, affecting the efficiency of mining production. 71

• Another constraint in using the LTCC method is potential spontaneous combustion. Coal is more likely to be left in the goaf area in the LTCC method than in the HSPL or the MSL methods. Hence, the LTCC method could face the potential hazards of spontaneous combustion more than two other methods. Therefore, the characteristics of the coal and the gaseous contents in the coal seam will also inﬂuence the application of the LTCC method. Since the coal seams in Quangninh have a low gas concent, this problem will not aﬀect this method in application there.

It is clear that the advantages outweight the disadvantages of the LTCC method. When compared with other mining methods, the LTCC method can be considered the most suitable application for the inclined thick seams in the Quangninh coalfield under suitable geological conditions. The LTCC method has been applied in the Quangninh coalfield since the 1990s. The underground mines in Quangninh that had utilised the LTCC method all demonstrate the favourable prospects of that method. Also, the miners are experienced with the method’s operations. Currently, the Vietnam National Coal-Mineral Industries Group (VINACOMIN) has been studying the feasibility of utilising mechanized equipment to improve the production of the LTCC method. Hence, the LTCC method is identified as the method with most potential applicable to the Quangn- inh coalfield. The success of an LTCC operation, from all three perspectives of safety, resource recovery and productivity, depends largely on the geotechnical environment and the management of both operational and geotechnical aspects. There- fore, it is essential to conduct a detailed study of operational and geotechnical issues associated with the application of LTCC methods for inclined thick coal seams. There are numerous considerations which have been investigated as part of this study. In summary, these are:

• Operational issues associated with the application of the LTCC method for inclined coal seams, such as equipment stability, intersections between face ends and gate roadways, equipment relocation and transportation of the 72

equipment in mines.

• Geotechnical issues: the eﬀect of the dipping factor on the caving mechanism of the top coal and stress distributions around the longwall face. In addition, other factors, such as the properties of the coal and rockmass, and the eﬀect of horizontal stress on the caving mechanism have also been pursued in the study.

3.4 Conclusions

This chapter has attempted to evaluate the most appropriate thick seam mining methods applicable to Quangninh’s geo-mining conditions. By an evaluation of the advantages and disadvantages among different mining methods currently applied around the world, coupled with the geo-mining conditions and the level of the mining industry in the Quangninh coalfield, the study re-confirmed that the LTCC method is the method with most potential for extracting the inclined thick seams in the Quangninh coalfield. In the literature review, five underground mining methods which have been, or are being applied for extracting thick coal seams in coal mines around the world were investigated:

• Room and pillar and Blasting Gallery methods;

• High reach Single Pass Longwall;

• Multi-slice Longwall;

• Hydraulic mining; and

• Longwall Top Coal Caving.

The room and pillar method was considered to have a limited application in the Quangninh coalfield due to the low percentage of coal recovery rate and the high rate of roadway development per tonne of coal production. The unsuitability of application in high cover depth and inclined seams were also constraints of 73 this method in Quangninh conditions. The blasting gallery method, which is a modification of the room and pillar method, was also reviewed. However, the difficulties of driving a large number of roadways along the seam dip and the high percentage of coal loss prevent this method from being a viable option in Quangninh’s conditions. The High reach Single Pass Longwall method was also considered to have a limited possibility in the Quangninh coalfield. It was apparent that technology was already gradually increasing both the shearer and supports height up to 5m and above. However, limitations such as equipment stability due to the dipping factor, large size and heavy weight, high capital cost of the mining equipment, and particularly the difficulties in the relocation and transportation of equipment in mines rule this method out as a suitable option for the Quangninh coalfield at the present time. The multi-slice Longwall method was also reviewed in detail with consideration of experiences both in other countries and Vietnam. However, through an evaluation of its advantages and disadvantages, the study revealed that this method contains many drawbacks, such as problems mining under the goaf area, the stability and protection of roadways, and difficulties in maintaining the slice direction in coal seams with a high variation in thickness. For these reasons, this mining method is only suitable for extracting those coal seams with a strong or competent roof which needs the pre-fracturing of immediate roof prior to the extraction of the lower slices. Hydraulic mining was found to have significant potential in the complicated geological conditions where mechanized methods encounter operational difficulties or are economically unfeasible. However, the current technology in coal processing and the infrastructure existing in mines prevent this mining method being applicable under Quangninh’s mining conditions. The main issues in applying hydraulic mining in Quangninh are: (1) the low technology in coal processing resulting in difficulties in beneficating a large propportion of fine coal product created from hydraulic mining; (2) complications in transporting the mixture of coal and water from the mine site to the coal preparation station and carrying the recycled 74 water back to the mine site under the existing infrastructure in Quangninh, (3) the difficulty in treating the spillage and contaminated runoff water discharged from hydraulic mining, and (4) the high rate of roadway development per tonne of coal production. In the future, when the technology of coal processing, and roadway development is improved, it is believed that hydraulic mining will have further chances to prove its viability in the Quangninh geo-mining conditions. The LTCC method has been proved suitable for thick seam mining, as it overcomes the disadvantages of the multi-slice longwall and High reach Single longwall methods. Its recent application in China shows that this mining method has achieved impressive results, such as high recovery, high production and productivity. Through caving coal from the upper section during mining the lower section of a thick seam, the LTCC method significantly reduces the development cost per tonne. The LTCC method enables up to 80 - 85% seam extraction in the 5 - 9m thickness range. It also overcomes the disadvantages existing in High reach Single Pass Longwall and Multi-slice longwall methods. Compared to High reach Single Pass Longwalling, the LTCC method offers a lower face height, resulting in smaller and less expensive equipment, and better face conditions. The experiences from underground coal mines in Quangninh, that have utilised the LTCC method, have returned favourable impressions and views of the prospects for the method in Quangninh underground mines. The output of coal per metre of gate road development from the LTCC method is much more than the multi- slice longwalling, and the entry layout of the LTCC method is much simpler than that of the multi-slice longwall method. Hence, the Longwall Top Coal Caving is further confirmed as the method with the most potential for extracting the inclined thick seams in Quangninh. Chapter 4

Operational issues for the extraction of inclined thick seams by the LTCC method

4.1 Introduction

As discussed in Chapter 3, the LTCC method is considered the most suitable application for the extraction of inclined thick seams in the Quangninh coalfield. Nevertheless, the complication of applying this mining method in the Quangninh coalfield is that a large proportion of the thick seams belong to inclined seams. Therefore, understanding the operational issues during the mining processes in steeply dipping conditions is necessary for successfully using the LTCC method for extracting the inclined thick seams. This chapter aims to highlight the operational issues that may arise when applying the LTCC method for the extraction of inclined thick seams in the Quangninh coalfield, and to offer practical solutions to these problems through the analysis of current operational practice. The study begins by analysing the advantages and disadvantages of the face directions for the extraction of inclined thick seams. The equipment required for the LTCC method for extracting in steeply dipping conditions is then identified. Finally, the issues associated with

75 76 the operations of longwall faces in inclined coal seams and practical solutions for these problems are proposed.

4.2 Face directions in inclined seams

One important operational concern for the extraction of steeply dipping seams is the direction of the longwall face. Compared to ﬂat seams, the longwall face for the extraction of inclined thick seams can retreat in one of the three following directions:

• Down the seam dip

• Up the seam dip

• Along the strike.

The choice of the direction of the LTCC face will certainly affect the efficiency and economical profitability of this method. For the extraction of a inclined seam, the three directions of the longwall face have reportedly been used in different coal mines. For example, for extracting inclined thick coal seams by the multi-slice longwall method with backfill method in Poland, the face is mostly oriented in an up-dip direction to allow better drainage of the hydraulic backfill (Palarski, 1999). For extracting the lower section of inclined thick seams in Russia by the LTCC method, top coal caving under steel mats (the 2.0 - 2.5m top section is extracted by the conventional longwall method, laying steel mats on floor during operation), the face was oriented in a down-dip direction (Adam, 1976) probably to attract more coal from the middle layer between the two slices. However, in many inclined coal seams in Vietnam, the longwall face usually retreats along the strike. There is a need to understand the advantages and disadvantages of each direction of the face when extracting an inclined seam to make better operational decisions. When the face retreats along the strike, the maingate and tailgate are horizontal, only the longwall face is steeply dipping. In contrast, when the longwall face retreats up-dip or down-dip, the longwall face, the maingate and the tailgate are all steeply dipping, as illustrated in Figure 4.1. 77

Figure 4.1: Diﬀerent directions of the LTCC face for extraction of an inclined thick seam 78

If the face retreats up-dip of an inclined seam, due to the main entries being driven into the upper border of the mine site, a less time is needed to set up the first longwall panel compared to two other directions of the face. In addition, since the water can locate at the goaf area, better drainage at the longwall face is anticipated from this method. However, this face direction may have some disadvantages. The first is difficulties in driving the gate roadways (in the down-dip direction) of a longwall panel. The face of gate roadways during the development process is also at the lowest level so that it can collect the water. This results in complications in dewatering at the face during roadway development. In addition, because the transportation of the coal product during roadway development and longwall extraction is in the up-dip direction, its cost is more expensive than that when the face retreats down-dip or along the strike. Another difficulty for the face retreating up-dip of the seam is to maintain the stability of the supports and other equipment at the face during mining operations. Due to the effect of the dip, the supports, or other mining equipment, may face problems of slippage, especially when the goaf area is not filled completely by the broken rockmass. Correcting the supports when they slip into the goaf area would be a dangerous and complicated process. Figure 4.2 presents the trend of slippage of supports and the conveyor for the face retreating up-dip of an inclined seams.

Figure 4.2: The potential slippage of supports for the face retreating up-dip in an inclined seam

The face retreating down-dip in an inclined seam may ensure more stability of the supports and other equipment than the face retreating up-dip of the 79 seam, because the supports of the longwall face can lean directly on the face. In addition, the transportation of coal from the roadway development and the longwall extraction consume less energy than the two other methods. However, this method requires a long development time for the first longwall extraction compared to the two other face directions. As shown in Figure 4.1, the main entries are first developed at the lower border of the mine site, and then the gateroads of the longwall panel can be driven. As a result, the time to operate the first longwall face will be longer than the face retreating up-dip or along the strike. The ventilation of the single gate roadway during its development in the case of the up-dip direction is also a challenge because the return air has to flow in a down-dip direction. Moreover, controlling the intersection between the face ends and the gate roadways would be difficult because both gateroads and the longwall face are steeply dipping. In terms of the operational point of view, the face retreating along the strike has many advantages compared to the face retreating up-dip or down-dip of an inclined seam. The first advantage is the driving of the roadway development along strike, where the maingate and the tailgate are both horizontal (only the roadway for the longwall set up is steeply dipping) is easier than driving in inclined situations. In addition, the ventilation of the single gateroad during development, in this case, is much simpler than in the case where the face retreats down-dip or up-dip of the seam. As a result, the length of the longwall panel along the strike can be developed far more than that of the face retreating down-dip of the seam dip. Driving the horizontal gate roadways with very slight dipping (2-50/00) also can allow underground water to drain to collection points at the main entries so that it can de-water during face extraction. In addition, transportation and gathering of supports and other equipment in horizontal conditions is much easier than at dipping conditions. A study on the differences in the caving mechanism between three different face directions in Chapter 5 concluded that face retreating along the strike has better cavability compared to the face retreating up-dip or down-dip of the seam. For these reasons, the following sections only discuss the LTCC when the direction of the longwall face is along the strike. 80

4.3 LTCC face equipment

It is understood that a proper selection of the supports is the key element to a successful operation of the LTCC method. Different geological conditions require different kinds of face supports. In the Quangninh coalfield, underground mining companies have been struggling to find the most suitable equipment for the extraction of inclined thick seams by longwall methods, including the LTCC method. As noted in Chapter 2, many attempts have been made over the last decade to improve coal production, recovery rate and safety management for the extraction of thick seams in the Quangninh coalfield by the LTCC method. Since 2000, for instance, semi-mechanized shields (Figure 4.3) have gradually replaced the wooden props and cribs, and hydraulic single props (Figure 4.4) to strengthen the support capacity of the LTCC faces (Du, 2003; Dac and Tuan, 2006).

Figure 4.3: The longwall face supported by semi-mechanized shields (source: Tuan and Thang, 2003)

Operation of this supports require a large manual work. As illustrated in Figure 4.5, the operating sequence of semi-mechanized supports starts with face extraction by the drilling and blasting method with a web width of 0.6 - 0.8m. After some of the broken coal is loaded, a new web of wire-mesh is laid on the roof area and the ﬂipper is pushed into a fully extended position to support the wire-mesh layer and prevent the roof falling. After broken coal along the bottom of the face is completely loaded, the conveyor advances to the new position. The supports then advance into the new web and set against the roof. The movement of the supports is undertaken by two side beams of the canopy 81 with the help of double-acting rams. To advance a shield, the left beam of the shield is ﬁrst lowered, then it is advanced by the double-acting ram of the right beam to a step forward position. This side beam is then set against the roof and holds the left side beam in the forward position. The right beam of the shield is then lowered and is pulled up to the new position by its own double-acting ram, and set.

Figure 4.4: Longwall face supported by hydraulic single props, (source: Tuan and Thang, 2003)

The major advantage of this type of support is that it is easy to set up the face, and relocation of the face due to the light weight of the mining equipment (the maximum weight of a semi-mechanized shield is only 980kg). Another advantage is that this type of support equipment is less expensive than mechanized supports. Despite the obvious advantages of these supports, several problems exist. The most significant of these is the poor load distribution demonstrated by the supports. This leads to excessive mechanical breakdowns of the supports and consequent damage of the immediate roof, as has been witnessed in many longwall faces in the Quangninh coalfield. Another problem is that because there is a large gap between the beams of the shield and between the shields, a potential roof fall may occur. As witnessed by the author during field visits, although all LTCC faces applying these types of support using wire-mesh on the roof to 82 reduce the potential of roof fall, roof fall cavities still commonly occurred as wire- mesh was torn. In addition, also due to the roof fall, the beams of the supports cannot make even contact with the roof, therefore support tilting is common, and influence the efficiency of the mining operation. Figure 4.6 presents the problems caused by using semi-mechanized shields at the longwall face with steeply dipping conditions.

Figure 4.5: Operational sequence of semi-mechanized shields used at LTCC faces in the Quangninh coalﬁeld

Due to the many problems occured during mining operations and the large 83

amount of manual work required, production and productivity are still limited. The production of one LTCC face supported by this kind of support (model XDY- 1T2/LY), and with face extraction by the drilling and blasting method is limited to 150,000 tonnes per year (Tuan, 2005).

a. Supports tilting due to roof sagging b. Roof fall due to wire-mesh tearing

Figure 4.6: Operational problems when using the semi-mechanized shields in the Quangninh coalﬁeld

Since 2007, some underground coal mine companies have experimentally applied another semi-mechanized shield model to support the LTCC face. In this type of shield the canopy is wide and has no space for the broken rock falling from the roof to the face. Additionally, to prevent the potential of the support tilting at a steeply dipping face, a system of beams located under the canopy of each shield is linked together during the mining operations, as shown in Figure 4.7. Since there is no space between the shields, laying of the wire-mesh is not needed during mining operations. As a result, much manual work has been saved. In addition, the problem of the support tilting is minimised. Results from initial applications show that the production of an LTCC face supported by this type of support range between 200,000 - 250,000 tonnes per year (Du, 2009b). However, some disadvantages still exist in this type of support. The ﬁrst is that this type of support is only suitable for extraction by the drilling and blasting method, because there is no base canopy for aiding double-acting rams. Hence, pushing the conveyor forward is very diﬃcult. In addition, the advancement of 84 the supports needs a large amount of time as each shield requires the digging of holes for the toe of the props. As a result, production is still limited. Moreover, there is no cover on the rear of the support, therefore, preventing the broken rock in the goaf area falling into the working face is also a challenge.

Figure 4.7: Model ZH1600/16/24Z semi-mechanized support (source: Du, 2009b)

In another attempt to improve the technology in underground thick seam mining, a longwall face supported by mechanized supports model VINAALTA - 2.0/3.15 was also trialed at No.8 seam, Tay Vangdanh, Vangdanh Coal Company in November 2007. This support model was similar to the European designed shields which have a chute window located at the middle height of the shield canopy, as illustrated in Figure 4.8. The output of this longwall face has increased dramatically to approximately 450,000 tonnes, but the face is limited at gently inclined coal seams (an average seam dip of 140). It is clear that in order to increase coal production, productivity, and standards of safety management in the extraction of thick seams with steeply dipping conditions, the level of mechanisation should be improved, so that all processes of mining operations, such as face extraction, coal loading, and support move- 85 ment should operate automatically as much as possible. In addition to this, other systems in mining operations, such as coal transportation, and roadway development also need to improve so that the coordination between different production stages operates efficiently and effectively.

Figure 4.8: Mechanized support model VINAALTA- 2.0/3.15 trialed in Vangdanh coal mine (source: Du, 2009b)

For extracting an inclined thick seam by the LTCC method, mechanized equipment should be used in a manner that maximizes production, productivity, and minimizes technical diﬃculties. In general, the equipment components for a mechanized LTCC face can fall into the four following categories (Hebblewhite et al., 2002):

• Hydraulic Supports

• Shearer

• Armoured Face Conveyor (AFC), and

• Auxiliary equipment.

In a mechanized LTCC face, the majority of the face equipment, such as the shearer, chain conveyor(s) and pantechnicon can be described as conventional. The main difference is the face supports which have specialised caving functions 86 for drawing coal from the top coal section during the extraction of the cut section. For this reason, the following section will only discuss those issues regarding the longwall supports and the different mechanisms for drawing the caved top coal. Detailed descriptions of other components of equipment for a longwall face can be found elsewhere (Stefanko (1983), and Peng and Chiang (1984)). Hebblewhite et al. (2002) and Xu (2004) among others, undertook comprehensive literature reviews of the LTCC supports which have been used in international thick seam mining practice. These reviews identified a number of different configurations of support that have been used throughout the world. These configurations can be separated into two categories based on the position of the caving window: “European” and “Chinese” designed shields (Hebblewhite et al., 2002). European designed shields are comprised of a chute window located in the middle of the shield canopy, as illustrated in Figure 4.9. The operating sequence for this type of support starts with the extraction of the lower section of the coal seams, as does the conventional longwall method. After the shearer completely extracts and loads the coal in the new web, the conveyor will be advanced. The support then advances a step forward into the new position and is set against the roof (the process of setting the support against the roof aids in fracturing the top coal). The chute is then extended and the caving door in the support canopy is opened by the operation of hydraulic double-acting rams. Caved coal from the roof is allowed to flow onto the face conveyor. The operation of the supports at steeply dipping conditions can be seen in Figure 4.10. European designed shields have both advantages and disadvantages. This kind of support has demonstrated excellent stability and ground control, and provided a safe working environment for operators. Experimental application of the high- opening support, designed in China (which is very similar to European caving support designs) noted that very few accidents were reported during operations (Xiaoyan, Deyun and Baoli, 1999). However, some disadvantages also exist in this design. Because coal is allowed to flow onto the face conveyor, simultaneous cutting and caving operations are impossible with this design. This affects the 87

Figure 4.9: An example of the European designed shield (after Hebblewhite et al., 2002)

Figure 4.10: Operating Range of the European designed support with steeply dipping conditions (after Hebblewhite et al., 2002) 88 efficiency of mining production. In addition to this, due to the caving chute being located at a high position in the face, caving operations produce large concentrations of dust on the face, and the support design provided only marginal levels of top coal recovery (Hebblewhite et al., 2002). Chinese designed shields have been involved in a great deal of iterative design and experimental work that has been developed since the 1980s. During different stages of development, a large number of shield designs have been developed through parallel testing so that the most appropriate support configuration could be established. The most notable improvement is the change of the position where top coal is caved from the middle to the rear of a shield, in addition to an arrangement of another rear conveyor, as shown in Figure 4.11. The rear caving canopy that was attached to the main roof canopy for cover, can be articulated and the canopy insert retracted, so as to direct the caved top coal onto the conveyor. This LTCC support design has the following advantages (Hebblewhite et al., 2002):

Figure 4.11: An example of Chinese style shield (after Hebblewhite et al., 2002)

• Since the rear caving canopy is long, the area for eﬀective top coal drawing is large, therefore this design is very productive, and the level of top coal recovery increased with this design.

• Due to the caving point being much lower than the middle height caving window of the European style shields, the coal dust created by the caving process is signiﬁcantly reduced. 89

• Due to the large space created underneath the rear caving canopy, maintenance in this area is much easier than it was for mid-opening designs.

After decades of testing and development, the new low-opening shield design (seen in Figure 4.11) has become the basis for all of the newer LTCC support designs used in China since 1992 (Hebblewhite et al., 2002). In addition, this design is structurally simple and exhibited excellent load distribution, strength and stability. The LTCC face supported by this shield design has proved to have a very high production capability. For example, one LTCC face at Dongtan Mine (Yankuang Mining Group Corporation Ltd.) supported by this shield design in 1998 produced 5.1 million tonnes (Hebblewhite et al., 2002). In order to efficiently apply the LTCC supports into complicated geological conditions, light duty LTCC support models, which are smaller, easier to relocate and contain multi-functional equipment were developed in China. These support models were the result of a modification made to a LTCC support design used in steeply dipping seams. Figures 4.12, 4.13, and 4.14 show examples of models of light duty support used in China. Table 4.1 presents specifications of the major types of light duty LTCC supports used for small or medium coal mines.

Figure 4.12: Model ZFSB 2200/16/24A light duty LTCC Support (after Xu, 2004) 90

Figure 4.13: Model ZFSB 2200/16/24B light duty LTCC Support (after Xu, 2004)

Figure 4.14: Model ZFQ 2000/16/24 light duty LTCC Support (after Xu, 2004)

Table 4.1: Speciﬁcations of 11 types of light duty LTCC supports (after Xu, 2004)

Support Working Support Yield Base Weight Type height (m) width (mm) Load (KN) pressure (MPa) (tonne) ZFB2200/16/24 1.6-2.4 1156-1296 2011-2275 0.6-1.2 5.8 ZF2200/16/24B 1.6-2.4 1450 2104-2285 0.5-1.1 8.2 ZFSB2200/16/24A 1.6-2.4 1450 2200 0.6-1.2 6.9 ZFS2400/16/24B(A) 1.6-2.4 1450-1622 2247-2459 0.3-0.9 8.7 ZFS1800/16/24B 1.6-2.4 1000 1703-1967 1.0 3.7 ZFBZ2200/16/24 1.6-2.4 1430-1570 2024-2228 0.3-1.47 6.2 ZFB2400/16/24D 1.6-2.4 1400-1570 2132-2417 0.39- 0.83 7.4 ZFB2000/16/24 1.6-2.4 1450-1590 1920-2056 0.6-1.2 6.9 ZFZ2000/16/24 1.6-2.4 1220-1360 1920-2056 0.6-1.2 6.3 ZFB2800/16/24 1.6-2.4 1400-1570 2487-2819 0.42-1.3 7.8 ZFZ2000/15/23 1.5-2.3 1220-1360 1905-2053 0.6-1.2 6.5 91

The light duty LTCC supports have all the same advantages of the low- opening support models as well as being easy to relocate and compact. In addition, the performance capability of these supports is attractive compared to the production and mechanization level in Vietnam. For example, the production of a LTCC face supported by the ZFS 2200/16/24B in Huatiang mine, Gan- shu province, achieved 1 million tonnes per year (Xu, 2004). As a result, light duty LTCC supports are considered to be the most suitable for application in the Quangninh coalﬁeld, where many coal seams have a short panel length and steeply dipping conditions. The reasons why there is a need to have light support for transport in a mine with steeply dipping conditions is discussed in detail in Section 4.4.

4.4 LTCC supports for steeply dipping seams

To adapt LTCC supports to inclined thick seams in Quangninh, some modi- ﬁcations should be made so that the overall eﬃciency of the LTCC operation is enhanced. To successfully operate LTCC supports and other equipment for steeply dipping seams, the following issues should be considered:

• Support tilting

• Slippage of supports and other equipment

• Transport in mine

• Installation and mantling.

4.4.1 Support tilting

One of the potential hazards associated with the operation of an inclined longwall face is the tilting of the supports. Support tilting occurs when stability is lost. The major factors inﬂuencing support tilting consist of seam inclination, the extraction height, roof and ﬂoor conditions, the operator’s skills, and the support maintenance (Peng and Chiang, 1984). 92

Supports tilt wherever the projection of the centre of gravity falls beyond the edge of the base plate. When static, there are two forces that cause instability of the supports: the weight of the supports and the weight of the broken rockmass on the roof of the supports. The tilting moment of the support (M0) around the corner 0 (as shown in Figure 4.15) is presented in Equation 4.1 (Peng and Chiang, 1984):

Figure 4.15: Analysis of support tilting (modiﬁed after Peng and Chiang, 1984)

B M = (W H + P.H )sinα − (W + P )cosα (4.1) 0 r r s 2 r where:

Wr = the total weight of broken rocks applying on the support

Hr = the height of the centre of gravity of the rock pile on the support P = the weight of the support

Hs = the height of the centre of the support B = the width of the support α = the dip of the longwall face

It can be seen from equation 4.1 that support tilting will occur when M0 > 0. This means that at a certain value of seam dip, since the weight and the height of a shield are unchanged, the stability of the supports largely depend on the conditions of the loading force from the roof. If the roof is strong, the increase of 93

the force in the legs of the supports will increase the stability of the support due to an increase of the normal force of the supports. In reality, the most common cause of support tilting in an inclined face is due to the roof fall cavity above the supports, as illustrated in Figure 4.16. In this case the support is unstable because of the rotational moment generated by unbalanced contact between the canopy and the uneven roof. If the roof rockmass has contact, the support is stable due to the resistant force of the support acting upon the roof. In the case where there is no contact between the canopy of the support and the roof rockmass, the support becomes a free block and its stability depends on body weight and the centre of gravity of that support.

a. A strong roof helps keep supports stable b. Support tilting caused by roof fall

Figure 4.16: Support tilting in an inclined face

Support tilting might also occur at a steeply dipping face during the movement of the supports. When a support starts to advance, it has to be lowered, and the canopy of the support does not maintain contact with the roof. When the dipping face is too large and the centre of gravity is too high, tilting can occur. As mentioned previously, this potential problem has been commonly occured in the Quangninh coalfield, where the LTCC face is supported by the semi-mechanized shields (shown in Figure 4.6). The common method of preventing the potential tilting of the supports in an inclined longwall face is to reduce the height of the longwall face, hence keeping 94 the support’s centre gravity as low as possible. The height of an inclination longwall face plays an important role in maintaining the stability of the support. As can be seen from Figure 4.15, in the case when the properties of the rockmass piling up on the support are unchanged, the moment of a support around the corner 0 depends largely on the height of the centre of support’s gravity. As a result, the greater the height of the longwall face, the less the lateral stability of the support. Under the same seam dip, the higher the longwall face, the less the stable the support. In addition, by keeping the height of the longwall face low, the weight of the support will be reduced. This can make transporting the supports in mines easier. With the current average height of the workforce, the height of a longwall face, ranging from 2.0 to 2.5m is reasonable because it is high enough for miners to work comfortably. 2.0 to 2.5m is also a reasonable height for the miners to reach to solve potential problems such as roof fall or roof cavity. Another method to reduce the potential tilting of the supports in an inclined longwall face is to link the supports together. When applying mechanized LTCC supports, and cutting the face by shearer, each shield is connected to the armoured face conveyor by its double acting ram. Due to the connection with the conveyor, the potential problems of support tilting will be reduced. The supports can also be linked together through a system of one or two beams located near the roof, as trialed in the Quangninh coalfield shown in Figure 4.7, or the beam can be located on the floor, which has been practised in the Western coal mine, USA, as presented in Figure 4.18.

4.4.2 Slippage of equipment at an inclined longwall face

Another problem faced by the operation of the LTCC face in steeply dipping conditions is the potential slippage of the supports and conveyor. They may slice along the seam dip when advancing due to their own weight or the impetus of the shearer during face extraction. Peng and Chiang (1984) state that due to the effects of many factors, including the weight of the conveyor, the cutting resistance of the shearer, and the influence of loaded coal, the conveyor will move 95 downward very slightly in each advance. This force causes the advancing ram and the frontal guide of the support to tilt toward the down-dip side. Consequently, whenever the advancing ram is activated to push the tilt conveyor, the conveyor moves farther toward the down-dip side, and thereby increases the amount of slippage. Eventually, the support will be forced to alter its direction of advance and moves down on the dip side, as illustrated in Figure 4.17. The slippage in each shear may be small, but the production will be seriously affected if such operation continues without any adjustment. The most complicated situation occurs when the supports cut into the maingate. Consequently, adjusting face alignment is both labour intensive and time consuming.

Figure 4.17: Factors leading to support slippage (Peng and Chiang, 1984)

In literature it was found that many attempts have been made to prevent the slippage of the supports and conveyor in steeply dipping seams. One of these is to group three shields together, which has been successfully applied in a Western US coal mine where the seam dips range from 30 to 350 (Peng and Chiang, 1984; Curth and Listak, 1986). In this method, three shields are grouped into one set, called a “troika”, using a 4.5m long floor beam, and these sets are advancing independently from the face conveyor, as shown in Figure 4.18. The centre shield of each troika set is rigidly attached to a floor beam. To advance a troika set, the centre shield is first lowered from the roof, then the double-acting rams of the two outer shields push the floor beam (and the centre shield as well 96 because the floor beam and the centre shield are rigidly connected together) which advances a step forward by the double-acting rams of the two outer shields. The centre shield is then set against the roof and holds the floor beam in the forward position. The lower and the up-dip shield are then lowered, pulled up to the floor beam by their own double-acting rams, and set against the roof in a sequenced order.

Figure 4.18: A method to improve the stability of the supports in a dipping face (after Peng and Chiang, 1984)

The face conveyor is pushed by a double-acting ram which is connected to the up-dip shield of each troika set (but not attached to the conveyor) and the conveyor cannot be retracted. An in-face anchorage system of rams connects every troika set to the conveyor to prevent the slippage of the conveyor caused by its own weight and the impetus of the shearer. The anchorage must be released for each troika advance. This type of support has two advantages. Firstly, due to the increase of the width of the support (three different shields have linked into one rigid shield) it reduces the probability of support tilting. Secondly, it can prevent the possible slippage of the supports during advancement. However, some constraints may exist. The major difficulty is to operate in adverse floor conditions and the accumulation of debris. In addition, keeping the face straight and controlling the roof when traversing strata discontinuities, such as faults, proves impractical with 97 the Troika approach (Curth and Listak, 1986). Another measure is to use side rams to adjust the alignment of the shields. This has been used in Germany for the extraction of steeply dipping seams (Curth and Listak, 1986). In this method, each shield is connected to the conveyor by its double acting rams, as in flat seams. Each shield has three side rams which are used to adjust alignment during mining operations, as presented in Figure 4.19. The conveyor is hitched by a system of in-face anchorage in order to prevent the potential slippage of conveyor. Figure 4.20 illustrates the in face anchorage system.

Figure 4.19: Aligning shield in steeply dipping face (after Curth and Listak, 1986)

Figure 4.20: Tensional in face anchorage (after Curth and Listak, 1986) 98

4.4.3 Support transport

The in-mine transport is another operational issue that should be considered for the application of the LTCC method in steeply dipping conditions. Most currently available vehicles are designed for transporting equipment in horizontal or gently inclined conditions. The maximum dip the rubber-tired equipment can negotiate is 220 (Curth and Listak, 1986). For higher dipping seams, other alternatives should be used. A monorail system, or a system of rail tracks with pulling winches may be considered convenient methods for transport in inclined faces. However, transport in steep dipping conditions is more complicated than in horizontal conditions. Using a monorail system for transportation is the case, where the equipment of the longwall face is transported along a monorail that is hung on supports of the gate roadway, as shown in Figure 4.21. The greatest challenge for this transport method is to hang the monorail system from supports of gate roadways strong enough to carry the heavy equipment of the longwall face. As the support capacity of the shields increases, their size and therefore their weight also increases. Extremely heavy equipment may need special requirements in the setting of the monorail system. Additionally, heavy and large sized equipment may cause diﬃculties during the process of installation and de-installation in a face with steep dipping conditions.

Figure 4.21: Transport support by monorail system (after Peng and Chiang, 1984) 99

Another consideration for transporting equipment in steep dipping conditions is the possibility of using different methods of transportation from the surface to the set up room or when moving from one panel to another. In a flat coal seam, where all gate roadways are horizontal, this would not be a problem because the equipment can be moved by one vehicle from the surface to the set-up face. However, some constraints may exist when transporting equipment in an inclined seam where the route from the surface to the set up room consists of both horizontal and steep dipping conditions. Since transporting the equipment by rubber tyred vehicles or a locomotive with wagons (where the conditions are suitable) is fast and efficient, these types of transport are usually preferred. When the route includes both horizontal and dipping conditions, a combination of different transport methods is commonly chosen. However, transferring the equipment between the two above mentioned transport methods would be time consuming. In this case, smaller and lighter equipment could certainly be transported more conveniently than larger and more heavy equipment. Therefore, the selection of supports should also be taken into consideration in order to make sure they are well engineered to manage this equipment at steeply dipping conditions.

4.5 Transition between the face ends and the gateroads

An additional problem that may occur when using the LTCC method for the extraction of inclined thick seams, is the diﬃculty in controlling at the intersection between the face ends and the gate roadways. A common operational requirement of face ends in LTCC mining is to have enough room for driving heads and the transfer points. Since the roof supports at the inclined longwall face always tend to have instability problems, such as tilting or slippage, it is essential that the supports at the maingate not only have caving capability, in order to minimise the coal losses at the face ends, but are also strong enough so that the supports in the longwall face can lean on them and thus avoid the possibility of instability. 100

However, the challenge is that if the face supports lean on the face end supports too heavily, the advancement of the face end supports might be diﬃcult. One possible alternative is to let the face supports lean directly on the supports of the gate roadway, as illustrated in Figure 4.22.

Figure 4.22: Transition between the face end and maingate

Another difficulty in operating the longwall in steeply dipping conditions is the transition between the face end and the tailgate. In normal operations, when the LTCC face extracts the lower section of a thick seam, one face end of the face connects to the floor of the tailgate, as illustrated in Figure 4.23. This situation, at the face ends, can cause difficulty in controlling the roof at the tailgate as the coal portion at the intersection between the longwall face and the gate roadways can easily be broken. Gradually leaving coal portions at the lower section of the seam so that the roof of the longwall face can reach the roof of the tailgate, as illustrated in Fig- ure 4.24, is an alternative for improving roof control at the intersection between the face end and the tailgate. However, high pressure caused by the heavy weight of the supports and the loading from the roof could be a significant problem in preventing breakage of the lower coal portion. Consequently, this issue needs to be further investigated during experimental work. 101

Figure 4.23: Transition between the face end and the tailgate

Figure 4.24: Another method of transition between the face end and the tailgate 102

4.6 Conclusions

This chapter highlighted the operational issues that may occur during mining operations of inclined thick seams using the LTCC method in Quangninh’s conditions. By analysing current operational practices of extracting coal seams at steeply dipping conditions both in Vietnam and overseas, some conclusions have been drawn: For the extraction of an inclined thick seam, the longwall face can retreat in either one of three different directions: up-dip, down-dip or along the strike. Of those directions, a face retreating along the strike has more operational advantages than a face retreating up-dip or down-dip. The first advantage is that the development of the heading gates (the maingate and the tailgate) of a longwall panel for a face retreating along the strike, where the gateroads are in horizontal conditions, is more convenient than that of a face retreating up-dip or down-dip of the seam where the gateroads are at steeply dipping conditions. Because of the ease of the operating processes (face extraction, coal transportation, and material transportation) and ventilation (especially in driving a single entry), the longwall panel for a face retreating along the strike is likely to be longer and take less time in development than the panel developed along the seam dip. As a result, less frequency in face movement is anticipated for the face retreating along the strike. Another advantage is the convenience of transporting and gathering supports and other mining equipment in the horizontal gateroads. Therefore, from an operational point of view, it is recommended that for the extraction of inclined thick seams, the face should retreat along the strike. After decades of the development of underground thick seam mining, it is clear that in order to improve coal production, productivity, and the standards of safety management, the mechanized equipment needs to be used in the LTCC face to minimise manual work. The literature review revealed that a number of support configurations have been developed for the LTCC face, which can be separated into two categories based on the position of the caving window. Among two shield designs (European and Chinese designed shields), the Chinese designed 103 shields are considered to be more appropriate for LTCC faces. European designed shields proved to have good stability and provide a safe working environment for operators. However, they produce a large concentration of dust on the face during the caving process because the caving chute is located in the high middle of the canopy. The Chinese shield design, however, has a greater recovery rate because the rear caving canopy is long and the area for top coal drawing is large. In addition, due to the caving point being located in the lower height of the face, the creation of dust by the caving process is much reduced when compared to European designed shields. When the face retreats along the strike, some potential problems may exist when using the LTCC method to extract thick coal seams with steeply dipping conditions compared to horizontal seams. The biggest problem faced is the instability of the supports and other equipment (the possibility of tilting and slippage) during operations due to the dipping factor. Another problem is the complications in transporting the supports and other equipment in different dipping conditions. As transporting the equipment by rubber tyred vehicles or a locomotive with wagons (where the conditions are suitable) is fast and efficient, these types of transport are usually preferred when possible. In the case the transport route includes both horizontal and dipping conditions, a combination of different transport methods would be needed. This would be time consuming and costly because of the processes of transferring the equipment between various types of transport. As a result, the supports in steeply dipping conditions should have sufficient lateral stiffness to maintain their stability against any loading parallel to the face line (to prevent support tilting and/or slippage). In addition, they must provide adequate bearing pressures at the roof and floor interfaces to avoid excessive fracturing, compacting or movement of the strata that would compromise ground control or support stability. The support should have the ability to provide adequate roof and goaf cover to avoid the hazardous infiltration of roof material into the working area. The support also should have a reasonable weight so that transporting in mine or face move is economically reasonable. The transition between the face ends and the gate roadways is considered a 104 weak point that may create, not only labour and safety issues, but also reduce advance rate. As a result, the transition between the face ends and the gate roadways, when using the LTCC method for extracting inclined thick seams, needs a considerable amount of further investigation. Chapter 5

Eﬀect of Seam Dip on the Cavability of Top Coal in the LTCC method

5.1 Introduction

There are many factors influencing the applicability of the LTCC method, such as ventilation, spontaneous combustion, development systems, cultural acceptance and geotechnical issues (Hebblewhite et al., 2002). In terms of geotechnical issues, the cavability of the top coal and the above rockmass is one of the critical factors affecting the LTCC method’s successful application. The accurate estimation of the cavability of the top coal and the above rockmass can help mine planners to predict the applicability of the LTCC method at the feasibility stage as well as to develop a detailed plan of operational measures in order to prevent potential hazards during mining operations. Therefore, an investigation of the factors influencing the top coal when extracting thick coal seams using the LTCC method is important for successful assessment and prediction. Previous work studying the cavability of top coal in the LTCC method applies mostly to flat coal seams. Both empirical (Adam, 1976; Zhang, Wu and Wang, 1995; Jia, 2001; Mark and Molinda, 2005) and numerical approaches (Xie, Chen

105 106 and Wang, 1999; Yasitli and Unver, 2005; Unver and Yasitli, 2006; Humphries, Poulsen and Ren, 2006; Vakili, Cai and Hebblewhite, 2008) indicate that the cavability of the top coal is influenced by many factors, including coal strength, discontinuities in seams, top coal thickness, cover depth, overlying strata, in-situ stress, and moisture sensitivity. Results from these studies have considerably contributed towards the understanding of the factors affecting the cavability of top coal in the LTCC method. Compared to horizontal seams, inclined coal seams have received less attention from researchers, probably because they are less common. Therefore, many geotechnical factors influencing the performance of the LTCC method in inclined thick coal seams have not been investigated thoroughly. This chapter is aimed at improving the understanding of the effect of seam dip, coupled with other parameters, on the caving mechanism of top coal in the LTCC method. To achieve these objectives, a literature review is first undertaken in order to gain a better understanding of the factors affecting the caving mechanism. Then an investigation of the seam dip on the cavability of top coal is analyzed. This chapter contains three main elements:

• A literature review on the factors inﬂuencing the cavability of top coal is discussed to highlight the issues that need to be addressed.

• An investigation of the eﬀect of face orientation on the cavability of the top coal for the extraction of inclined thick seams by the LTCC method.

• A demonstration of the diﬀerences in the cavability of top coal between ﬂat and inclined seams for the LTTC method when the faces retreat along the strike.

5.2 Literature review

As discussed in Chapter 3, the Longwall Top Coal Caving (LTCC) method is a modified version of the conventional longwall operation. The main difference between the two methods is that for the conventional longwall method, the coal 107 product is only from the face, while in LTCC method, a thick coal seam is divided into two sections: the cut and the top coal. The longwall face is installed and operated in the cut section, which is located at the lower portion of the seam. The top coal section lies immediately above the power supports. As a result, the space in the goaf area, that needs to be filled by the caved fragments for the LTCC longwall face, is larger than that of conventional longwalling. Therefore, the movement of the overburden strata above the coal seam in the LTCC method can be considered the same as in the conventional longwall method. Figure 5.1 outlines the differences in the roof strata between conventional and the LTCC methods.

Figure 5.1: Roof strata in conventional longwall and LTCC methods (adapted from Peng and Chiang, 1984; Vakili, 2009)

According to Peng and Chiang (1984), there are 3 zones of disturbance in the overburden strata in response to longwall mining. The caved zone has a thickness of 2 to 8 times the height of extraction. In this zone the strata are broken and then, under gravity fall on to the mine ﬂoor in irregular shapes and sizes. As a result, the broken rock volume in its broken state is larger than that of the original intact strata. Above the caved zone is the fracture zone where the strata are broken into blocks by vertical cracks due to strata movement, and horizontal cracks due to bed separation. The individual blocks in the strata remain fully or partly in contact across the vertical fractures, through the horizontal force can be transferred. Above the fracture zone is the continuous deformation zone. The strata in this zone deform without causing any major cracks, therefore, they do not loose their integrity. Figure 5.2 presents the three diﬀerent zones in the 108 overburden strata due to longwall mining.

Figure 5.2: Three zones in the overburden strata due to longwall mining [(after Brady and Brown, 2004), adapted from Peng and Chiang (1984)]

In general, any rockmass will cave if a large enough area is undercut (Mawdesley, 2002). However, the challenge is to reliably predict the dimensions of initial and continuous caving in the initiation, propagation and continuous stages, so that maximum coal recovery will be achieved. Therefore, an evaluation of the factors influencing the cavability of the top coal is important for the feasibility of the application of the LTCC for thick coal seams in certain geo-mining conditions. The cavability of top coal is a critical consideration not only in terms of overall mineability, but also for the rate of recovery. Coal caves behind the supports should be of suitable sizes so that they can be drawn through the windows of the supports. Overly large blocks of the caved coal may cause difficulties in drawing coal from the longwall face, resulting in either high coal loss or a high cost in the second fragmentation. The cavability of the rockmass lying above the seam also needs to be taken into consideration during the estimation of the cavability of top coal in the LTCC method. It is believed that the natural properties of the immediate roof and the main roof affect the cavability of top coal. The caving process must continue to propagate through the layers of rockmass to fill the air gap in the goaf area as coal is extracted. Therefore, understanding the caving mechanism of the overlying roof strata is also crucial for predicting the cavability of top coal in the LTCC 109 method. The factors that influences the cavability of top coal in the LTCC method have been broadly investigated by several researchers (Xu, 2004; Mark and Molinda, 2005; Whittles, Reddish and Lowndes, 2006; Vakili, 2009). Those factors which are considered to have major effects on the cavability of top coal are: coal strength, the cover depth of the seam, discontinuities in the coal seam, the thickness of top coal, characteristics of the overlying roof strata, and the presence of ground water in the mining area. The following sections will briefly discuss the effect of those factors on the cavability of top coal in the LTCC method.

5.2.1 Coal strength

Strength is one of the main criteria in estimating the mechanical properties of coal or rockmass. The strength of coal or rock material refers to the maximum stress level which can be carried by a specimen (Hoek, Kaiser and Bawden, 1995). To understand the behaviour of jointed rockmass, the characteristics of the intact rock material and individual discontinuities, which make up the rockmass system, are investigated. Much information on the methods of estimating the strength of coal and rockmass has been published during the last few decades. Figure 5.3 summarises the ranges of rockmass characteristics, testing methods and theoretical considerations. The most widely performed test to study the strength of intact coal and rock is the uniaxial compression strength (UCS) of cylindrical specimens prepared from drill cores. This test is used not only to determine the UCS, but also elastic constants, Young’s modulus, and Poisson’s ratio of the rock material (Hoek et al., 1995). Another method to indirectly determine the strength of coal and/or rock, in areas where the preparation of specimens and the standard uniaxial compression test are not available, is the use of the Point Load Index. Techniques for determining the strength and deformability of rock materials are published in many textbooks and it would be inappropriate to attempt to review all this information here. Reference is made to the International Society for Rock Mechanics 110

Commission on Standardization of Laboratory and Field Test (1981), and Brady and Brown (2004).

Figure 5.3: Summary of rock mass characteristics, testing methods and theoretical considerations (after Hoek et al., 1995)

Strength is one of the main parameters in standard classification systems for predicting the strength of an in-situ rockmass during mining operations. For instance, Whittles et al. (2006) compared fifty different parameters described in 257 published articles relating to the use and development of rockmass classification systems, and identified that coal strength (presented by Uniaxial compressive strength) is the most influencial factor on the process of strata deformation. In 111 the Coal Mine Roof Rating (CMRR) system, developed by National Institute of Safety and Health (NIOSH) of the USA (Mark and Molinda, 1994; Mark and Molinda, 2005), which is used to assess the stability of the roof during underground mining operations, the UCS accounts for approximately one-third of the CMRR rating. Chinese statistical analysis methods (Zhang et al., 1995; Jia, 2001) also indicate that coal strength is one of the main factors affecting the cavability of top coal.

5.2.2 Discontinuities

The common types of failure in mining excavations in high jointed rockmasses are those involving wedges falling from the roof or sliding out of the excavating sidewalls (Hoek et al., 1995). These wedges are formed by intersecting structural features, such as bedding planes, joints, and other structural fractures which separate the rockmasses into interlocking pieces. When rockmasses are excavated, a free face will be created. If the bounding planes are continuous or rock bridges along the discontinuities are broken, any wedges located next to the free face may fall or slide from the surface by gravity, due to the removal of the restraint from the surrounding rock (Hoek et al., 1995). Therefore, discontinuities generally have a significant effect on the stability of coal and rockmass. Brady and Brown (2004) state that “the term discontinuity is commonly used in rock mechanics as a collective term for all fractures or features in rockmasses such as joints, bedding planes, shears, faults or dykes, and contacts that have low tensile strength”. The terminology used in this thesis is limited to the minor fractures that affect the stability assessment of top coal and rockmasses within longwall panels. It includes bedding and cleats/joints. The larger scales of fractures or features, such as the faults or dykes that influence overall mining activity are outside the scope of the current study. Bedding was a factor that was consistently cited as causing roof problems in coal mines, especially when the rockmass consists of weak laminated layers of shale and/or thinly interbedded sandstone and shale. This is because the bedding 112 planes of these rock types are not only closely spaced, but also the bedding surface is very weak (Mark and Molinda, 2005). Brady and Brown (2004) point to the slip on the bedding that causes inward displacement towards the span centreline of the bed, which tends to detach the lower bed from the one immidiately above it. Separation at the bedding plane of the lower coal bed from its upper neighbour indicates the loss of roof support by the overlying beds. Cleats in the coal seam are another parameter that has a significant effect on caving stimulation in underground mining. The attribute of cleats closely resem- bles those fractures termed as ’joints’ in other rock types (Ting, 1977; Laubach, Marrett, Olson and Scott, 1998). Cleats are fractures that usually occur in two sets that are, in most instances, mutually perpendicular and also perpendicular to bedding (Laubach et al., 1998). Cleats commonly occur without any observable shear offset, and are thus termed opening-mode fractures rather than faults, and they generally are planar (Laubach et al., 1998). Cleats are sub-vertical in flat lying beds, as illustrated in Figure 5.4.

Figure 5.4: Schematic illustration of coal cleat geometries (Laubach et al., 1998)

The analysis of the effect of discontinuities on the stability of roof rockmasses in many geomechanical classifications systems, such as the ’Q’ Index (Barton, Lien and Lunde, 1974) , Rock Mass Rating (RMR) (Bieniawski, 1973; Bieniawski, 1976), and Coal Mine Roof Rating (CMRR) (Mark and Molinda, 2005), involves the assessment of the shear strength of discontinuities. In rock mechanics, the Mohr-Coulomb failure criterion is commonly used to describe the shear strength of 113 discontinuities, in which the peak shear strength, τ is defined by the equation 5.1 (Hoek et al., 1995; Brady and Brown, 2004):

τ = c + σ0 · tanφ (5.1)

where:

c = cohesion strength of the failure surface σ0 = eﬀective normal stress acting on the failure surface φ = friction angle of the failure surface.

For the discontinuities containing no infilling, cohesion will be zero and strength will be defined by the friction angle. The shear performance of the discontinuity surfaces depends on the combined effects of the surface roughness, the rock strength at the surface, the applied normal stress and the amount of shear displacement (Wyllie and Mah, 2004). The rough surface, which is initially undis- turbed and interlocked, will have a peak friction angle. For the case when normal stress and displacement increase, the asperities will be sheared off, and the friction angle will be reduced, as illustrated in Figure 5.5.

Figure 5.5: Eﬀect of surface roughness and normal stress on the friction angle of the discontinuity surface (after Transportation Research Board, 1996) 114

For the discontinuities containing an infilling, both the cohesion and the friction angle of the surface are influenced by the thickness and the properties of the infilling. For example, the presence of gouge or clay layers can decrease both stiffness and shear strength, while vein materials, such as quartz, can serve to increase shear strength (Brady and Brown, 2004). Mark and Molinda (2005) indicated that roughness has a significant effect on shear strength only when cohesion is moderate or weak. When the discontinuities have slickenside surfaces or the strength of the bedding is most likely equal to or greater than that of the intact rock, the surface roughness has little effect on the shear strength of the discontinuities. An illustration of the effects of discontinuity characteristics on shear strength is shown by the Mohr diagram in Figure 5.6.

Figure 5.6: Relationships between shear and normal stress on sliding surface for diﬀerent discontinuity geological conditions (modiﬁed after Transportation Research Board, 1996)

The appearance of the discontinuities certainly affects caving performance, both for top coal and roof rockmass. Discontinuities with low shear strength could enhance cavability as they have a low resistance to in-plane shear stress and have a greater tendency to unravel (Mahtab and Dixon, 1976). This is true for coal seams that demonstrate the two principal engineering properties of bedding planes which are significant in underground mining context. These are (Brady and Brown, 2004): firstly, low or zero tensile strength in the direction 115 perpendicular to the bedding plane, and secondly, a relatively low shear strength of surface compared with that of the intact rock. Both of these properties play an important role in the failure of mining activity as well as of the top coal in the LTCC method. While these properties are more important for discontinuities that are widely spaced, spacing between the bedding planes or cleats is also crucial in a high density environment of discontinuities, like coal seams. The more closely spaced the discontinuities, the greater the weakening effect they have on the cavability of top coal and the rock mass (Mark and Molinda, 2005).

5.2.3 The cover depth of the seam

The cover depth of the coal seam is related to the level of vertical stress on the top of the coal seams. In a uniform layer of coal or rock with a free surface, the vertical stress of the rockmass strata is usually determined by gravity, as illustrated in Equation 5.2 as follows:

Pv = g · γ · H (5.2)

where:

Pv = the vertical stress g = gravity γ = the material density H = the depth of cover.

Before mining operations, the stress destribution around a coal seam is in fundamental equilibrium. When mining below ground, the average vertical load is not removed and it must be equal to the initial cover load (Wilson, 1982). There- fore, a local decrease in the load in the particular area induced by mining must be balanced by an equivalent increase in the load nearby. Studies by various researchers (Whittaker and Singh, 1979b; Whittaker, 1982; Wilson, 1982; Xu, 2004) have shown that longwall mining operations give rise to the major redistribution of strata pressures. The re-distribution of stress in front of the longwall 116 face will form four main vertical stress distributions (Figure 5.7), which are (Wilson, 1982; Peng and Chiang, 1984):

Figure 5.7: Vertical stress re-distribution in an across section at the middle of the longwall face (modiﬁed after Whittaker, 1974)

• Zone 1: The zone where the pressure formed from the load of the overburden weight presses on the caved fragments. Whether or not the gob pressure reaches the overburden weight depends on panel width and the distance from the gob area to the longwall face. The broken fragments near the longwall are subjected to almost no pressure from the upper strata. As the face advances further away, the broken rock from the goaf area is gradually subjected to the load of the weight from the overburden and/or the downward sagging of the upper strata.

• Zone 2: The yielding zone. In this zone there is a reduction in vertical stress associated with the fracture of coal and rockmass materials. Here the strata has passed its failure limit (Wilson, 1982).

• Zone 3: The front abutment strata pressure area. The magnitude of the abutment stress is not accurately known, but is approximately between 0.2 and 6.4 times the vertical stress, depending on the geological conditions, face location and the adjacent mined-out area (Peng and Chiang, 1984). 117

This abutment stress zone can be beneﬁcial in promoting the weakening of the coal front (Whittaker, 1982).

• Zone 4. The zone where the strata pressure slowly decreases to cover load.

From the nature of the stress distribution around the longwall face, it can be seen that the cover depth plays an important role in the deformation of top coal in front of the longwall face. According to Wilson (1982), a “yield zone” occupied by strata exists between the longwall face and the peak of the abutment stress. The nature of the yielded rock will vary across the yield zone. Near the face line it will exhibit visible fractures, but close to the abutment peak its behavior may be more plastic (Wilson, 1982). These fractures in the top coal and roof rockmass will enable them cave easier.

Figure 5.8: Area involved in Load balance (after Wilson, 1982)

The width of the yield zone is determined by the following equation (Wilson, 1982): " 1 # M q k−1 x = · − 1 (5.3) b 2 p + p0 where:

M = the height of opening q = cover load p = horizontal pressure supplied by the ribside support p’ = the unixial compressive strength of the failed broken materials k = triaxial stress factor. 118

Equation 5.3 indicates that the cover depth has a large effect on the magnitude of vertical stress as well as of the abutment pressure. The more the cover depth, the greater will be the magnitude of the abutment peak. As a result, the possibility of the deformation of the top coal in front of the longwall face is greater. A field investigation on the fracture behaviour in different zones of top coal in front of an LTCC face undertaken at Wangzhuan mine, the Luan Group in China (cited in Xu, 2004, no original reference) clearly shows the effect of the cover depth on the deformation of top coal in the LTCC method, as presented in Figure 5.9.

Figure 5.9: Fracture zones in top coal (after Xu, 2004) I=Intact zones II=Fracture development zones III=Fracture extension zone IV=fragmenting and broken zones

5.2.4 Overlying strata

Clearly, the properties of rockmasses in the stratigraphic sequence around the seams also play an important role in the cavability of top coal in the LTCC method. According to Peng and Chiang (1984), although stratigraphic sequences above the coal seam vary from seam to seam, the following are possible outcomes of overburden movements of the rockmass above a conventional longwall face:

• Type 1 stratigraphic sequence: The immediate roof is thick enough to ﬁll 119

the whole gob space where coal was extracted. In this case the main roof is supported by the caved rock piles and maintains equilibrium. Sometimes the immediate roof is not suﬃciently thick, and the main roof is well jointed. Its sagging will have little eﬀect on the roof pressure at the face area. As a result, the longwall face supports the weight of the fracture immediate roof.

• Type 2 stratigraphic sequence: The immediate roof is not thick enough to ﬁll the gob space, and leaves a gap between the top of the caved rock piles and the lower border of the uncaved strata in the main roof. Under such conditions, the main roof will break periodically at a certain distance. This type of stratigraphic sequence will induce clear period weighting.

• Type 3 stratigraphic sequence: The immediate roof is hard and strong but not very thick. It breaks only after overhanging a certain distance in the gob as a cantilever. This type of stratigraphic sequence induces a strong period weighting.

• Type 4 stratigraphic sequence: The immediate roof is massive, very hard and strong and only caves after overhanging in the gob over a large area. When caving it causes strong and storming winds that may cause roof stability problems and safety hazards to the people and mining equipment in the longwall face.

• Type 5 stratigraphic sequence: Here, the immediate roof is strong but very well jointed. During the advancement of the longwall face, the immediate roof sags gradually until it touches to the ﬂoor in the gob and forms a semi- arch. Although the semi-arch also breaks periodically, it is usually strong enough for these breaks to be detected. This type of sagging roof is more likely to occur in thinner seams.

Compared to the conventional longwall method, where all broken rockmasses remain in the goaf, part or all of the fragmented top coal will be drawn in parallel with the extraction of the cut section of the LTCC. As a result, the space required for by the caved fragments in the LTCC method is larger than in conventional 120 longwalling. Therefore, from an operational and safety point of view, the LTCC method should be applied to thick seams which have stratigraphic sequences ranging from type 1 to type 3. If the LTCC method is applied to thick seams which have the stratigraphic sequences of type 4 or 5, it could cause potential hazards to mining equipment and personnel working in the longwall face. The literature review revealed that stratigraphic sequences have a large effect on the cavability of top coal. During mining operations of the lower section of the thick coal seam, the immediate roof, depending on the rock type, would cave immediately or with little delay to the cave of the top coal. The delay time for separation and downward sagging between adjacent strata depends primarily on the thickness and the strength of the strata (Peng and Chiang, 1984). In the case where the immediate roof is weak and thin, due to the load from the upper strata and its own weight, the extraction of the lower section of a thick seam will cause a convergence of the immediate roof to the top coal section. The increasing stress from the immediate roof could create cracks or fractures in the top coal and, as a result, make the top coal collapse easier. Conversely, if the immediate roof is strong and thick, it can resist the load from the upper strata, and delay or decrease the convergence of the immediate roof to the top coal section. As a result, the top coal layer can resist for a longer period of time before collapsing. This will affect the cavability of the top coal during mining operations. The movement of the main roof also affects the stability of the immediate roof and thus the top coal section in the face area. The main roof generally breaks periodically along the direction of the face advance and imposes periodic roof weighting on the face area (Peng and Chiang, 1984). If the bedding layers in the main roof are weak and thin, they will cave simultaneously with the immediate roof. As a result, the primary effect of the main roof on the movement of top coal is the weight of the broken rockmass of the main roof. However, if the main roof caves with periodic breakage, its movement may cause a more significant effect on the movement of the immediate roof because in addition to bearing the weight of the immediate roof, the top coal also intermittently, has to take the weight of the main roof. 121

5.2.5 Top coal thickness

The characteristics of the top coal play a crucial part in the success of the LTCC method (Xu, 2004; Vakili et al., 2007). In this method, it is desirable that following coal extraction from the cut section of a thick seam, the top coal will collapse immediately after the advancement of the supports. In addition, the caved coal should be small in size so that it can be easily drawn from the longwall face. If the top coal is hanging and only caves after a certain time, it may cause coal loss because the caved coal will be too far away from the reach of the miners who stand within the longwall face protected by the supports. Another important property of the top coal is that it must be strong enough to withstand collapsing for certain period of time when undercut by the cutting machine in order to maintain stability before being supported by mining equipment. The process of fracturing and collapsing the top coal takes place in different stages of mining operations. In the first stage, when the longwall face approaches, the top coal section is subjected to abutment stress induced by the extraction of the cut section of the thick seam. In the second stage, when the lower section has been cut by the cutting machine, the top coal section, which might have been fractured by the abutment stress, is fractured and collapsed due to the load of the upper layer rockmass. Du (2003), in a literature review of the results from research undertaken by the VNIMI institute in Russia on the effect of seam thickness on the cavability of top coal, states that if the top coal section is considered as a beam, there are components of both external and internal forces that may cause the failure of top coal, as illustrated in Figure 5.10. They include the loading force from the upper roof rockmass, the weight of the top coal itself, the loading force from the support, and the force of the broken rockmass from the goaf area acting on the top coal. The loading force from the goaf area can help improve the stability of the top coal, but this force appears only when the broken rock fills all the gaps in the goaf area and it is impossible to determine the value of the loading force from the goaf area. 122

Figure 5.10: Loading forces acting on top coal (after Du, 2003)

As seen in Figure 5.10, the failure of the top coal will occur when it can not resist the loading forces acting on it. The resistance of the top coal depends on the coal’s strength and its thickness. A thinner top coal is more likely to fail compared to thicker coal in the case where the strength of the coal is the same. The presence and nature of one or several band layers in the coal seam also aﬀects the cavability of the top coal. A band layer in a coal seam is a thin layer of rockmass present within the coal seam. If the band is soft rock, it can encourage the cavability of the top coal. In contrast, if the band is created by a strong rockmass, it may prevent the cavability of the top coal.

5.2.6 Pre-mining stress regime

The design of an underground structure is undertaken in the in-situ rockmass when initial stress has been established prior to excavation. Therefore, understanding the eﬀects of the pre-mining state on problems of underground stability is necessary for design analysis. The widely known concept indicates that the in-situ state of stress is composed of two stress components: vertical and horizontal. The vertical stress component, at a given sub-surface point in underground, is generated by the weight of overburden (σv = γ.g.H). The horizontal stress components include both major and 123 minor stress components. A common but unjustiﬁed assumption is that the horizontal normal stress components occur as a result of Poisson’s restraint and are given by (Brady and Brown, 2004):

ν ! σ = · σ (5.4) h (1 − ν) v where:

σh = The horizontal normal stress ν = Poission’s ratio for the rock mass

σv = The vertical normal stress.

Observations of field investigations (Hooker and Johnson, 1969; Brown and Hoek, 1978) indicate that for the determination of depths of stress, equation 5.4 is rarely satisfied, and the vertical is seldom a principal stress direction. According to Brady and Brown (2004), the ambient state of stress in an element of rock in the ground is determined by both the current loading conditions and the stress path defined by its geological history. Mechanical processes such as fracture generation, slipping on a fracture surface and viscoplastic flow through the medium, can produce both complex and heterogeneous states of stress. The effect of horizontal stress on roof stability has been intensively studied by a number of researchers (Mark, Mucho and Dolinar, 1998; Chen, 1999; Mark and Molinda, 2005; Whittles et al., 2006). These studies state that the main problems of roof stability attributed to high horizontal stresses are compressive- type roof failures, commonly called “cutter roof”, and “guttering”, as illustrated in Figure 5.11. There is also an argument which maintains that although horizontal stress is a cause of cutter roof failure, it constrains the transformation of loading from the upper to the lower strata, resulting in less cavability of the top coal and rockmass above. Kendorski (1978) states that horizontal stress existed in rockmass can not only aid the caving process by overwhelming the tendency to form tension above the undercut area, but it also can influence cavability by locking rock blocks 124 together and cancelling tensional forces that would otherwise favour caving. As a result, this issue needs to be further addressed.

Figure 5.11: Roof instability due to horizontal stress (after Kripakov, 1982)

5.2.7 Moisture sensitivity

According to Wyllie and Mah (2004), the presence of ground water in the rockmass can have a detrimental eﬀect on roof stability because water pressure decreases the strength of the rockmass by reducing the shear strength of the discontinuities. The presence of ground water in the rockmass can cause changes in the moisture content of some rock, particularly shale, which results in diminish- ment of the shear strength in discontinuities. The appearance of ground water may encourage potential slip, or separation at the bedding planes when they are the common discontinuities in sedimentary rockmasses. Therefore, moisture sensitivity is also an important factor inﬂuencing the cavability of the top coal. 125

5.2.8 Discussion

Over the last decades, much research has been conducted on the caving mechanism of the top coal in the LTCC method and they have successfully improved the understanding of this topic. Parameters that have a significant effect on the cavability of top coal are: coal strength, persistence and intensity of discontinuities, roof strata, top coal thickness, the cover depth of the coal seam, horizontal stress, and the moisture sensitivity of the coal seam. Those studies, however, concentrated on horizontal seams, and consequently, the effect of seam dip has not been investigated thoroughly. The following geotechnical questions arise for the extraction of inclined thick coal seams: (1) Are there any differences in the caving mechanism of the top coal and above rockmass in different directions of the longwall face? (2) Are there any differences of the caving mechanism between horizontal and inclined coal seams? (3) How important is the seam dip factor, compared to others on the cavability of top coal and the roof rockmass? These questions will be discussed in the following sections by reference to results from numerical analysis. Prior to discussing these geotechnical issues, it is considered appropriate to briefly review numerical analyses of rock mechanics for underground mining, in order to gain an appreciation for the extent of these analyses.

5.3 Numerical analyses for underground mining

In underground mining, numerous efforts have been made to study and predict rock mass behavior generated by mining activity. Empirical approaches in problem solving in mining engineering are based on past experiences, and trial and error procedures (Pande, Beer and Williams, 1990). Mines designed by empirical techniques tend to be over or under safe and are applicable only to situations similar to those developed by the designers. Conventional attempts for parameter studies in mine design involve studies by physical models. The aim of physi- 126 cal modelling is to study and develop predictive capacities for application in the real life situations represented by the models. Brady and Brown (2004) point out that the difficulty in this procedure is maintaining similitude in the material properties and load applied to the models. They also note that physical modelling is more suitable for basic research rather than complex design applications. Another disadvantage of physical modelling is the expense and time needed to design, construct and test models in order to answer specific mine design questions. As a result, Brady and Brown (2004) concluded that physical modelling is limited in its potential as a predictive tool in mine designs. Engineers today may face many complicated problems with no past experience. Although rockmass behaviour around complex excavation shapes may be obtained from closed form solutions that approximating simple shapes, it is sometimes necessary to seek a detailed understanding of rock behaviour for more complicated configurations within different geological conditions. Therefore, conventional physical analysis clearly faces difficulties in investigating excavation design problems in mining engineering. With the rapid development of computer technology over the last few decades, numerical methods of analysis have been developed to expand and diversify into all the major fields of scientific and engineering studies, including mining engineering. They provide an alternative approach to physical models that require a significant amount of time and expense (Pande et al., 1990). According to Lorig and Varona (2004), numerical methods are a prominent tool used to investigate rock mechanics because of three main reasons: Firstly, numerical analyses can be conducted for the multiple possibilities of geological models, failure modes and design options. Secondly, numerical methods can incorporate comprehensive and more realistic information about key geological features, and therefore assist researchers in understanding of rockmass behaviour more exhaustively than do other analytical methods. In addition, compared to empirical methods, numerical models can be extrapolated outside their database. Starfield and Cundall (1988) argue that because measurements and experiments in rock mass are costly, difficult to perform or even impossible to carry out, numerical modelling, as an 127 alternative form of experimentation, is an essential ingredient of both science and design disciplines. For these reasons, numerical methods are becoming increas- ingly popular for solving problems in rock mechanics for underground mining where enough geological information is not always available or cannot easily be obtained.

5.3.1 Types of numerical methods

According to Jing and Hudson (2002), the three following numerical methods are most commonly applied to problems of rock mechanics:

• Continuum methods, including the finite differential method, the finite element method and the boundary element method.

• Discontinuum methods, including the discrete element method, and the discrete fracture network method,

• Hybrid continuum/discrete methods.

In order to choose the appropriate numerical code for a realistic presentation of speciﬁc geological conditions, it is necessary to know the principle of each numerical method. The fundamentals and application each method have been discussed in several books and articles, including Pande et al. (1990) Hoek et al. (1995), Jing and Hudson (2002), and Jing (2003). Therefore, it is not further review within this thesis.

5.3.2 The numerical methods used in the current study

The choice of suitable numerical methods for rock mechanics depends on many factors, and mainly on the specific problems, scale, and the discontinuities system geometry used. Continuum methods are suitable for analysis where few discontinuities are present and if fracture opening and complete block detachment are insignificant. Discontinuum methods are well suited for representing a highly jointed rockmass, and allow the representation of the detachment of elements 128 in the models. In some cases, some disadvantages of each type can be avoided by combining continuum-discontinuum methods in hybrid models. Due to the ability of discontinuum methods to present the detachment of elements in the model, required for cavability analysis, the 3DEC software, which is based on discontinuum element methods and developed by the Itasca Consulting Group (2002) was chosen for this study’s assessment of the caving mechanism of top coal in the LTCC method. It was soon recognized that the simulation of a detailed model of the jointed rockmass by the 3DEC software would be very time consuming. For that reason, the investigation of specific problems that did not need three dimensional models and for models requiring more details of jointed rockmass systems, the UDEC programme complemented the 3DEC programme. The UDEC programme, which was also developed by the Itasca Consulting Group (2004) is a two dimensional discontinuum element method. In addition, due to the large advantages in stress analysis and faster solution times of continuum methods, the FLAC programme, which is a finite differential method and, again, developed by the Itasca Consult- ing Group (2005), was used for stress analysis in this study.

5.4 The eﬀect of face orientation on the cavability of inclined thick seams

5.4.1 Scope of the modelling

This study aims to investigate the eﬀect of the orientation of the longwall face on the caving mechanism of top coal and above rockmass when extracting inclined thick seams using the LTCC method. The study focuses on analyzing the diﬀer- ences in the caving mechanism when the longwall face retreats in three directions: along strike, up-dip, and down-dip by reference to the results of numerical analysis. The UDEC software developed by the Itasca Consulting Group (2004) was used for assessing the caving mechanism of top coal in the LTCC method. This numerical modelling software is based on discontinuum methods and has been 129 discussed in detailed in Section 5.3.1.

5.4.2 Mining geometry

A total of 3 geometric models were simulated to represent the cross-section at the centre point of the longwall face with diﬀerent seam dips of 00, 150, and 300. The diagram showing the location of the UDEC analysis is presented in Figure 5.12.

Figure 5.12: Diagram shows the location of the UDEC analysis

The models consisted of 9 different layers, including a coal seam and a number of rockmass layers. Detailed strata both above and below the coal seams in the models are presented in Figure 5.13. The coal seam has a thickness of 6.0m and was divided into two sections: the cut and the top coal. The height of both sections was assumed to be 3.0m. To represent bedding and cleats, the entire top coal section was divided into 6 bedding layers in parallel with the seam and rockmass layers, with the thickness of 0.5m for each layer. The cleats were created in each layer in a regular rectangular pattern with a cleat spacing of 0.5m, as illustrated in Figure 5.13. This type of joint pattern is commonly observed in Quangninh coalfield as well as some other coalfields (Laubach et al., 1998; Su, Feng, Chen and Pan, 2001). 130

Figure 5.13: Overall context of UDEC model

5.4.3 Material properties and constitutive models

The properties of the coal and roof rockmass were obtained from both laboratory testing and the available references from the Vietnamese mining industry (Sy, 1999; Sy, 2002). It is assumed in the UDEC analysis, that the failure mechanism in coal containing a high density of discontinuities is controlled by their presence. Therefore, under the current study, all blocks in the models have been simulated elastically, but joint contacts represent nonlinear constitutive behaviour. The UDEC programme requires material properties for both the intact blocks and the discontinuities. All material models for deformable blocks assume the behavior of an isotropic material in the elastic range and are described by two elastic constants, bulk modulus, K, and shear modulus, G. The equations to convert from Young’s modulus, E, and Poisson’s ratio, ν to bulk modulus, K, and shear modulus, G are (Itasca Consulting Group, 2004):

E K = (5.5) 3(1 − 2ν)

E G = (5.6) 2(1 + ν) One parameter of the discontinuity properties that needs to be assigned in 131 the UDEC programme is the discontinuity stiffness (normal and shear stiffness). Often, the only way to guide the choice of appropriate parameters is by comparison to similar joint properties derived from field tests. However, values of normal and shear stiffnesses for coal and rock discontinuities were not available in the Quangninh coalfield. Published data (Kulhawy, 1975; Rosso, 1976; Ban- dis, Lumsden and Barton, 1983) indicated that the values of normal and shear stiffnesses for rock joints range from approximately 10 to 100 MPa/m for joints with soft clay in-filling, to over 100 GPa/m for tight joints. In addition, there is a range of joint stiffnesses that are reasonable to use in a UDEC model. A too high stiffness requires a significantly longer solution time and the displacement of the blocks does not represent a real life situation. A too low normal stiffness, contrastingly, can cause problems during running, such as block interpenetration. A previous study which deals with the UDEC programme (Vakili, 2007) recommends that the optimized values of normal and shear stiffness of bedding planes and cleats in coal seams should be 100GPa/m and 10GPa/m, respectively. As a result, these values of joint stiffness were also used in this study. The shear strength of the discontinuities was modelled using the Mohr-coulomb failure criterion on the basic estimate provided by Vakili et al. (2007). Details of assigned shear strength parameters are given in Table 5.1. For the purpose of comparison, the properties and the stress applied for all three models were the same, with an average cover depth of 500m. In the current analysis, the initial stress condition was assumed by equation 5.7 .

σv = γgH (5.7) and

σh = σv (5.8) where:

σv = the vertical stress

σh = the horizontal stress g = gravity 132

γ = density of coal or rockmass H = the cover depth.

Table 5.1: Materials and joint properties of diﬀerent layers used in the models (after Vakili, 2007)

Density Bulk Shear Joint Joint friction Joint tensile Material (T/m3) Modulus Modulus cohesion angle strength (GPa) (GPa) (MPa) (degree) (MPa) Coal 1.6 varies varies 0.002 15 0 immediate roof 2.6 2.45 1.47 0.002 20 0.002 main roof 2.65 3.02 1.81 0.005 20 0.005 immediate ﬂoor 2.6 2.45 1.47 0.002 20 0.002 main ﬂoor 2.65 3.02 1.81 0.005 20 0.005

Because there is no information on horizontal stress available in the Quangn- inh coalfield, various values of the ratio between the horizontal and vertical stresses of 0.5, 1, and 2 were assigned during sensitivity analysis. For all the models, the horizontal stress refers to the principal horizontal stress and was assumed to be perpendicular to the vertical stress. In addition, all the stress paths simulated in the models were assumed to be perpendicular to the face direction. It should be noted that the depth of cover in the inclined seam presented in numerical modelling was determined by the average depth of the panel length. As a result, when the longwall face retreats up-dip, the depth of cover in the first stage of the face retreat will be greater than that of the flat seam. In the case where the longwall face retreats down-dip, the depth of cover in the first stage of the retreat will be less than that of the cover depth in the face retreating along the strike.

5.4.4 Simulation of face extraction and top-coal caving

Mining was simulated by removing 1.0m slices at the bottom of coal seam (3.0m in height), followed by advancing the supports. The face supports were modelled 133 by combining two elements with a high Young’s modulus: the canopy and the base. A FISH function (a programming language embedded within the UDEC to enable user defined tasks) was created to simulate the advancement of the supports. In order to investigate differences in the caving mechanism of the models, the following two monitoring tools, adapted from a study undertaken by Vakili et al. (2007), were created to measure the cavability in different geotechnical conditions, as illustrated in Figure 5.14:

• Top coal caving distance (CD). This is deﬁned as the distance from the ﬁrst line of the opening entry to the face distance where the whole thickness of top coal section starts to cave.

• Total average vertical displacement of top coal (TAVD). After every two shears, the average vertical displacement of all blocks in the top coal section, within the extracted longwall panel was calculated. The interval panel has a length ranging from the edge of the back pillar to the faceline of the longwall face. The TAVD was then determined by averaging all the average vertical displacements of the top coal in the panel, as presented in Equation 5.9.

Figure 5.14: Monitoring tools used for the investigation of caving mechanism 134

n X 1 T AV D = AV Di · (5.9) i=1 n where:

AV Di = the average vertical displacement of top coal blocks at investigation number i n = number of investigation points.

5.4.5 Discussion of numerical modelling results

Results from the UDEC analysis show that the orientation of the longwall face has a large effect on the cavability of the top coal for the extraction of inclined thick seams by the LTCC method. Figure 5.15 represents the cross-section at the centre point of the longwall face retreating in two directions: along the strike and in the up-dip direction (300) at different retreat distances. Results in Figure 5.15 indicate that, when extracting an inclined thick seam, a face retreating along the strike has a better top coal caving performance compared to a face retreating up-dip. The differences in the caving mechanisms of a face retreating along the strike and a face retreating up seam dip are not only seen in the top coal section of the seam, but also in the rockmass above the seam. As can be seen in Figure 5.15, at a 60m retreat distance, when the face is retreating along the strike, the immediate roof and a part of the main roof caves completely, while in the case of the face retreating up-dip and also at a 60m retreat distance, only a small area of the immediate roof collapses, while the main roof has not yet started caving. This comparison suggests that an overall better caving performance is anticipated when the face retreats along strike as compared to a face that retreats up dip. 135

00, L = 20m 300, L = 20m

00, L = 40m 300, L = 40m

00, L = 60m 300, L = 60m

Figure 5.15: Diﬀerences in the cavability of top coal between two face directions at diﬀerent face retreat distances (L) 136

150, L = 20m 300, L = 20m

150, L = 40m 300, L = 40m

150, L = 60m 300, L = 60m

Figure 5.16: The cavability of top coal for a face advancing down the dip at diﬀerent face retreat distances (L) 137

To investigate the cavability of the top coal in the LTCC method when a face retreats down the dip of the seam, two models that represent a face retreating down-dip 150 and 300 were simulated. Results from the numerical modelling also indicate that a face retreating along the strike gives better caving performance compared to a face retreating down dip. Figure 5.16 represents a face mining down-dip at 150 and 300, at different retreat distances. The effect of face orientation on the cavability of top coal in the LTCC method can also be shown by two monitoring tools presented earlier: CD and TAVD. Results from the UDEC analysis indicate that the caving distance is shorter when the longwall retreats along the strike when compared to that of a face retreating upward or downward on the seam dip. In addition, a face that retreats along the strike has a greater value of TAVD than that of a face that retreats up-dip. Table 5.2 presents the relationship between the three different directions of the LTCC face in relation to the CD and TAVD.

Table 5.2: CD and TAVD for diﬀerent directions of longwall Face direction Seam dip CD TAVD (degree) (m) (m) Along strike 0 22 1.62 Up-dip 15 25 1.48 Up-dip 30 28 1.02 Down-dip 15 27 1.39 Down-dip 30 30 0.89

As can be seen in Table 5.2, differences in CD and TAVD between the face retreating up-dip and down-dip is probable because of differences in the depth of cover at the initial opening of the panels between the two scenarios. As discussed previously, the depth of cover in the inclined seam, represented in the numerical model, was determined by the average depth of the panel length. As a result, when the longwall face retreats up-dip, the cover depth in the first stage of the face will be greater than that of the face retreating along strike. In contrast, in the case of the face retreating down-dip, the initial opening of the face is under 138 a lower cover depth than that of a face retreating along the strike, resulting in differences in CD and TAVD between the face retreating up-dip and down-dip. Two more detailed models, which simulated a greater length of longwall panel in a flat (00) and an inclined seam (300), further confirmed the conclusion that when mining an inclined thick seam, a face retreating along the strike gives a better caving performance compared to a face retreating up-dip or down-dip of the inclined seams. Figure 5.17 presents the differences in the caving mechanisms of the top coal and the rockmass above, at a panel length of 100m.

a. 00, L = 100m

b. 300, L = 100m

Figure 5.17: Cavability of the top coal at the retreat distance of 100m with diﬀerent seam dips

The failure mechanism is certainly a complex process and is influenced by 139 many factors. Results from the UDEC analysis show that horizontal stress also has an effect on the caving mechanism of the top coal and the rockmass above. In the first phase of movement from the setup entry to the distance where the top coal section starts to cave, the higher the horizontal stress, the easier the cavability of the top coal. For instance, with k = 0.5 (where k is the ratio between the horizontal and vertical stress), the distance of the top coal (CD) for flat seam is 26m. When k = 1, the CD value is 22m, and the CD value decreases to 15m if k = 2. These results indicate that horizontal stress applies the load normally, causing buckling of both the top coal and the rockmass. These results agree with the observations widely reported in the past. However, after the first weighting of the top coal and/or immediate roof, as the face proceeds, the lower the horizontal to vertical stress ratio, the better the caving performance. These results from the UDEC analysis support the argument that although horizontal stress is a cause of cutter roof failure in coal mine roof, it constrains the transformation of loading from the upper to the lower strata, which results in less cavability of the top coal and above rockmass. Figure 5.18 compares the differences in the caving mechanisms of the top coal and the above rockmass for flat and inclined (300) seams with different values of the horizontal to vertical stress ratio. The strength of the coal is also a factor that affects the cavability of the top coal and the rockmass above. Results from the UDEC analysis show that the lower the strength of the coal, the better the caving performance. However, all the results suggest that a better caving performance is predicted when a face retreats along the strike compared to a face retreating either up-dip or down-dip of the seam. Table 5.3 presents the correlation between the various influencing parameters and the seam dip to the CD and TAVD values. It should be noted that the heights of the longwall face in all three seam dip scenarios were assumed to be the same, 3.0m. However the vertical distance from the roof to the floor of the longwall face in the inclined coal seam was slightly higher than that of the flat lying seams. 140

a. 00, k = 0.5 d. 300, k = 0.5

b. 00, k = 1 e. 300, k = 1

c. 00, k = 2 f. 300, k = 2

Figure 5.18: Eﬀect of horizontal stress on the cavability of the top coal with diﬀerent seam dips (@60m of retreat distance) 141

The analysis of the UDEC programme also can be used to represent the cross-section just behind the support for the face retreating along the strike, as illustrated in Figure 5.19. In this analysis, the simulation of the models is similar to the investigation of the eﬀect of the seam dip of the face on cavability, as described in section 5.4.2. The only diﬀerence is that the support is simulated in the models.

Table 5.3: CD and TAVD for diﬀerent inﬂuenced parameters

Young’s Seam dip Horizontal/ CD TAVD modulus (GPa) (degree) vertical ratio (m) (m) 3.0 0 1 22 1.62 3.0 15 1 25 1.48 3.0 30 1 28 1.02 4.53 0 1 24 1.42 4.53 15 1 31 1.18 4.53 30 1 34 0.85 3.0 0 0.5 26 1.47 3.0 15 0.5 30 1.34 3.0 30 0.5 31 1.39 3.0 0 2 15 1.70 3.0 15 2 17 1.72 3.0 30 2 19 1.59

Figure 5.19: Diagram shows the location of the UDEC analysis for the face retreating along strike 142

The results from the UDEC analysis indicate that when the face retreats along the strike, better caving mechanism of the top coal and above rockmass is anticipated in a ﬂat coal seam as compared to an inclined ones, as seen in Figure 5.20. Also from Figure 5.20, it can be seen that for ﬂat seams, the cavability of the top coal near the maingate and tailgate is similar, and better cavability is anticipated in the middle of the longwall face. For inclined seams, better cavability of the top coal is predicted in the uphill portion of the longwall face compared to the downhill ones.

a. ﬂat coal seam (00)

b. inclined coal seam (300)

Figure 5.20: Cavability between ﬂat and inclined seams

In an attempt to understand the reasons for diﬀerences in the caving mech- 143 anisms of top coal between ﬂat and inclined thick seams, the stress distribution around the longwall face was investigated as it is believed that the redistribution of the stresses has a major impact upon roof stability.

a. 00

b. 300 Figure 5.21: Diﬀerences in vertical stress distribution around the longwall face between two face directions (at 40m of retreat distance)

As can be seen in Figure 5.21, where the faces retreat along the strike, the vertical stress of ﬂat seams is distributed symmetrically on both sides of the longwall panel, and the failure zone is located in the middle of a longwall panel. 144

For inclined seams, vertical stress tends to concentrate on the downhill portion rather than the uphill portion of the longwall panel. These results indicate that for inclined seams, the concentration of vertical stress is divided into two loading components. The first is normal force acting perpendicularly to the lower strata and the second is the sliding of the rockmass layers, which concentrates on the downward chain pillar of the longwall panel. As a result, the force unloading the lower strata of inclined seams is less than that of flat seams, causing less cavability of the top coal and rockmass on the downhill side of the longwall panel, compared to uphill. Results from the UDEC analysis agree with the results obtained from the FLAC programme which indicates that with flat coal seams, better caving is anticipated in the middle portion of the longwall face, while with inclined seams, better caving is predicted at the uphill portion of the longwall face compared to the downhill portion. Details of the FLAC analysis is provided in Section 7.3.

5.5 Conclusions

This chapter provided a detailed discussion of the effect of seam dip on the cavability of the top coal and above rockmass in the LTCC method. Before this study, investigations of the cavability of the top coal and above rockmass largely focused on horizontal thick seams. The literature review shows that many geotechnical factors influence the cavability of the top coal and upper strata. The major factors affecting the cavability of top coal are: coal strength, persistence and intensity of discontinuities, the roof strata, the thickness of coal seam, the depth of cover, and the moisture sensitivity. However, the effect of seam dip on the cavability of the top coal has not been studied thoroughly. It is clear from the UDEC analysis that the orientation of the longwall face has an effect on the cavability of the top coal and above rockmass during the extraction of an inclined thick coal seam by the LTCC method. Results from the analysis show that a better caving performance is predicted when the face retreats along the strike compared to the face retreating up-dip or down-dip of the seam. It is therefore recommended by this study that in order to extract 145 inclined thick seams, the LTCC face should retreat along the strike. In the case of the faces retreating along the strike, results from the UDEC analysis also indicates that better caving performance in flat coal seams is anticipated compared to that of inclined seams. For a flat coal seam, better cavability of the top coal and above rockmass is anticipated in the middle of the longwall panel, while in an inclined thick seam, the cavability is better in the uphill portion of a longwall panel than that of the downhill ones. Horizontal stress is also believed to be one of the factors influencing the cavability of the top coal and above rockmass through two contrasting aspects: com- presses and crushes the top coal layers which help to form tension and aid caving process. However, it also blocks the top coal and rockmass together and reduces the tension forces that otherwise encourage the caving progresses. The difference in the re-distribution of vertical stress is considered the main reason for the difference in the caving mechanism of the top coal and above rockmass between flat and inclined coal seams. In flat seams, vertical stress is distributed symmetrically on both sides of the longwall panel, and all the vertical force unloads on the lower strata causing the cavability of the top coal. For inclined seams, vertical stress, which is the main cause of the fracturing of the top coal and above rockmass, is divided into two loading components. The first is the normal force acting perpendicularly to the lower strata and the second is the sliding of the rockmass layers, which concentrates on the downward chain pillar of the longwall panel. As a result, the normal force unloading on the lower strata of the inclined seam is less than that of flat seams, causing less cavability of the top coal and rockmass on the downhill side of the longwall panel compared to that of the uphill side. As a result, better cavability of the top coal and above rockmass is anticipated in flat coal seams compared to those that are inclined. Chapter 6

Caving mechanism of the Top Coal in the LTCC method for Inclined Thick Seams

6.1 Introduction

In Chapter 5, an investigation of the eﬀect of the seam dip on the face orientation of the longwall face was undertaken by the UDEC analysis. Results from that analysis concluded that for the extraction of an inclined thick seam, the face retreating along strike has a better caving performance compared to retreat directions of up-dip or down-dip. Initial results from the UDEC analysis also pointed out that, for the face retreat along strike, better caving is predicted for a ﬂat coal seam compared to an inclined one. However, it became apparent that the UDEC programme, which is a two-dimensional discontinuum modelling code, can not fully explain the caving progression of the top coal. For inclined seams, it is impossible to simulate both the orientation of the longwall face retreating along the strike and the seam dip in UDEC models. One geotechnical question arising from the investigation of the cavability of top coal in the LTCC method is that, for the faces retreating along the strike, does the top coal section of an inclined thick seam cave within a certain distance from the longwall face when

146 147 compared to a flat face, so that the miners standing at the working area can reach the caved coal. In order to compare these differences between flat and inclined seams, it is necessary to understand the caving progression of the top coal. For that reason, a three-dimensional numerical code was chosen for simulating the caving mechanism in the top coal for the faces retreating along the strike. The aim of this analysis was to further study the following issues:

• To identify the diﬀerences in the cavability of the top coal in the LTCC face retreating along the strike between ﬂat and inclined coal seams.

• To gain a greater insight into the eﬀect of seam dip, combined with other factors, on the cavability of top coal by parametric studies from modelling analysis. The results of the study will then be compared to the statistics of top coal caving performance in typical coal mines to determine a criterion of caving assessment of top coal in the LTCC method with inclined seams.

To address the above mentioned tasks, the 3DEC programme, developed by the Itasca Consulting Group (2002), was used for this analysis. The 3DEC programme, which is based on a distinct element method, is a commercial numerical code for modelling the behavior of brittle rock, which includes fracture propagation and mechanical stability in the rockmasses around underground openings. This programme has been developed speciﬁcally to study the complex failure mechanism and has been used with much success in rock mechanics. Some of the works in which it has been used include slope stability (Corkum and Mar- tin, 2004), slope failure mechanism (Franz, 2008), and the caving mechanism of top coal (Vakili et al., 2008).

6.2 Mining geometry

A total of four geometric models were established to represent coal seams with different seam dips of 00, 150, 300, and 450. Models in 3DEC analysis were created first by a single block with a size that encompassed the physical region being analyzed. Then, this block was cut into smaller blocks whose boundaries 148 represent both geological features and engineered structures in the models. The longwall faces with different seam dips simulated in the models were assumed to all retreat along the strike. The x direction was along the strike of the longwall panel, the y direction was the vertical, and the z direction was along the seam dip. For simplicity, the maingate and tailgate were not included in the models. A typical mining geometry in 3DEC analysis is illustrated in Figure 6.1.

Figure 6.1: Typical mining geometry in the 3DEC analysis

As discussed in section 5.3.2, one of the complicated steps in the model setup was to balance the representation of joints and rockmasses for reasonable accurate modelling against the requirement of reasonable solution times. Too many details of the joints represented in the models will result in a significant solution time which is not desirable for the analyzing stage. In contrast, too few details may cause the models fail to represent a real life situation. For this reason, only the areas of interest (the top coal section and immediate roof) close to the mining openings are modelled in detail. The far-field areas are represented by a rockmass with fewer joints. By that modification, the solution time for each run of the model at current personal computer speed (CPU 3.2GHz, 2.0GB of RAM), ranges from 15 to 25 days. 149

The models consisted of 6 different layers, including a coal seam and 5 different rockmass layers. The coal seam, with different thickness values, was divided into two sections: The cut in the lower section of a thick seam has a thickness of 2.5m in all models and the rest thickness of the seam is the top coal section. The longwall panel was modelled with a width of 30m (z direction) and a length of 20m (x direction). Figure 6.2 presents the details of the longwall panel in 3DEC modelling.

Figure 6.2: Typical mining geometry in the 3DEC analysis

To represent joints and cleats in the top coal section, the thickness of the top coal was divided into bedding layers parallel with the roof of the longwall face. The cleats were created perpendicular to the bedding layers. The cleat spacing in the z direction was assumed to be 1.0m in all models, while the other cleat spacing (x direction) and the spacing of the bedding vary for sensitivity analysis.

6.3 Material properties and constitutive models

It was assumed in the 3DEC analysis that rockmass failure behaviour is completely controlled by the behaviour of discontinuities. Therefore, all blocks in the models have been simulated elastically, but joint contacts represent nonlinear constitutive behaviour. The joint constitutive model and properties in different layers were used the 150 same as in the UDEC analysis described in Chapter 5.4.3. That is, the joint area contact (Coulomb slip) model was applied, which provides a linear representation of joint stiffness and yield limit and is based upon elastic stiffness, frictional, cohesive and tensile strength properties and dilation characteristics. Assigned joint normal stiffness was chosen at 10GPa/m and shear stiffness at 1GPa/m. The properties of materials and discontinuities in different layers were presented in Table 5.1. For the purpose of comparison, the material properties and the initial stress applied to the models were kept unchanged (except for other states) with the vertical stress assumed at a given point in the underground generated by overburden, as in equation 5.7. In addition, because there are some different opinion on the relationship between the vertical and the horizontal stress as discussed in Section 5.2.6, and there is no information of that relationship in the Quangninh coalfield, the horizontal stress is assumed to be equal to the vertical stress.

6.4 Simulation of face extraction and top coal caving

After cycling the models to the equilibrium state under the influence of pre- mining stresses, the simulations were continued by multiple extractions of the longwall face. Each excavation step was simulated by removing a 1m longwall face advance (x direction) and 30m of the length of the longwall face (z direction) at the lower section of coal seam (2.5m height), followed by the advance of the supports. Then the models continued to run to achieve equilibrium. The supports of the longwall face were modelled by combining two elements of high Young’s modulus: the canopy and the base. A FISH function (a programming language embedded within 3DEC to enable user defined tasks) was created to simulate the advancement of the supports. To model the caving processes for the LTCC mining, after the processes of extracting the coal face and advancing the supports, any blocks of top coal that 151 dropped to the floor and located within 2m behind the supports, was deleted. A FISH function was created to simulate these processes. To investigate the caving mechanism of the top coal, a monitoring tool called caving performance index (CPI) was created. The CPI was determined by the ratio of the total number of deleted blocks and the total number of blocks of top coal within the interval longwall panel. The interval panel was calculated by the length ranging from the back pillar of the set up entry to the rear of the supports of the longwall face, and the width ranges from the maingate to the tailgate.

6.5 Discussion of the numerical modelling results

Although the main purpose of the 3DEC analysis was to investigate the effect of seam dip on the cavability of top coal in the LTCC method retreating along the strike, some other important parameters which critically influence the top coal were also examined in the current study to evaluate the importance of the dipping factor on the caving mechanism of top coal compared to the others. A total of 24 different mining geometry models were completed and 6 different parameters studied in the 3DEC analysis. Those are listed in Table 6.1, with the default values given in bold type.

Table 6.1: Parameters studied in the 3DEC analysis

Study parameter Variation (default in bold) Young’s modulus (MPa) 1.35 1.64 1.98 2.83 3.68 Seam dip (degree) 0 15 30 45 Seam thickness (m) 0.5 1.0 1.5 2.0 2.5 Vertical spacing (m) 0.375 0.5 0.75 1.0 Horizontal spacing (m) 0.30 0.375 0.50 0.75 Cover depth (m) 200 300 400 500

Some of the critical parameters inﬂuencing the cavability of the top coal are 152 presented in following sections below.

6.5.1 Eﬀect of the seam dip

Results from the 3DEC analysis show that the dip of the coal seam has an effect on the cavability of the top coal in the LTTC method. Figure 6.3 presents the average vertical displacement of the top coal section of the LTCC face retreating along strike in four seam dip scenarios: 00, 150, 300, and 450. As can be seen from Figure 6.3, the greater the seam dip, the less the average vertical displacement of the top coal section above the longwall face. It should be noted that the height of the longwall face in three cases was assumed to be similar. The vertical distance from the roof to the floor (the vertical displacement) of the longwall face in inclined coal seams was, however, higher than the distance in the flat ones.

Figure 6.3: Comparison of the average vertical displacement in diﬀerent seam dips

The difference in the caving mechanism between flat and inclined coal seams can be observed when comparing the number of top coal blocks remaining in the goaf area with different seam dips. Figure 6.4 presents the status of top coal blocks being left unmined in the goaf area at a retreat distance of 20m. As can be seen, the number of top coal blocks remaining in the flat seam case is smaller than those in the inclined seam case. In addition, the top coal near both sides of 153 the longwall panel caves similarly in the flat seam, while for the inclined seam, better cavability is expected in the uphill portion compared to the downhill of the longwall face. These results are similar to the results obtained from the UDEC and FLAC analyses. The difference in the caving mechanism in different portions along the longwall face can be seen by the comparison of the average vertical displacement of top coal blocks along the longwall face with different seam dips. Figure 6.5 presents the average vertical displacement of top coal blocks located within 3m behind the supports and along the longwall face at an advance distance of 15m. It can be seen from Figure 6.5 that in the flat coal seam, better caving is predicted at the middle of the longwall face, and equal caving performances of the top coal near both face ends are predicted. With inclined seams, the caving performance of top coal in the uphill portion of the face is better than in the downhill portion. The effect of seam dip on the cavability of top coal in the LTCC method can also be presented by the caving performance index (CPI). Results from the 3DEC analysis show that a higher CPI value is anticipated for flat seams than for inclined ones. Figure 6.6 shows the relationship between the CPI of the top coal section and the dip and strength of coal seam. Results indicate that the greater the dip of the coal seam, the less the cavability of the top coal section is expected. The same trend is also obtained in different scenarios of coal strength. A detailed investigation of the effect of coal strength on the cavability of the top coal is discussed in the next section. A discussion of the reasons why there are differences in the cavability of top coal and rockmass between flat and inclined seams is provided in section 5.4.5 above. The failure mechanism is certainly a complex process and influenced by many factors. Other factors such as coal strength, discontinuity spacing, the thickness of the coal seam, the depth of coal seam and their individual and combined effects are described below. 154

a. 00

b. 300

Figure 6.4: Status of top coal blocks with diﬀerent seam dips (20m of retreat distance) 155

Figure 6.5: Average vertical displacement of top coal along the longwall panel in diﬀerent seam dips

Figure 6.6: Caving perforamnce with diﬀerent seam dips 156

6.5.2 Eﬀect of the strength of coal seam

The evaluation of the eﬀect of coal strength on the cavability of top coal in the 3DEC analysis was based on comparing 5 alternative values of Youngs modulus of the coal seam. Results obtained from the 3DEC analysis show that the stronger the coal seam, the less the cavability of top coal anticipated. Figure 6.7 illustrates the relationship between the strength of the intact coal seam and the CPI of the top coal.

Figure 6.7: Coal strength vs. cavability of top coal

6.5.3 Eﬀect of discontinuity spacing

The appearance of the discontinuities will certainly affect the caving performance of both the top coal and roof rockmass. As discussed previously, in a coal stratification setting, coal seams and the other layers of rockmass lie conformably with each other, where the stratification is associated with bedding planes. There are two principal engineering properties of bedding planes which are significant in the underground mining context (Brady and Brown, 2004): The first is the low or zero tensile strength in the direction perpendicular to the bedding plane. The second is the relatively low shear strength of the surface compared with that of the intact rock. Due to the coal seams having a large density of discontinuities, 157 the spacing of vertical joint/cleats and horizontal (bedding) planes are evaluated in this study. Figure 6.8 is the analysed result of investigating the vertical spacing factor of the discontinuities that may affect the caving mechanism of the top coal. The plot predicts that the more densely the joints/cleats are spaced, the better the caving mechanism of the top coal. These results from the 3DEC analysis agree with the study undertaken by Vakili (2009).

Figure 6.8: Eﬀect of the joints/cleats spacing on the cavability of top coal

The evaluation of the eﬀect of the bedding (horizontal) spacing in the top coal was undertaken by comparing the caving performance obtained from 4 bedding spacing scenarios. Model 1 and 2 simulated thin bedded coal seams and bedding spacing was assumed to be at 0.3 and 0.375m intervals, respectively. Model 3 simulated moderately bedded top coal with the bedding spacing assumed to be at 0.5m intervals, while for model 4, the bedding spacing was at 0.75m intervals, representing a thickly bedded top coal. The 3DEC analysis shows that, for the same thickness of top coal, an early caving of top coal is anticipated with a thin bedded roof. It is expected that top coal caving, in particular the recovery rate of the caved top coal, will perform better for thin bedding spacing than for thicker bedding spacing. Figure 6.9 compares the status of the top coal after 20m of face retreat between model 3, 158 which has a bedding spacing at 0.5m intervals, and model 4 where the bedding spacing was at 0.75m intervals.

a. bedding spacing = 0.5m b. bedding spacing = 0.75m

Figure 6.9: Status of top coal blocks with diﬀerent bedding spacing (at 20m of retreat distance)

Figure 6.9b indicates that the slip on bedding causes inward displacement towards the span centreline of the beds; the tendency is detachment at the bedding plane of the lower bed from the one immediately above it. These results agree with previous studies which were described in detail by Brady and Brown (2004). Figure 6.10 shows the relationship between the bedding (horizontal) spacing factor and the CPI of the top coal. The results from the study indicate that the greater the density of the bedding planes, the better the caving mechanism of the top coal. 159

Figure 6.10: Eﬀect of the bedding (horizontal) spacing on the cavability of top coal

6.5.4 Eﬀect of the thickness of the top coal

Results from the 3DEC analysis show that the thicker the top coal, the more delayed the caving process will be, as reflected in Figure 6.11. This obviously affects the caving performance in the early stages. In practice, however, the overall impact will be minimal if operational measures are taken to improve the top coal caving performance and recovery rate. Another effect of the thickness of the top coal is that with the same caving angle, as the top coal gets thicker, in some cases, the upper portion of top coal will cave to the floor at distances too far away to be reached from the longwall face. In the LTCC method, only top coal that caves at a certain distance from the support can be drawn through the top coal drawing window. Figure 6.12 plots the relationship between the thickness of coal seam in cor- respondence with the CPI values. Notably, the caving performance improves as the top coal gets thinner. 160

a. Top coal thickness = 1.0m

b. Top coal thickness = 2.0m

Figure 6.11: Caving situation between diﬀerent top coal thickness (face advance distance 14m)

Figure 6.12: Eﬀect of the seam thickness on the caving mechanism of the top coal 161

6.5.5 Eﬀect of the depth of cover

The 3DEC analysis confirmed that the depth of cover has a large effect on the caving performance of the top coal. A clear trend indicates that poorer caving performance may be expected at a shallow depth and it is expected that the deeper the cover depth of the coal seam, the better the caving performance. The relationship between the depth of cover of the coal seam corresponding to the CPI is presented in Figure 6.13. The results indicate that the caving mechanism may experience some difficulties due to the shallow depth of cover. More detailed models may be required to further confirm the depth cover below which the cavability of the top coal will no longer be affected.

Figure 6.13: Caving performance vs depth of cover

6.6 Conclusions

The following conclusions were drawn below on the conditions that the rock properties for both coal and roof materials were kept unchanged among the models.

• It is clear that the dip of the seam has some eﬀects with regard to top coal cavability. For faces retreating along strike, a better caving performance of the top coal is predicted for ﬂat coal seams compared to inclined ones. 162

Comparing the cavability in diﬀerent parts of the top coal along the longwall face, results from the study illustrate that for a ﬂat seam case, the cavability of the top coal near both endgates of the longwall face (maingate and tailgate) is expected to be similar, and better cavability of the top coal is anticipated at the middle of the longwall face. While for an inclined seam, better caving performance in the uphill portion of the longwall face is anticipated than in the downhill portion.

• Coal strength is another important factor aﬀecting the cavability of the top coal. In general, the weaker the coal material, the better the expected caving performance result.

• Discontinuities presented in coal seams (bedding planes and cleats) also have a large impact on the caving performance of the top coal. A clear trend shows that the more closely spaced the discontinuities, the greater the weakening eﬀect on the cavability of the top coal.

• The thickness of the top coal is conﬁrmed as a critical factor which may inﬂuence the caving performance of the top coal. Normally, a thinner top coal will cave right behind the support, therefore a better recovery rate is expected. With a thicker top coal, caving at early stages may not be as good as for thinner top coal due to the delay in the caving of the upper layers of the top coal. However, as the face advance distance increases, the top coal caving performance will improve. In practice, some operational measures may be considered to improve the recovery rate for thicker top coal.

• The cavability of top coal in the LTCC method is believed to be depth dependent. The caving performance of top coal is expected to have some diﬃculty at shallow cover depth, and is believed to improve as the depth increases. Chapter 7

The eﬀect of seam dip on stress distribution in an LTCC panel

7.1 Introduction

Another geotechnical issue associated with the extraction of inclined thick seams, retreating along strike, is the designing of the locations of the gate roadways for the next panel, downdip of those already mined. Currently, almost all of the longwall panels in the Quangninh coalfield were developed using a single entry system. A longwall panel is set up by two gateroads (a tailgate and a maingate); the maingate will be protected so it can be a usable tailgate for the next longwall panel during the progression of mining operations, while the tailgate is left unsupported. In order to protect the maingate, a submaingate is driven in parallel to the maingate, as illustrated in Figure 7.1. As a result, the designing of the chain pillar width between the sub-maingate and the maingate is a crucial part of successful operations of thick seam mining in the Quangninh coalfield. However, the challenge is that designing a chain pillar for inclined thick seams in the Quangninh coalfield is developed through trial and error. Many mines trialled a pillar size that had been used successfully in flat seams. Once a pillar size proves to be successful, it is often used over and over again even though its size is wasteful in many applications. Therefore, an improved understanding of

163 164 stress distribution in the chain pillar for inclined seams is necessary.

Figure 7.1: Typical longwall layout in the Quangninh coalﬁeld

The study outlined in this chapter aims to investigate the effect of seam dip on the stress distribution in the chain pillar for a face retreating along the strike. In addition, the study also aims at extending the understanding of the caving mechanism in the top coal section when using the LTCC method to extract inclined coal seams. In order to achieve these objectives, a brief literature review on longwall pillar design is first undertaken. The aim of the review is to identify the effect of the chain pillar on the stability of the gate roadways, especially the effect of abutment stress applied on the pillar. Then, the effect of the seam dip in the abutment stress location is discussed by reference to the results of numerical analysis. Such knowledge can assist mine planners to design the location of gate roadways for the next longwall panel, downdip of those already mined during the extraction of inclined thick seams, in order to avoid or minimize the need for subsequent remedial work on the roadways, and improve coal recovery.

7.2 Literature review

The design of the chain pillar between longwall panels is a major consideration in the economics of mining operations. The main function of the pillar left between 165 the edge of the longwall panel and the gate roadways is to maintain the stability of the gateroads during the lifetime of the longwall operations. In addition, chain pillars are also used for isolating the current working of the planned panel in order to protect against possible mining hazards (Whittaker and Singh, 1979a). The critical importance of chain pillar design is not only to stabilise the longwall entries during the service life for the longwall panel, but also to minimize the size of the pillar to improve coal recovery, as the longwall pillars are seldom recovered (Mark, 1987). The designing of the chain pillar between the edge of a longwall panel and the maingate, where the longwall panel developed by a single entry system, plays an important part in the success of longwall operations. A gateroad usually lasts from performing as the maingate for one longwall panel, to performing as the tailgate for the next. Hence, it must remain serviceable for both machinery and personnel access, as well as for the ventilation of the longwall face. A maingate in a single entry system is subjected to a range of loading conditions during the progression of mining operations. They are:

• The initial state where the gate road is driven in the original stress ﬁeld,

• The abutment stress as the result of the longwall extraction, and

• The gateroad is subjected to a modiﬁed stress environment because of the previous extraction.

There are two main approaches to coal pillar design (Mark, 1987; Majdi, Hassani and Cain, 1991; Tarrant, 2006):

• The ultimate strength pillar approach which assumes that the load-bearing of a pillar will be reduced to zero when its strength is exceeded. The term “failure” is used to describe a pillar whose strength had been exceeded. In this case, designing pillar width is dependant on the loading applied to the pillar and the strength of the pillar, and this load is assumed to be constant across the entire pillar. 166

• The progressive yielding approach contends that yielding is initiated at the most critical point and propagates gradually to ultimate yielding. In this approach, the stability analysis is based on pillar strength and the load at a given point, and the pillar strength and stress concentration across a pillar are assumed to be variable.

Several approaches have been developed to estimate the pillar width of the coal seam for both the single-entry system (Wilson, 1972; Wilson, 1982), and multi-entry system (Carr and Wilson, 1982; Hsiung and Peng, 1985; Mark, 1987). Although both the strength pillar and yield pillar design approaches are undertaken based on diﬀerent assumptions, their advocates agree that stress distribution in the chain pillar plays an important part in pillar design. A number of past cases have proved that the roadways placed at a high stress concentration were subjected to poorer ground conditions compared to other regions (Whittaker and Singh, 1979b; Koehler, Demarco and Wuest, 1996). Therefore, understanding the intensity of pillar stress and the location of associated strata pressure abutments is a crucial part in pillar design.

7.2.1 Distribution of the stress in the chain pillar

Studies by a number of researchers (Wilson, 1972; Whittaker, 1974; Whittaker and Singh, 1979b; Wilson, 1982) have pointed out that longwall mining operation produces a change in the stress distribution of strata around a longwall panel. As mining progress, the development of the load applied to the coal pillar, named the abutment load, is a complex phenomenon aﬀected by various geometric and geologic parameters, such as the cover depth of the coal seam, panel width, extraction height, in-situ stress and the characteristics of the coal and rockmass (Mark, 1987). Figure 7.2 presents the re-distribution of stress around a single longwall face for conditions in the United Kingdom developed by Whittaker (1974), which has been widely accepted for many years. As can be seen in Figure 7.2, the main stress abutments are indicated by:

• The front abutment stress zones which are located a certain distance in the 167

chain pillar from the faceline, depending on the characteristics of the coal and roof rockmass, and it travels with the extraction of the longwall face. Front abutment stress is a major factor aﬀecting the stability of the coal front along the longwall face (Whittaker, 1982). A discussion of the eﬀect of the abutment stress on the cavability of top coal has been discussed in section 5.2.3.

Figure 7.2: Redistribution of vertical stress around a single longwall [(after Brady and Brown, 2004), adapted from Whittaker (1974)]

• The side abutment pressure zone, which is located in the chain pillar next to the face ends of the longwall panel. Side abutment stress is a major feature aﬀecting the stability of the gateroads in the next longwall panel.

Several research studies have been undertaken to estimate the magnitude and location of stress in the ribside pillar of a longwall panel. One of these was an approach undertaken by Wilson (1972), who studied geo-mining conditions in the United Kingdom. Wilson states that the constancy of cover load is one of the fundamental conditions of equilibrium. When mining below ground, the cover is not removed, hence, the decrease in the load on the longwall panel must be taken up by a rapidly increasing load on the pillar nearby, which rises to an abutment peak in a comparatively short distance, then decreases slowly to 168 cover load. As mentioned in Section 5.2.3, Wilson (1972; 1982), based on the concept of a progressive failure approach, argues that the re-distribution of stress in the ribside located next to the longwall panel will form three main vertical stress distributions: yield zone, the abutment stress zone, and the zone where the strata pressure slowly decreases to cover load. The yield zone area is where the strata have passed their failure limit. Wilson states that any “load deﬁciency” in the goaf area must be taken up by the pillar. By the investigation of the convergence of roadways caused by the longwall extraction and the compaction of broken rock, Wilson concludes that the load carried by the waste increases from zero at the extraction edge to the original cover load at a distance of 0.3H. If the waste is less than 0.6H across, the stress in the centre will not reach cover load but a linear stress rise will still occur, as illustrated in Figure 7.3. Figure 7.4 presents the area involved in load balance in the ribside pillar.

Figure 7.3: Distribution of the stress in the goaf area (after Wilson, 1982)

Figure 7.4: Load balance in the ribside pillar (after Wilson, 1982) 169

From the above mentioned concept, Wilson (1982) proposed a formulation to calculate a minimum pillar width, Wrp, for various roof conditions, as in equation 7.1 for the protection of a gateroad.

Wrp = 2 · (xb + d) (7.1)

where:

xb = The yield zone, determined by equation 5.3 d = an exponential decay factor, determined by equation 7.2 A + q · x − A d = W b b (7.2) Amax − q

AW = the load of abutment stress on the chain pillar For panel widths greater than 0.6H:

2 AW = 0.15γ · H (7.3)

For panel widths less than 0.6H:

W W A = γ(H − ) (7.4) W 2 1.2

W = the panel width q = cover load

Ab = the load on yielding zone H = the depth of cover

Amax = the magnitude of the abutment stress.

Amax = k · q + σ0 (7.5)

k = Triaxial stress factor

σ0 = in-situ uniaxial compressive strength of the coal seam.

The basis for this pillar width was that a roadway should be placed at a shadow area of the vertical abutment stress (where the vertical stress decreases toward the cover load), outside the peak section of the abutment curve. The greater the vertical stress increases, the higher the roadway convergence (Wilson, 1982). 170

Another study on stress behaviour was undertaken by Whittaker and Singh (1979b). This empirical study examined the deformation of 82 gate roadways in the Midland coalﬁelds, England. Results from this study concluded that the gate roadways, located in the range 10 - 30m from the edge of the longwall panel, were subjected to a higher gate roadway closure than other locations in the ribside pillar (pillars which have a greater width than 10 - 30m or pillars having smaller width, which ranges from 0 - 5m). Figure 7.5 plots a graph showing the relationship between the gate closure and the pillar width.

Figure 7.5: Relationship between rib pillar width and the vertical closure in gate roadways (after Whittaker and Singh, 1979b)

The ﬁndings from another empirical study undertaken by Koehler et al. (1996) in Sunnyside, Utah, USA also agrees with the trend in stress behaviour in the chain pillar identiﬁed by Wilson (1972; 1982), and Whittaker and Singh (1979b). From the results obtained in that study, Koehler et al. (1996) concluded that the roadway placed at a location of high stress concentration can promote poor ground conditions. Figure 7.6 presents the relationship between the pillar width and the roadway stability in the longwall extraction. 171

Figure 7.6: Relation between the pillar width and gateroad performance (after Koehler et al., 1996)

The results obtained from field investigations undertaken by Whittaker and Singh (1979b) and Koehler et al. (1996) are similar to the analytical approach proposed by Wilson (1972; 1982). The stress distribution in the ribside pillar can be divided into three zone areas. The first zone area in the coal pillar is where coal has been deformed. This is the “yield zone” in Wilson’s method, and is not expected to carry the abutment load. The second zone is the high intensity zone of the stress abutment. The location of maximum abutment vertical stress depends on the physical properties of the roof and floor, seam thickness and cover depth (Peng and Chiang, 1984; Koehler et al., 1996). The third zone is the area where the abutment stress gradually decreases to the original vertical stress. Among the three above zones, the gate roadways located in the low stress area is thought to have a better ground conditions than in the peak abutment stress by reducing the stress concentration near the gate entries. For example, the use of gate entries located in the “yield zone”, known as skin-to-skin roadways, has been practised in the United Kingdom for over 100 years (Yavuz and Fowell, 2004). According to Yavuz and Fowell (2004), although the no pillar system (the pillar width is from 0 - 5m) was economical in terms of not abandoning reserves of coal in place and resulted in reducing the risk of spontaneous combustion, roadways would not be considered safe when both the support system and time are 172 regarded as critical factors for stability. As a result, in reality, the mine planners tend to design a long continuous chain pillar between longwall panels to protect the roadways. The determination of the location of high stress concentration is therefore very important in designing the pillar width between the panel edge and the gate roadways.

7.2.2 Discussion

Previous work in relation to chain pillar design is relevant but mostly for traditional longwalling in horizontal coal seams. Both empirical and theoretical research indicate that the extraction of a longwall can produce re-distribution and create abutment stresses on the coal pillar both in front and beside the longwall face. These abutment stresses are the direct result of stress transfer from the adjacent extracted longwall panel (Whittaker and Singh, 1979b). Roadways located within this region would suﬀer more damage compared to the other regions. When considering the design of a gate roadway it is usual to pay attention to the location of peak abutment stress, so that the gate roadway can be located in the low vertical stress area. However, the eﬀect of seam dip on the stress distribution in the chain pillar has not been thoroughly investigated. As a result, a better understanding of the geomechanical behaviour in the chain pillar between the face end of the longwall panel and the maingate for the extraction of inclined thick seams is necessary to provide assistance for support design.

7.3 The eﬀect of seam dip on stress distribution in the chain pillar

The following section aims to investigate the diﬀerences in stress distribution in the chain pillar of the face ends in the LTCC method between diﬀerent seam dips by reference to the results of numerical analysis. Numerical modelling was carried out using the FLAC software based on continuum methods. 173

7.3.1 Mining geometry

A total of 4 geometric models were simulated to represent diﬀerent seam dips of 00, 150, 300, and 450. The models consisted of a coal seam and a number of rockmass layers. The coal seam has a thickness of 6m and was divided into two parts: the cut and the top coal sections. The heights of the cut and the top coal sections were both assumed to be 3m. The longwall layout simulated by the FLAC programme was a commonly used single-entry system. The simulation of the models representing the mined- out panel width in the LTCC face operated along the strike at the location where the longwall face has been far away and the stress re-distribution was assumed to have achieved a stable condition. The diagram showing the location of the FLAC analysis is presented in Figure 7.7.

Figure 7.7: Diagram shows the location of FLAC analysis

7.3.2 Material properties and constitutive models

Four constitutive models (materials) were employed in the numerical simulations. The strain-softening material was used to represent the coal seam with strength properties being a function of plastic strain. The Mohr-Coulomb material was used to represent the rockmass surrounding the coal seam. The far field region 174 away from the coal seam was represented by an elastic material. The double yield material was used to represent the broken material in the goaf area in order to reflect the behavior of waste materials in realistic situations where volume decreases when compacting occurs, depending on the stress level. In order to achieve suitable calibrated modelling results, the laboratory UCS values of coal and rock samples taken from the testing were calibrated to take into account the in-situ rockmass effect. To calibrate the model it was assumed that the initial caving (the distance of longwall face where the broken rockmass can fill the space created by the conventional longwall method) was 15 - 20m. This was based on experience that practiced in many underground mines in the Quangninh coalfield. Other material properties, such as the density and the angle of friction of the coal and surrounding rockmass were assumed according to similar rock types and conditions. In the FLAC analysis, the relation between the unconfined compressive strength, σc, cohesion, c, and friction angle, φ, is given by:

φ σ = 2 · c · tan(45 + ) (7.6) c 2 The input data in relation to the physical and mechanical properties of the coal and rockmass were extracted from geological reports from the Vietnamese industry, and are given in Table 7.1.

Table 7.1: Properties of coal and rockmass used in FLAC models

Name of Density Bulk modulus Shear modulus Cohesion Friction Tensile rockmass (T/m3) (MPa) (MPa) (MPa) angle (0) strength (MPa) coal 1.6 2.0 1.2 11.5; 17; 25 29 1 immediate roof 2.62 4.7 2.8 3.5 32 1.5 main roof 2.6 6.0 3.6 5.4 33 1.5 immediate ﬂoor 2.62 4.7 2.8 3.5 32 1.5 main ﬂoor 2.6 6.0 3.6 5.4 33 1.5

For comparison, material properties and the initial stress applied in the four models were kept unchanged with average cover depth of 600m. In the current analysis, vertical stress was assumed to be determined by equation 5.7. Diﬀerent values of vertical/horizontal stress ratio will be undertaken for sensitivity analysis. 175

7.3.3 Simulation of face extraction and top coal caving

The excavations were simulated by the removal of every 2m panel width in the direction from the tailgate to the maingate at the bottom section of coal (3m height). This represents an increase of the panel width (or the face length) in the LTCC face, followed by the drawing of 3m in the top coal section. The total simulated longwall lengths in the four models were assumed to be 100m. The simulation of the caving process of coal and rock above the longwall face used the Mohr-Coulomb failure criterion, the same as the methodology used in a research study undertaken by Hebblewhite et al. (2002). After the extraction of 2m of the cut section, every zone which represented coal and rock above the longwall face was checked. Once any zone reached the state of yield in tension or/and in shear, it was removed and stresses in the rockmass around the caved area were then re-distributed until the model reached equilibrium. In the FLAC programme, the basic criterion for material failure is the Mohr- Coulomb relation, which is a linear failure surface corresponding to shear failure (Itasca Consulting Group, 2005):

q fs = σ1 − σ3 · Nφ + 2 · c Nφ (7.7)

where: 1 + sinφ N = (7.8) φ 1 − sinφ

σ1 = major principal stress (compressive stress is negative);

σ3 = minor principal stress φ = friction angle; and c = cohesion.

Shear yield is detected if fs <0 and tensile yield is detected if ft >0. Note that the tensile strength cannot exceed the value of σ3 corresponding to the apex limit for the Mohr-Coulomb relation. This maximum value of tensile strength, τmax, is given by: c τ = (7.9) max tanφ 176

To represent the caving progress of the rockmass, the backfill material was created in the goaf area once the condition in equation 7.10 was satisfied. In this FLAC analysis, it was assumed that the recovery rate of the cut section is 95%, and the recovery rate for the caved coal section is 75%. As a result, by taking into account the bulking effect, the volume of caved rockmass and coal which remained in the goaf area can fill the space of extracted coal from both cutting and caving after the face has advanced a certain distance. A FISH function (a programming language embedded within FLAC to enable the performance of user defined tasks) was created to simulate the bulking cycling.

V1 · 95% + V2 · 75% + V3 = (V1 · 5% + V2 · 25% + V3) · K (7.10)

where:

V1 - the yielding zones of the coal cut section

V2 - the yielding zones of the top coal section

V3 - the yielding zones of the caved roof rockmass K - the bulking factor.

Figure 7.8: Relationship between the applied stress and the volumetric strain used in FLAC analysis

In order to represent the broken materials in the goaf area where volume decreases when compacting, the double yield model, available in FLAC, was used 177 for backﬁll material in the models. It is assumed in the FLAC analysis that the bulking factor at the initial state was 1.5 for both coal and rockmass. The bulking factor will reduce according to compressive strength, and the backﬁll material can stand for a high loading force when the bulking factor reduces to 1.35, equivalent to a 10% reduction in volumetric strain. The relationship between the applied stress and the volumetric strain is illustrated in Figure 7.8.

7.3.4 Simulation of the gateroad in the rib pillar

The gate roadway was simulated in the second stage, after the settlement of the caved waste at 100m of panel width. The purpose of the simulation of the gate roadway in the FLAC analysis is to determine the sensitivity of gateroad behaviour to diﬀerent seam dip scenarios. The gateroad was simulated to locate within the coal seam with pillar widths of 5m, 10m, 15m, and 20m for two dipping scenarios: a ﬂat coal seam and an inclined seam (300). The gate road was supported by liner elements available in FLAC. Liner elements, which are two- dimensional elements with three degrees of freedom (x-translation, y-translation and rotation) at each end of node, can simulate the behaviour of ductile or brittle materials such as steel, concrete, and shotcrete (Itasca Consulting Group, 2005). The properties of liner element are presented in Table 7.2.

Table 7.2: Properties of liner element used in the FLAC analysis

Parameters Value Density, kg/m3 3.5 Young’s modulus, GPa 20 Compressive strength, MPa 100 Tensile strength, MPa 10 Thickness of the liner layer, m 0.15 178

7.3.5 Discussion of the numerical modelling results

A typical stress analysis result for a LTCC mining panel in ﬂat coal seams is presented in Figure 7.9. This indicates that due to stress redistribution from a longwall extraction, a high stress concentration is located at the rib coal pillar beside the maingate and tailgate of the longwall panel, and the peak magnitude of stress concentration is located within the coal seam.

Figure 7.9: Vertical stress distribution around a longwall panel in a ﬂat seam (location of the cross section refers to Figure 7.7)

Figure 7.10 shows the vertical abutment stresses obtained from the ﬂat seam model at the mid-height of the top coal from 5m in the goaf area of the longwall panel to a position of 40m in the rib pillar of the coal seam. As can be seen in Figure 7.10, the peak stress steadily increases with the increase of the longwall panel, indicating that the top coal in the rib pillar is subjected to a greater stress concentration. For every ten metre panel width analysed, the location of the peak vertical abutment stress (beyond the yield zone) and the yielding zone in the top coal section gradually move further into the coal pillar of the coal seam. Within the panel width of 100m analysed, a high stress concentration was located at a 179 distance of 8 - 11m depth in the coal pillar, and the top coal yielded within this 8 - 11m zone depended on the panel width of the longwall. A more detailed model may be required to further conﬁrm whether, and to what mining distance, the peak stress should reach a practically constant level.

Figure 7.10: Vertical stresses at the centre of top coal in ﬂat seams at diﬀerent panel widths

Results from the FLAC analysis show that the location of the peak stress concentration has been affected by the dipping factor of the coal seam. This is illustrated in Figures 7.11 and 7.12, where Figure 7.11 represents the status of vertical stress in the chain pillar in the updip side of the longwall panel. Fig- ure 7.12 displays the distribution of the vertical stress in the chain pillar in the downdip side of the longwall panel. As can be seen in Figure 7.11, when the dip of the seams increases, the high stress concentration in the chain pillar in the updip side the longwall panel moves gradually from the centre of the top coal section of the coal seam downward to the floor rockmass area. This result indicates that more potential problems of floor heave the tailgate of the inclined seams may suffer than that of the flat seams. The far more significant effect of the seam dip on vertical stress distribution 180 is the difference in the location of the abutment stress area at the rib pillar updip of the longwall panel, as indicated in Figure 7.12. When the dip of the seam increases, the peak vertical stress in the chain pillar, downdip side of the longwall panel moves upward to the roof rockmass area.

a. α = 00 b. α = 150

c. α = 300 d. α = 450 Figure 7.11: Vertical stress distribution in the chain pillar, updip side of the longwall panel in diﬀerent seam dips (α)

Figure 7.13 presents the vertical stresses at the middle height of the top coal section of the LTCC panel from 4m at the goaf area to 40m depth in the coal pillar downdip of the longwall panel with different seam dips. Results in Figure 7.13 indicate that the rib pillar of an inclined coal seam is subject to less abutment stress than that of a flat one. The effect of seam dip on the stress distribution at the rib pillar of the coal seam also can be seen in the differences in the behaviour of the gateroad between 181

a. α = 00 b. α = 150

c. α = 300 d. α = 450 Figure 7.12: Vertical stress distribution in the chain pillar, downdip side of the longwall panel in diﬀerent seam dips (α)

Figure 7.13: Vertical stress at centre of coal seam, downdip side of the longwall panel in different seam dips 182 two different seam dip scenarios: 00 and 300. For a flat seam, the gateroad with a pillar width of 10m is subjected to a higher loading force than the other alternatives of 5m, 15m, and 20m in pillar width. For example, the maximum axial force of the liner around the gateroad at a pillar width of 10m is 64MN, while for the case of pillar width of 5m, 15m, 20m are approximately 59MN, 56MN, and 49MN, respectively. The same trend is obtained for the inclined seam case. However, the value of maximum axial force in an inclined seam is smaller than that of a flat one with the same pillar width. These results indicate that with the same width of the rib pillar between the longwall panels, gate roadway in an inclined seam is more stable than in a flat one. Figure 7.14 presents the value of maximum axial force in different locations of gate roadway for flat and inclined seams. Vertical contours and the distribution of the axial forces of the supports of the roadway at different pillar widths for flat and inclined seams (300) are presented in Appendix B.

Figure 7.14: Values of maximum axial force for diﬀerent locations of gateroad between ﬂat and inclined seams

The FLAC analysis also clearly showed that the yielding zones above the longwall face were affected by the dipping factor of the coal seam. As can be seen in Figure 7.15, the maximum yielding zone in a flat coal seam is located at the middle of the longwall panel. When the dip of the seam increases, the maximum yielding zone gradually moves to the uphill portion of the longwall panel. Re- 183 sults from the FLAC analysis indicate that with flat coal seams, better caving is anticipated in the middle portion of the longwall face. With inclined seams, better caving is predicted at the uphill portion of the longwall face compared to the downhill portion. It also can be seen from Figure 7.15 that the lower the seam dip, the larger are the yielding zones in the roof above the LTCC longwall face. This indicates better cavability of the top coal and above rockmass for a flat seam than for an inclined one. The results from the FLAC analysis agrees with the results obtained from the 3DEC analysis discussed in Chapter 6.

a. 00 b. 150

c. 300 d. 450 Figure 7.15: Yield zones above the roof of longwall face in diﬀerent seam dips

The strength of the coal seam is clearly a factor affecting the stress distribution in the chain pillar. An attempt to investigate the effect of the coal strength was undertaken by comparing three different values of UCS of 11.5MPa, 17MPa, and 25MPa of the coal material in flat (00) and inclined (300) seams. Results from the FLAC analysis show that the weaker the coal material, the further the yielding 184 zone extends into the coal seam, resulting in a higher distance from the edge of the longwall panel to peak stress distribution in the chain pillar. This conclusion also applies to dipping seams. In all the results, however, the magnitude of the peak stress applied in the chain pillar of an inclined coal seam is smaller than a flat one. Figure 7.16 presents the effect of coal strength on the vertical abutment stresses obtained at mid-height of the top coal from the edge of the longwall panel to a position of 40m in the chain pillar, downdip side of the longwall panel.

Figure 7.16: Coal strength vs depth of peak stress distribution at the chain pillar, downdip side of a longwall panel

The horizontal/vertical stress ratio will certainly affect the distribution of stress in the chain pillar of face ends. As discussed previously, the state of pre- mining stress, both in terms of loading conditions and the stress path, is complex. Mechanical processes such as the generation or slip of fracture and viscoplastic flow through the rock mass, can be expected to produce both complex and heterogeneous states of stress (Brady and Brown, 2004). As a result, the orientation of the principal horizontal stress may be of any angle intersected to the direction of the longwall face. The purpose of the current study is to investigate a mecha- nistic explanation of the effect of horizontal stress on the stress distribution in the chain pillar of face ends. Under the current study, three different stress ratios k = 0.33, k = 1, and k = 2 (where k is the ratio between the horizontal and the vertical stresses) were simulated for two seam dip scenarios: flat and inclined (300) 185 seams. For all the models, the horizontal stress refers to the principal horizontal stress and was assumed to be perpendicular to the vertical stress. In addition, all the stress paths simulated in the models were assumed to be perpendicular to the face direction. It was revealed from the FLAC analysis that the ratio of horizontal/vertical stress had an effect on the stress distribution in the chain pillar of the coal seams. In the case of flat coal seams, little effect of the ratio between the horizontal and vertical stress was seen on the location of peak stress distribution in the chain pillar, as illustrated in Figure 7.17.

a. k = 0.33 b. k = 1

c. k =2 Figure 7.17: Vertical stress distribution in the chain pillar beside the maingate in a ﬂat coal seam with diﬀerent vertical/horizontal stress ratios (k)

Figure 7.18 shows the vertical stress at the centre of the middle height at the top of ﬂat coal seams from the edge of longwall panel to a depth of 40m 186 in the chain pillar with three diﬀerent values of horizontal/vertical stress ratios. Results in Figure 7.18 indicate that the chain pillar is subject to less vertical abutment stress as the ratio of horizontal to vertical stress increases. Bearing in mind that horizontal stress is a major cause of the separation of bedding layers, and perhaps resulting in instability of the gate roadways, this analysis and interpretation clearly requires further investigation.

Figure 7.18: Vertical stress at the centre of a ﬂat coal seam in diﬀerent vertical/horizontal stress ratios (k)

Results obtained from the models representing an inclined (300) coal seam show a large difference in the location of the peak stress area when changing the values of the horizontal/vertical stress ratio. The location of the abutment stress area gradually moves far away from the coal seam as the ratio of horizontal and vertical stress decreases. In contrast, when the horizontal/vertical stress ratio increases, the peak stress area gradually moves towards the coal seam, as illustrated in Figure 7.19. Figure 7.20 presents the vertical stress at the middle height of the coal seam in the chain pillar, downdip side of the longwall panel. Results indicate that at under the same geological conditions, the chain pillar with higher horizontal/vertical stress ratio is subjected to a higher abutment stress than it is under lower the horizontal/vertical stress ratio. The trend obtained from the inclined coal seam model is different from the results observed in the flat seam models, where the chain pillar under a higher horizontal/vertical stress ratio is subjected to less abutment stress than it is under a lower horizontal/vertical stress ratio. 187

a. k = 0.33 b. k = 1

c. k =2 Figure 7.19: Vertical stress distribution in the chain pillar, downdip of a longwall panel for inclined seam under diﬀerent vertical/horizontal stress ratios (k)

Figure 7.20: Vertical stress in the chain pillar, downdip side of an inclined (300) longwall panel under diﬀerent vertical/horizontal stress ratios (k) 188

7.4 Conclusions

This chapter discussed the effect of seam dip on stress distribution in the chain pillar, downdip of the longwall panel, when extracting inclined thick coal seams using the LTCC method, and the cavability of the top coal and above rockmass in the LTCC method. The literature review shows that a pillar left between the longwall panel and the gate entries has been proposed as a means to improve the stability of the roadway during longwall mining as well as to prevent potential mining hazards. Designing the proper size of the coal pillars so that the roadways can be maintained in a stable condition and minimizing the coal loss from leaving the pillar unmined, are always crucial parts of a longwall operation. There are two main approaches to coal pillar design. The first is the ultimate strength pillar approach which designs the pillar width based on both the loading applied to the pillar and the pillar supporting full abutment loads. The second is the progressive yielding approach which designs the pillar width based on the stability analysis at a given point. Although both the strength pillar and the yield pillar design approaches are undertaken based on different assumptions, their advocates agree that stress distribution in the chain pillar plays an important part in pillar design. The abutment stress in the chain pillar is considered a major factor influencing the stability of the roadway. The magnitude and the location of the abutment stress are affected by many factors such as cover depth, and the properties of the roof and floor of the coal seams. It has been proposed that gate roadways be placed in the the shadow area of the high stress distribution in order to obtain better ground conditions. It is clear from the FLAC analysis that the dip of the seam affects the stress distribution in the chain pillar of the face ends in the LTCC method. In a flat coal seam, the peak stress area is located within the coal seam with different values of horizontal stress, while in an inclined coal seam, the location of the abutment stress area is largely dependant on the ratio between the horizontal 189 and vertical stress. If the horizontal stress is smaller than the vertical stress, the abutment stress area is located in the roof rockmass area above the coal seam. As the horizontal stress increases, the abutment stress moves downwards to the coal seam area. However, all results indicate that the maximum vertical abutment stress of an inclined seam is smaller than that of a flat coal seam. This means that the same pillar width for the gate roadway of an inclined seam is subjects to less vertical abutment stress than it would for a flat coal seam. Results from the FLAC analysis in this chapter also confirm that the seam dip has a large effect on the cavability of the top coal and above rockmass when extracting a thick seam using the LTCC method. Results from the FLAC analysis also show that for flat coal seams, better caving is anticipated in the middle portion of the longwall face, while with inclined seams, better caving is predicted in the uphill portion of the longwall face compared to the downhill portion. In addition, better cavability of the top coal and above rockmass is anticipated in flat coal seams than in inclined ones. Chapter 8

Development of a cavability assessment criterion for the LTCC method in inclined thick seams

8.1 Introduction

The purpose of the parametric studies described in Chapters 5 and 6 was to lay the foundation for an improved assessment system to predict the cavability of top coal when using the LTCC method. This chapter describes this system in detail. This chapter begins with a literature review of existing cavability assessment techniques for the LTCC method. An improved assessment system for predicting cavability of the top coal when extracting inclined thick seams by the LTCC method is then presented. In the ﬁnal section of the chapter, back analysis is undertaken for the realistic application of the suggested cavability assessment system.

190 191

8.2 Literature review

The response of rockmass surrounding a mining extraction is commonly predicted from geotechnical data collected in the exploration, or feasibility study stage by using analytical tools to assist the selection of an appropriate mining method. The two most widely known geomechanical classifications used in tunnelling design and some other applications are the ’Q’ index (Barton et al., 1974) and rock mass rating (RMR) system (Bieniawski, 1976). However, according to Mark and Molinda (2005), the standard rockmass classification systems are not feasible for use in coal mining because of two reasons. Firstly, these classification systems tend to focus on joint properties, while bedding is generally the most significant discontinuity in the coal mine roof. Secondly, the standard rockmass classifications only rate one rock unit at a time, when the coal mine roof often consists of several different layers of coal or rockmasses. For underground coal mining, some classification systems have been developed for predicting the stability of the roof and pillars. They are the Coal Mine Roof Rating (CMRR) (Mark and Molinda, 2005), and the Coal Measure Classification (CMC) (Whittles et al., 2006). In addition, a number of methods have been developed for assessing the cavability of the top coal for the LTCC method. These are empirical approaches developed by Zhang et al. (1995), Jia (2001), and the Shandong Coal Mining Administration in China [cited by Xu (2004), no information about original reference]. Recently, an analytical approach (using numerical methods combined with back analysis) was also developed by Humphries et al. (2006) and Vakili (2009) for predicting the cavability of top coal in Australian geo-mining conditions. These methods have made important contributions to the knowledge, but all of them are limited in predicting the behaviour of top coal in horizontal coal seams. In the following sections, some of the most published assessment cavability methods for the LTCC method will be critically reviewed. The purpose of these reviews is to evaluate the sensitivity of the input parameters on the cavability of the top coal, and to discuss the advantages and the limitations of each method. 192

As a result, the attributes required for an improved method will be identiﬁed.

8.2.1 Method of vertical displacement in top coal

Zhang et al. (1995) [cited by Xu (2004) and Vakili (2009)] developed a cavability assessment method which was based on mine investigations and laboratory tests. In this method, it was suggested that the accumulative vertical displacement of the top coal at a certain location from the faceline could be used as a criterion for predicting the cavability of the top coal. The accumulative vertical displacement of the top coal is given by equation 8.1:

Lm · (S + 10) = C (8.1)

Where: L is the accumulative displacement of top coal; S is the distance away from coal the faceline (positive values indicate investigated position located ahead of the faceline, and negative values refer to the position located behind the faceline; m is the power of displacement; and C is a constant. In order to predict the cavability of top coal, a set of observation holes would be drilled to desired height levels (mainly between 5m and 6m from the seam floor) of the top coal for collecting vertical displacements of certain positions in the top coal. The displacements then can be used as criteria for assessing the cavability of the seam. It is suggested, in the vertical displacement method, that the coal seam has excellent cavability when the accumulative displacement is greater than 100mm at the investigated distances of greater than 2m ahead of the faceline (S > +2m). The top coal or rockmass in this category may cause stability problems, such as roof fall or face spalling during mining operations. On the other hand, if the displacement is only 100mm when 2m away behind the faceline (S < -2m), the seam is considered to be difficult for top coal caving. Figure 8.1 plots the relation between the cavability of the top coal and the locations in top coal where the accumulative vertical displacement reaches 100mm at height levels (H) from the seam floor of 5m and 6m with different UCS values. 193

Figure 8.1: The face distance of positions where the vertical displacement reaches 100mm with diﬀerent UCS values (after Zhang et al., 1995)(cited by Xu, 2004; Vakili, 2009)

This cavability assessment method appears to have two main disadvantages. One limitation is the need to obtain the vertical displacement at the site in order to predict the cavability of the top coal. Consequently, this method is time consuming, costly, and can not be used to predict the cavability of the top coal at the feasibility stage when the monitoring results from the mine site are not available. In addition, this cavability assessment method indicates that the UCS of the coal seam is the only factor that aﬀects the cavability of the top coal. As commented by Vakili (2009) and discussed in Chapter 6, the cavability of the top coal is aﬀected by a number of factors. As a result, the UCS value alone can not present a measurement for a good decision making in feasibility studies.

8.2.2 The fuzzy cluster evaluation method

As cited by Xu (2004), Jia (2001) used the fuzzy cluster evaluation method to classify the cavability of top coal by comparing 7 different parameters of coal and roof rockmass to the caving performance of the LTCC faces witnessed in some coal seams in China. The seven main parameters used in Jia’s classification are: coal strength, the cover depth of the coal seam, bedding/cleat spacing, the ratio of cut and top coal thickness, band thickness in the top coal, characteristics 194 of the immediate, and characteristics of main roof. By comparing the caving performance witnessed at the coal seams extracted by the LTCC method with the seven different geological parameters above, the cavability of the top coal was classified into 4 categories: 1 - excellent cavability; 2 - good cavability; 3 - medium cavability; 4 - poor cavability, as presented in Table 8.1. As stated by Vakili (2009), the caving performance of some LTCC practices in China, shown in Table 8.1, provides a useful source of information on the relationship between the geological parameters and the caving mechanism of the top coal, which can be used for future study. However, since there are no guidelines for classifying the cavability of the top coal, the classification method developed by Jia (2001) could not be used to predict the cavability of the top coal at the planning stage when geological parameters obtained from exploration are available. Vakili (2009) evaluated the importance of these parameters in the cavability method developed by Jia (2001) and identified that the coal strength and the spacing of discontinuities are the most important factors influencing the cavability of top coal, while three other parameters (immediate roof, main roof and the cover depth) have a minor impact on the top coal’s cavability. This can be observed in Table 8.2. 195 y 1 2 2 1 3 2 2 4 1 2 2 1 2 2 Category Cavabilit cut 1.4 1.4 2.8 5.22 1.54 2.36 1.14 1.45 1.571 1.679 1.462 1.589 2.674 1.571 Ratio Cave/ 2 2 2 1 3 2 2 4 1 2 2 2 2 2 Category Main Roof of 0 1.6 0.5 1.3 1.46 1.543 1.167 0.325 1.143 0.312 1.134 0.406 1.159 0.525 Immediate Roof Filling Factor 0 0 0.2 0.4 0.2 0.2 0.3 0.4 0.3 (m) 0.12 0.15 0.25 0.31 0.21 Band Thickness 1 3 2 1 4 2 2 5 1 2 3 2 3 2 Category Joint/Cleat Depth - 150 480 250 130 210 240 254 350 230 360 412 210 150 (m) Cover 20 10 11 7 20 30 38 30 25 32 20 24 15 30 5 - 5 - (MPa) 15 - UCS of coal Cavability classification of some Chinese LTCC faces developed by Jia (2001) (after Xu, 2004) Mine, Xingtai zhuang Mine in Yanzhou 2 Mine, Datong site Mine, Shengyang 6 Mine, Hebi Table 8.1: liang Mine, Xuangang ghejing Mine, Xuzhou Shuiyu Mine, Fengxi hung, Zhendzhou 1 Mine, Yanquan No 1 Mine, Pinding Shan 4 Mine, Yanquan angzhuang Mine, Luan No enghuangshang Mine, Jincheng No Mic 11, Puhe No F No W liujia Xingtai Xinlong Shan Name of mine Seam 3, Seam 8, Seam 15, Seam 15, Seam 3, Seam 4, Seam 3, Seam 11,12, Seam 5, Seam 2, Seam 3, Seam 10, Da Seam, Seam 7, 196

Table 8.2: The eﬀect of individual parameters on cavability of the top coal in Jia’s classiﬁcation method (Vakili, 2009)

Parameter Weighting (%) UCS 38.20 Joint spacing 36.30 Cave/Cut ratio 7.20 Band thickness 7.10 Immediate roof 4.10 Cover depth 4.00 Main roof 3.10

8.2.3 Empirical cavability assessment method

The Shandong Coal Mining Administration in China [cited by Xu (2004), no information about original reference] also developed an empirical method to predict the cavability of top coal for the LTCC method. This cavability assessment system was created by rating 7 geological parameters which have close relations with the caving performance of the top coal. They are: coal strength, cover depth, characteristics of the immediate roof, characteristics of the main roof, the ratio of cut and top coal thickness, discontinuities of spacing, and the characteristics of the band in the seam. The effect of individual parameters on the cavability of the top coal is rated in accordance with certain geo-mining conditions, as presented in Tables 8.3 to 8.9. The cavability of the top coal is calculated by the summation of those 7 different affecting factors given in equation 8.2.

7 X M = Ai · µi (8.2) i=1 where:

Ai = the weighting of the aﬀecting factor number i, given in Table 8.10

µi = the aﬀecting factor number i. 197

Table 8.3: Ratio of cover depth to coal UCS versus µ1 (after Xu, 2004)

H/Rc <5.5 5.6 - 10 10.1 - 15 15.1 - 20 20.1 - 30 30.1 - 40 >40

µ1 0.2 0.3 0.7 0.8 0.85 0.9 0.95

Note: H = cover depth of the seam; Rc = coal uniaxial compressive strength.

Table 8.4: Category of immediate roof versus µ2 (after Xu, 2004)

Immediate roof Unstable Medium stable Stable Strong

D = Rc.C1.C2 ≤30 31 - 70 71 - 100 >100 First weighting distance ≤8 9 - 18 19 - 25 >25

µ2 0.9 0.8 0.7 0.4

Note: D = strength index; C1 = Joint spacing index; C2 = bedding thickness index.

Table 8.5: Category of main roof versus µ3 (after Xu, 2004)

Main Roof Weak Medium Strong Very Strong

µ3 0.4 0.6 0.7 0.8

Table 8.6: Ratio of cut to cave versus µ4 (after Xu, 2004)

Rc (MPa) ≥25 <25 Cut/cave ratio 1:0.5 1:(0.5-1) 1:(1-1.5) 1:(1.5-2) 1:(2-4) 1:0.5 1:(0.5-1) 1:(1-1.5) 1:(1.5-2) 1:(2-4)

µ4 0.5 0.7 0.9 0.85 0.6 0.4 0.5 0.7 0.89 0.8

Table 8.7: Cleat spacing versus µ5 (after Xu, 2004)

Rc ≤10 11 - 15 16 - 20 20 - 30 >30 Cleat space <0.18 0.19 - 0.3 0.31 - 0.4 0.41 - 0.47 >0.47

µ5 0.9 0.85 0.8 0.5 0.3

Table 8.8: Band thickness versus µ6 (after Xu, 2004)

Band thickness, mm <100 100 - 200 200 - 300 >300

µ6 1 0.8 0.5 0.1 198

Table 8.9: The UCS of band layer versus µ7 (after Xu, 2004)

UCS of band layer <10 10 - 20 21 - 30 >30

µ7 1 0.8 0.4 0.2

Table 8.10: Weightings of diﬀerent factors (after Xu, 2004)

Factor H/Rc Immediate Main Cut/cave Cleat Band The UCS roof roof ratio space thickness of band

Weighting Ai 0.23 0.12 0.1 0.14 0.14 0.12 0.15

The calculated result obtained in equation 8.2 is then compared to the cavability classiﬁcation Table 8.11.

Table 8.11: Cavability classiﬁcation developed by Shandong Coal Mining Admin- istration (after Xu, 2004)

M 0.9-1 0.8 - 0.9 0.65 - 0.8 0.5 - 0.65 <0.5 Cavability Very good Good Medium Poor Very poor

This cavability assessment method is one of the most comprehensive systems as it covers the effects of many parameters influencing the cavability of the top coal. The main advantage of the Sangdong Coal Mining administration’s cavability method, compared to the classification developed by Jia (2001) is that since the requirement of the data collection is clear, this cavability criterion is easy to apply. As seen in Table 8.10, among the investigated parameters in this method, coal strength and cover depth are twice as important as other parameters, accounting for approximately 10 - 15% of the overall important factors of the cavability.

8.2.4 CSIRO’s cavability criterion

Humphries et al. (2006) developed an assessment method using numerical analysis to predict the cavability of top coal for the LTTC method for Australian geo- 199 mining conditions. The study began by evaluating the parameters influencing the cavability of top coal, including coal strength, cover depth, top coal thickness, immediate and main roof conditions, and support setting pressure. Results from the study concluded that the support capacity, the immediate roof strength, and the main roof strength have a low significant effect on the cavability of top coal. The study then created a new cavability criterion based on the three most important factors: coal strength, cover depth, and top coal thickness, by using a multiple regression, given in equation 8.3:

CI = −2.64 + 0.0395H − 0.72σc + 0.191T (8.3)

where:

CI = the caving index, H = the cover depth of the seam,

σc = the UCS of the coal seam, and T = The top coal thickness.

The caving index (CI), calculated from equation 8.3, was then back analysed with the data obtained from Chinese practices. The categorisation of the performance of the LTCC method based on the caving index is expressed in Table 8.12.

Table 8.12: CSIRO’s cavability classiﬁcation (Humphries et al., 2006)

Success Excellent Good Moderate Poor Very Poor Top Coal Recovery (%) >80 65 - 79 50 - 64 30 - 49 <30 Caving Index (CI) >0.7 0.7 to -4.8 -4.8 to -10.3 -10.3 to -17.64 <-17.64

However, like the same method of vertical displacement in top coal developed by Zhang et al. (1995), the cavability method developed by Humphries et al. (2006) seems to be too simple, as it assumes that only three parameters (coal strength, cover depth, and top coal thickness) affect the stability of the top coal. However, the effect of discontinuities, which are one of the important factors 200 influencing the cavability of top coal, was not investigated. Experience of the LTCC method has proved that their three chosen parameters are not good enough to predict the cavability of the top coal.

8.2.5 Numerical cavability assessment method

Most recently, Vakili (2009) conducted a numerical analysis to develop a cavability criterion. The analysis, undertaken by a discrete element numerical code, evaluated the eﬀect of 5 diﬀerent parameters on the cavability of top coal. These parameters were: coal strength, top coal thickness, bedding spacing, cleat spacing, and the cover depth of the seam. The cavability criterion is based on two caving indexes, MCD and TCR which are given in equations 8.4 and 8.5.

−0.0136σh MCD = 0.397·UCS +0.46T +12.56·Jh +19.56·Jv +22.417e −19.5 (8.4)

−0.1514 2 TCR = −1.432·UCS+61.322·T −82.447·Jv+0.015·σv +0.0000401·σv+75.42 (8.5) Where:

MCD = Main caving distance, m TCR = Top coal recovery, % UCS = the uniaxial compressive strength of the coal seam, MPa T = The top coal thickness, m

Jv = spacing of vertical joints, m

Jh = spacing of horizontal joints, m

σh = major horizontal principle stress magnitude, MPa

σv = vertical principle stress magnitude, MPa.

The cavability assessment method developed by Vakili (2009) uses the same methodology of Humphries et al. (2006). The main diﬀerence is that in Vakili’s cavability assessment criterion, by using the discrete element modeling software 201

(UDEC and 3DEC), properties of in-situ coal or rockmass have been investigated in detail, as the nature of rockmass consists of properties of both intact rock and discontinuities. In his study, Vakili (2009) assumed that the failure mechanism of the top coal and the rockmass above was controlled by discontinuities. This assumption seems to be reasonable in the stratiﬁed strata where high densities of discontinuities are involved.

Table 8.13: Cavability classiﬁcation developed by Vakili (2009)

Category I II III IV V Cavability Weak roof Good Fair Poor Very poor MCD <7 7 - 15 15 - 23 23 - 32 >32 TCR >100 79 - 100 59 - 79 39 - 59 <39

8.3 Discussion

The assessment methods for predicting the cavability of top coal for the LTCC method, as described earlier, have made a comprehensive contribution to the understanding of the effect of various parameters on the cavability of top coal. Typically, a cavability criterion quantifies four to five parameters of coal and/or rockmass which are considered to have the most significant effect on mining operations. The cavability criteria for predicting the stability of top coal for the LTCC method, developed by both empirical and computational methods, indicated that the most significant parameters influencing the cavability of top coal are: coal strength, discontinuity spacing, cover depth, top coal thickness, and the characteristics of the immediate and main roof. However, these criteria are mainly used for predicting the cavability of top coal for horizontal coal seams. For the extraction of the inclined seams, the effect of seam dip should be considered. It is therefore desirable to have a cavability criterion developed specifically for the prediction of the stability of top coal for the extraction of inclined thick seams using the LTCC method. Because of these discussed reasons, a cavability criterion developed for predicting the cavability of top coal in steeply dipping 202 conditions is described in the following section.

8.4 Cavability assessment criterion for inclined thick seams

In this section, a cavability assessment method called “caving performance index” (CPI) is developed to predict the cavability of the top coal when extracting inclined thick seams by the LTCC method. The cavability assessment method was developed by evaluating the results of the computation analysis and back analysis from the databases obtained through practice. The methodology of this study is similar to the study undertaken by Humphries et al. (2006), and Vakili (2009). The main difference is that in the CPI method, the effect of the seam dip is investigated so that it can be used to predict the cavability of thick seams in steeply dipping conditions. The first stage in the development of the CPI system is to identify the most significant parameters affecting the cavability of the top coal which will be used in this cavability assessment method. In Chapter 5, a critical review of the impact of the parameters on the cavability of top coal revealed that the most significant parameters are: coal strength, cover depth, top coal thickness, the spacing of bedding and joints/cleats, characteristics of the immediate and main roof, the characteristics of band layer in the top coal, and the moisture sensitivity. Ac- cording to Vakili (2009), when choosing the parameters affecting the cavability of top coal, it is necessary to consider not only their importance, but also the simplicity of data collection during field investigations. By evaluating the importance of various parameters with reference to existing cavability assessment systems (Jia, 2001; Humphries et al., 2006; Vakili, 2009), the following six independent parameters were identified as being most influential on the cavability of top coal in inclined seams under this study:

• Coal strength

• Cover depth of the coal seam 203

• The dip of the coal seam

• Spacing of the bedding planes

• Spacing of the joints/cleats, and

• Top coal thickness.

It is recognised that several other parameters might have an effect on the cavability of the top coal. Taking into account the number of parameters that were a part of this investigation, simplification was required in the form of best estimates to reduce the number of simulations. The next step in the development of the cavability criterion was to quantify the effect of individual parameter on the cavability of top coal. In Chapter 6, the impact of each parameter on the cavability of the top coal was evaluated by using computational analysis (3DEC software). In the 3DEC analyses, the caving performance index (CPI), was used as a tool for evaluating the effect of an individual parameter on the caving performance of the top coal. Based on the results from that numerical analysis, regression analysis was undertaken to examine the relationship between the CPI and specified independent variables. The correlation statistics from multiple regression were summarised in Table 8.14. Details of the mathematical model of their relationship in the regression equation are described in Appendix C.2.

Table 8.14: Regression Statistics

Parameter Value Multiple R 0.919699 R Square 0.845847 Adjusted R Square 0.79144 Standard Error 5.43421 Observations 24

The multiple linear regression of the six parameters considered in the study 204 in relation to the caving performance index, CPI, is determined by equation 8.6:

−0.2291 CPI = 73.38 − 3.085E − 0.095α + 48.21T − 29.75Sv − 132.17Sh + 0.0336H (8.6) where: E = Young’s modulus of the coal seam, MPa α = the dip of coal seam, degree T = the top coal thickness, m

Sv = spacing of vertical joints/cleats, m

Sh = spacing of bedding planes, m H = the cover depth of the coal seam, m.

In mining engineering, the strength of the coal and rockmass is usually represented by the UCS. Therefore, it is necessary to convert Young’s modulus, which presents the coal strength in equation 8.6, to the UCS. Dincer, Acar, Cobanoglu and Uras (2004) investigated the correlation between the uniaxial compressive strength and Young’s modulus for andesites, basalts and tuﬀs and proposed the relationship between the UCS and Young’s modulus, E, as in equation 8.7. This equation adapted well to the available data of the UCS and Young’s modulus of stratiﬁed rockmass. Therefore, in this study, equation 8.7 was used to convert the UCS to Young’s modulus of the coal seam.

E = 0.17UCS + 0.28 (8.7) As a result, equation 8.6 can be represented in equation 8.8 as follows:

−0.2291 CPI = 72.51−0.524UCS−0.095α+48.21T −29.75Sv −132.17Sh +0.0336H (8.8) where: UCS = intact unconﬁned compressive strength of coal, MPa.

In an attempt to quantify the scale of the cavability of top coal for the LTCC method from parametric geological conditions, a categorisation scheme which consists of 4 categories is suggested. The individual cavability classes in the the CPI system are characterised as follows: 205

• Class 1: Excellent cavability. Top coal caves immediately after the advance of the support, and good fragmentation of caved top coal is anticipated. Therefore, the maximum recovery rate of the top coal is expected. However, due to the easy cavability of the top coal, some problems, such as face instability or/and roof fall cavity, may occur in the roof area between the faceline and the canopy tips. Those potential hazards need to be managed during coal production.

• Class 2: Good cavability. Top coal caves almost immediately after the advance of the support, but a larger fragment size of top coal than that of class 1 is anticipated, which might cause diﬃculties for top coal drawing and transportation. As a result, a lower recovery of top coal may be anticipated. Fewer roof instability problems are expected compared to class 1.

• Class 3: Medium cavability. Poor fracturing of the top coal results in a delay in cavability, poor fragmentation, and low recovery rate of top coal. Artiﬁcial top coal fracturing techniques may be needed to improve the recovery rate of the top coal.

• Class 4: Poor cavability. The strong top coal prevents the caving process, and roof hang-ups or delay in top coal caving are expected. Very poor fragmentation is anticipated and the recovery rate of the top coal is very small even if artiﬁcial fracturing techniques are applied.

In order to classify the cavability of top coal in equation 8.8 for application in real life situations, the determination of individual scale categories was achieved by comparing the results obtained from 3DEC analysis to realistic case histories. The main information of LTCC practices, used for back analysis in this study, was the caving performance obtained from Chinese LTCC practices and the geological database provided by Jia (2001). Table 8.15 presents the CPI values, coupled with some other cavability assessment methods in relation to the individual geological parameter of some Chinese LTCC faces. 206 (%) 77.2 55.3 58.6 79.4 21.5 60.7 61.9 11.7 86.9 68.0 50.8 64.8 50.6 CPI d metho 86.5 68.6 61.1 98.1 33.6 63.3 68.7 33.4 76.4 62.2 88.8 56.2 102.9 akili’s V d metho y assessment methods 0.73 0.71 0.61 0.84 0.51 0.63 0.69 0.49 0.76 0.63 0.70 0.74 0.66 abilit v Ca Shangdon’s d 1 2 2 1 3 2 2 4 1 2 2 1 2 metho Jia’s (m) coal 1.4 1.4 2.8 5.22 1.54 2.36 1.14 1.45 1.571 1.679 1.462 1.589 2.674 op kness T thic (m) 0 0 0.2 0.4 0.2 0.2 0.3 0.2 0.4 0.3 0.3 0.12 0.15 Band kness thic (m) t/Cleat 0.15 0.35 0.25 0.15 0.45 0.25 0.25 0.55 0.15 0.25 0.35 0.25 0.25 Join spacing depth 150 480 250 130 210 240 254 350 230 360 412 210 150 (m) er v Co coal a) 7 8 of 20 30 38 30 25 32 20 24 30 7.5 17.5 (MP UCS assumptions: Comparision of CPI and diﬀerent cavability assessment methods (modiﬁed after Xu, 2004; Vakili, 2009) uang Mine in Yanzhou Mine, Xingtai 2 Mine, Datong ng Mine, Xuangang ejing Mine, Xuzhou site Mine, Shengyang Shuiyu Mine, Fengxi hung, Zhendzhou 1 Mine, Yanquan No 1 Mine, Pinding Shan 4 Mine, Yanquan angzhuang Mine, Luan enghuangshang Mine, Jincheng No Mic 11, mine Puhe No F No W liujialia Shangh Table 8.15: of Constant strength of the band layer and overlying strataCutting conditions height for is all 3.0m, cases, and The seam dip is 0 degrees for all cases. 3, 8, 15, 15, 3, 4, 3, 11,12, 5, 2, Xingtai 3, Xinlongzh 10, 7, Some following • • • Name Seam Seam Seam Seam Seam Seam Seam Seam Seam Seam Seam Seam Seam 207

As can be seen in Table 8.15, there is a good correlation between the CPI and some other classiﬁcation methods. Figure 8.2 plots the relationship between the cavability category developed by Jia (2001) the CPI method.

Figure 8.2: Comparision between Jia’s classiﬁcation and the CPI method

The CPI values also have a reasonable correlation with the real recovery rate obtained from some LTCC practices reported by Humphries et al. (2006) and Vakili (2009). The relationship between the CPI and the recovery rate is presented in Table 8.16 and Figure 8.3.

Figure 8.3: The relationship of the CPI vs the recovery rate of top coal in Chinese LTCC faces 208

Table 8.16: Comparison between the database of Chinese LTCC mines and the CPI values (modiﬁed after Zhongming et al., 1995; Humphries et al., 2006; Vakili, 2009)

Mine/Face UCS Cover Cleat Top coal Recovery CPI (MPa) depth (m) spacing (m) thickness (m) rate (%) (%) Yangquan No.1 14.9 360 0.44 3.5 73.2 44.6 Yinying No.15 8.5 220 0.25 3.8 90.5 70.5 Wangzhuang No.3 16 200 0.35 4 75.05 49.3 Xingzhouyao 30 300 >0.47 6 36.3 <22.7 Fenghuangshan 35 140 >0.47 3 37 <20.2 Liujialiang No.5 2.8 140 <0.18 3.5 85.5 >82.3 Shuiyu No.10-11 6.5 190 0.25 4.7 60 68.8 Xianshan No.5 8.5 230 0.25 3 82.3 72.4

In comparison with the case histories from the LTCC practices, the categorisation of the caving performance of the LTCC method based on the CPI values can be expressed according to the back analysis of recovery rate (Table 8.17).

Table 8.17: Cavability classiﬁcation of the top coal based on the CPI values

Category 1 2 3 4 Classiﬁcation rating Excellent Good Medium Poor Caving performance index (PCI)(%) >75 50 to 75 25 to 49 <25

To weigh the effect of the individual parameters on the overall cavability of the top coal, the categorisation of each parameter must be rated. Based on the evaluation of the thick seam database in the Quangninh coalfield and analysing the previous classifications, the affecting parameters can be categorised, as shown in Table 8.18. From the categorisation of the individual parameters in Table 8.18 and the equation 8.8, the weighting of each parameter in the CPI criterion is estimated, as can be seen in Table 8.19. 209

Table 8.18: Categorisation of individual parameters in the CPI criterion

Category Coal Strength Seam dip Seam vertical/Cleat Bed (horizontal) Cover (UCS) (MPa) (degree) thickness (m) spacing (m) spacing (m) Depth (m) 1 <15 <20 <5.0 <0.20 <0.20 >350 2 15 - 25 20 - 25 5.0 - 8.0 0.20 - 0.35 0.20 - 0.25 200 - 350 3 25 - 35 25 - 30 8.0 - 10.0 0.35 - 0.50 0.25 - 0.32 100 - 200 4 >35 >30 >10.0 >0.50 >0.32 <100

Table 8.19: Weighting of individual parameters in CPI criterion Parameter Weighting (%) UCS 13.1 Seam dip 2.4 Seam thickness 31.8 Vertical spacing 10.4 Bedding (Horizontal spacing) 34.4 Cover depth 7.9

As can be seen from Table 8.19, the bedding spacing, seam thickness and coal strength are the three most significant factors influencing the cavability of the top coal. The seam dip has a minor effect on the cavability of the top coal compared to the other parameters.

8.5 Other case studies

The back analysis of the proposed cavability assessment method discussed in previous sections are on flat coal seams. To evaluate the applicability of the proposed method in steeply dipping conditions, another source from the Quangninh coalfield where the seams are in steeply dipping conditions was also used for back analysis in this study. The data was obtained from field visits and information available from previous reports in some of the underground coal mines in Quangn- inh. The data includes geological/geotechnical information of individual mines such as stratigraphical maps, drillhole logs, properties of the coal seam, joint 210 properties, and caving performance witnessed during mining operations.

8.5.1 No.8 seam Tay Vangdanh, Vangdanh Coal Company

No.8 seam Tay Vangdanh is owned by the Vangdanh Coal Company, and is located 10km north of Uongbi town, Quangninh province. The seam has a thickness of 4.59 - 9.14m, with an average of 7.3m, and an average dip of 140. There are 1 - 5 band layers in the seam, the total thickness of which ranges from 0.1 to 1.08m, with an average of 0.6m. The coal is reported to have a strength slightly less than 30MPa. A detailed measurement of the uniaxial compressive strength (UCS) of some samples from the coal seams undertaken by the author shows that the average coal strength is 247 - 250kg/cm2, approximately equal to 25MPa. The immediate roof consists of 0.4 - 0.7m of fine-grained sandstone and shale overlain by sandstones and coal bands. The immediate floor also consists of fine-grained sandstone. Figure 8.4 presents some typical stratigraphic sections of the No.8 seam, Tay Vangdanh.

Figure 8.4: Some lithological sections of No.8 seam, Tay Vangdanh (source: Vangdanh Coal Company, 2006)

At the time of the mine visit (July 2007), the coal seam was extracted by the 211

LTCC method at a level of +50m to +100m compared to sea level, equivalent to a cover depth of approximately 220 - 270m from the surface, an average of 250m. The face retreated along the strike with an extraction height of 2.2m, supported by semi-mechanized shields (as shown in Figure 4.3). An (2007), the head of the longwall face, stated that the monthly production of this longwall face was from 12,000 to 15,000 tonnes and the coal recovery rate was from 70 to 80%. From underground observation it was noted that pronounced bedding planes were lying parallel to the seam dip and had a spacing of 200 - 300mm. The cleats/joints are orthogonal at approximately 700 vertical to the bedding planes, but they are not clearly visible. The distance between the cleats is estimated to be 100 - 800mm, with an average of 300mm. Figure 8.5 shows the conditions of No.8 seam at diﬀerent locations of the longwall face. Observation from the underground inspection also shows that the caving angle of the top coal ranges from 65 to 800. The common fragment size of the caved coal ranges from 80 - 400mm. However, many large caved lumps were observed in the goaf area with the predicted maximum size being 1000mm. Figure 8.6 presents the diﬀerent status of caved coal behind the supports of the longwall face. 212

Figure 8.5: Face conditions in No.8 seam Tay Vangdanh 213

Figure 8.6: Status of caved coal behind the longwall face 214

In order to evaluate the applicability of the CPI system, the geological information of No.8 seam, Tay Vangdanh was used to assess cavability. Because the top coal section has some band layers which have a diﬀerent strength compared with the strength of the top coal, it is necessary to determine the average value of UCS in top coal. To calculate the average UCS of the top coal, the typical lithological section of No.8 seam can be represented, as illustrated in Figure 8.7.

Figure 8.7: Typical lithological section of No.8 seam, Tay Vangdanh

Because there is no information of the UCS value in each ply and each band layer, it is assumed that the UCS values in plies are the same as the average strength, and the strength of the band layers is 35MPa, as reported by previous study. The average UCS of the top coal is determined as follows:

(0.6x25)+(0.3x35)+(3.2x25)+(0.1x35)+(0.3x25)+(0.6x35)+(2.2x25) UCSa = 0.6+0.3+3.2+0.1+0.3+0.6+2.2 = 26.4MP a

The cavability evaluation of this coal seam by the CPI method is summarised in Table 8.20. Results from the cavability evaluation indicate that the CPI is in class 2. It can be predicted from the results of the CPI that, good cavability is anticipated in this coal seam. Large fragment sizes of caved top coal are sometimes expected in this category. This result agrees with observations at the mine site. It should be noted that because there is no shield cover at the rear of the longwall face, and the caved coal and broken rockmass were prevented from coming to the working 215 face by hydraulic props and wire-mesh (as can be seen in Figure 8.6), drawing these large caved coal lumps from the working face is possible when the face supported by semi-mechanized shields. This can explain why the recovery rate achieved is still high (70 - 80%) in realistic operations, despite many large caved coal lumps in the goaf area. However, when applying mechanized shields which have smaller opening windows (which are safer for the operators at the longwall face), the recovery rate may be reduced, as drawing the large caved coal lumps will be diﬃcult.

Table 8.20: Cavability assessment of No.8 seam Tay Vangdanh by the CPI method

Name of parameter Value Coal strength (UCS) MPa 26.4 Seam dip (α), degree 14

Bedding spacing (Sh), m 0.25

Cleat spacing (Sv), m 0.30 Top coal thickness (T), m 5.1 Cover depth (H), m 250 CPI value 57.0 Cavability rating Good (class 2)

8.5.2 No.6b seam, Lotri, Thongnhat Coal Company

No.6b seam, in the Lotri area, is managed by Thongnhat Coal Company and is located approximately 15m north of Campha town, Quangninh province. The working thickness of the No.6b seam ranges from 14.5 to 29.5m, with an average of 19.4m. The seam has a dip range from 14 to 280, with an average of 210. The coal is reported to have a strength of 200 - 250kg/cm2, equivalent to 20 - 25MPa. There are 4 - 5 band layers present in the coal seam, the thickness of each band layer ranges from 0.15 - 1.85m, depending on speciﬁc area, and the strength of the band layer ranges from 320 - 350kg/cm2. The immediate roof consists of 0.5 216

- 0.8m of mudstone overlain by sandstone and coal bands. The immediate ﬂoor also consists of mudstone layers. The geological section of No.6b seam in the Lotri area, Thongnhat Coal Company is shown in Figure 8.8.

Figure 8.8: Geological section of some boreholes in No.6b seam, Lotri, Thongnhat Coal Company (source: Thongnhat Coal Company, 2005)

According to the Company’s plan, the thick seam is divided into three extraction slices. The upper slice was extracted by the LTCC method with an average thickness of 5.0m. The current longwall face, at the time of the visit (July 2007), was extracting the second slice with an average thickness of 7.2m, and a cover depth from -35 to +8 (as compared to sea level), equal to a depth cover of 300m from the surface. The bottom slice is planned to have an average thickness of 7.2m. The LTCC face was 120m in length, supported by model ZH1600/16/24Z semi-mechanised shields (as shown in Figure 4.7), and the extraction height of the face was 2.5m. Coal at the face was extracted by the drilling and blasting method. Top coal drawing and dilution control during the caving process was via 217 visual inspection and manual operations. Observation at the longwall face found that there were visible bedding planes in the coal seam, the spaces of the bedding planes were measured 0.1 - 0.15m, and these spacings varied along the longwall face. The vertical spacings were measured 0.2 - 0.4m, with an average of 0.25m. Due to the concern of gas explosion, photographs were not permitted at the site. Through observation it was found that the top coal caves immediately after the advancement of the supports. The caved coal was fragmented with sizes ranging from 40 - 300mm; the larger sizes were also observed in some positions along the longwall face. Miners working in the face said that one of diﬃculties in top coal drawing was the dilution of the broken rocks from the upper slices. The caving angle was estimated to be 75 - 800, and the recovery rate was reported of 60 - 70%. To predict the cavability of the top coal, the CPI method is used. A summary of the geological parameters and the estimated rate of the caving performance is presented in Table 8.21.

Table 8.21: Cavability assessment of No.6 seam, Lotri area, Thongnhat Coal Company by the CPI method

Name of parameter Value Coal strength (UCS) MPa 22 Seam dip (α), degree 21

Bedding spacing (Sh), m 0.12

Cleat spacing (Sv), m 0.25 Top coal thickness (T), m 4.7 Cover depth (H), m 300 CPI value 79.6 Cavability rating Excellent (class 1) 218

8.5.3 No.7 seam, Thanthung, Nammau Coal Mine

No.7 seam is located at the Thanthung mine site, approximately 15km North- West of Uongbi town, Quangninh province. The mine is owned by Nam Mau Coal Company. The geological information, and the underground observations of caving performance conducted at this seam were mainly by Tuan and Thang (2003) during the process of evaluating the application of the LTCC method, supported by semi-mechanized shields. According to the technical designing report of the experimental application of the semi-mechanized shields in No.7 seam conducted by Dac, Tuan and Thang (2000), No.7 seam has a thickness from 5.2 to 10.5m, with an average of 6.4m. The seam dip ranges from 20 to 300, with an average of 250. There are 0 - 6 band layers in the coal seams, depending on particular locations. The total thickness of the band layers ranges from 0 to 0.6m, with an average of 0.3m. The immediate roof is the laminated mudstone layers and has a total average thickness of 6m. The cover depth of the seam is from 120 - 180m, with an average of 150m. The UCS test reported to have an average coal strength of 200 - 300kg/cm2, equal to approximately 20 - 30MPa. Some typical lithological sections of No.7 seam in some boreholes are shown in Figure 8.9.

Figure 8.9: Typical lithological section of 3 boreholes in No.7 seam, Thanthung, Nam Mau Coal Company (after Dac et al., 2000) 219

There is no information on the discontinuities of the coal seams. However, a previous study of the properties of discontinuities (Hoan, 1990) indicated that the bedding spacing and cleat spacing in the Thanthung area varies from 0.1 - 0.5m. In addition, the distance between the bedding planes can also be estimated from pictures taken during ﬁeld observations (as shown in Figure 8.10) which indicate that the discontinuity spacing ranges from 0.1 to 0.25m, with an average of 0.15m.

Figure 8.10: Face condition at No.7 seam Thanthung, Nam Mau Coal Company (source: Tuan and Thang, 2003)

The caving performance of this longwall face was reported by Tuan and Thang (2003) during the process of evaluating the applicability of the new supports into No.7 seam, Thanthung. This coal seam has been extracted by the LTCC method, 220 supported by semi-mechanized shields model XDY -1T2/LY made in China, as presented in Figure 4.3, since 2001. The longwall face had a cutting height of 2.0 - 2.2m, and the top coal section was drawn through the rear of the longwall face. In the designing stage, due to the uncertainty of the cavability of the top coal, it was planned that coal drawing would be undertaken after two times of face extractions (the width of each web was 0.8m), with the aim to have some time delay for the top coal to fracture. However, the reality of mining operations showed that after the advancement of the support (with the advance rate of 0.8m each time), the top coal started to cave. The common fragment sizes of caving coal in the goaf area ranged from 50 to 400mm, as presented in Figure 8.11. The drilling and blasting method was sometimes used to assist the fracturing of the top coal. The percentage of recovery rate was reported at 75 - 80%. In order to assess the practicability of the proposed cavability assessment method, the CPI method was used to evaluate the cavability of the top coal. Based on geological information, it is predicted that the top coal will cave readily after the advancement of the supports. In addition, very good fragmentation of the caved coal is anticipated. Result calculated from the CPI method is in reasonable correlation with observations from actual operations. Table 8.22 presents information of the geological parameters and the rate of the cavability using the CPI method.

Table 8.22: Cavability assessment of No.7 seam Thanthung by the CPI method

Name of parameter Value Coal strength (UCS) MPa 25 Seam dip (α), degree 25

Bedding spacing (Sh), m 0.15

Cleat spacing (Sv), m 0.15 Top coal thickness (T), m 4.2 Cover depth (H), m 150 CPI value 72.5 Cavability rating Good (class 2) 221

Figure 8.11: Caved coal size behind the longwall face (source: Tuan and Thang, 2003) 222

8.6 Conclusions

This chapter described the development of an assessment method to predict the cavability of the top coal when extracting inclined thick seams by the LTCC method. This cavability assessment method was developed by parametric studies combined with back analysis from LTCC practices. The classification criterion first assigned numerical values to six different geological parameters of the rockmass considered to be the most significant influence on the cavability of top coal in the LTCC method, by reference from the 3DEC modelling analysis. These parameters were: coal strength, top coal thickness, seam dip, spacing of the bedding planes, spacing of the joints/cleats, and the cover depth of the seam. Based on the results from the 3DEC analysis, an equation name as “caving performance index” (CPI) was formed through multiple-regression analysis to represent the numerical values of the effect of the above mentioned six different parameters on the overall cavability of the top coal in the LTCC method. Rating values of the CPI equation were then determined by back analysis. Back analysis was undertaken by correlating the values of the CPI equation to the caving performance obtained from Chinese and Vietnamese LTCC faces. The categorisation of cavability prediction was classified into four categorises ranging from 1 (excellent cavability) to 4 (very poor cavability). The analysis of case studies observed from the LTCC faces in practice confirmed the applicability of the CPI method for predicting the cavability of the top coal in the LTCC method. This cavability criterion can be used to assist mine planners in investigating the efficiency of the application of the LTCC method for the extraction of inclined thick seams from the geotechnical data collected in the feasibility stage of a mining project. In addition, this cavability method can also be used to predict the behaviour of coal or rock surrounding an underground excavation, so that the mine designers can outline the measures that help prevent potential problems during mining production. Chapter 9

Conclusions and recommendations

9.1 Summary and conclusions

9.1.1 Introduction

The extraction of thick coal seams by underground mining has long been a challenge in the Quangninh coalfield, because almost of all coal reserves in Quangninh are located in difficult geological conditions and many thick seams are steeply dipping. In addition, most of the underground mines in Quangninh are operating with low technology and simple equipment. The application of a mining method for inclined thick seams is mainly undertaken by implementing the methods which are applied to flat seams from overseas or other local coal mines, with little understanding of the operational and geotechnical issues associated with these methods when extracting dipping coal seams. As a result, there is a need to improve the understanding of operational and geotechnical issues associated with the application of underground thick seam mining methods in the steeply dipping conditions in the Quangninh coalfield. It is estimated that the demand for coal in the domestic market in Vietnam in the coming years will increase rapidly from 29 - 32 million tonnes in 2010 to 112 - 118 million tonnes in 2025, requiring a dramatic increase in coal production.

223 224

However, under the pressure of environmental protection of the area nearby, mining operations from openpit will have to be gradually scaled down and will end by 2014. Therefore, if Vietnamese coal operators are to fully exploit their thick seam deposits with higher production, and a higher standard of safety management, improvements in underground thick seam mining are essential. This thesis aimed to determine if success could also be achieved in Quangn- inh’s conditions when applying ﬁve methods (bord and pillar, high reach single pass longwall, multi slice longwall, hydraulic mining, and longwall top coal caving) and identifying the possible issues during mining operations. The main objectives of this research can be summarized as follows:

1. Evaluation of the most suitable underground thick seam methods applicable to inclined thick seams under the current mining conditions in the Quangninh coalﬁeld.

2. To improve understanding of the operational and geotechnical issues associated with the application of the LTCC method for the extraction of inclined thick seams.

3. To develop an assessment method to predict the cavability of the top coal when using the LTCC method to extract inclined thick coal seams.

This chapter summarises the conclusions of the study and gives recommendations for future work. Section 9.1.2 summarises the results from the investigation into the reasons why the LTCC method is most suitable for inclined thick seams under Quangninh’s geo-mining conditions. The operational issues associated with the application of the LTCC method in steeply dipping conditions is concluded in Section 9.1.3. Section 9.1.4 summaries the study into the effect of seam dip on the cavability of the top coal, and the differences in the caving mechanisms between flat and inclined seams. Section 9.1.5 presents a summary of an improved assessment method to predict the cavability of the top coal when using the LTCC method for extracting inclined thick seams. Recommendations for future work are given in Section 9.2. 225

9.1.2 Potential mining methods in the Quangninh coal- ﬁeld

The Quangninh coalfield, located approximately 150km North-Eastern of Hanoi, Vietnam, holds a large proportion of the coal resources in Vietnam. Coal resources in Quangninh to a depth of -300m (compared to sea level) are estimated to be 4.0 billion tonnes, and at a depth of greater than the level of -300m, are 6.9 billion tonnes. In order to have a clear view of the quality and quantity of thick coal resources in the Quangninh coalfield, this study analysed databases of the coal seams managed by six coal companies with total reserves of 560,845,000 tonnes, which will be the main target of coal production of the underground mining companies in the next decade. Results from the study found that a large proportion of coal reserves (57%) are thick seams (thickness greater than 3.5m), and 58% of those thick seams have a seam dip range from 15 to 350. 79% of thick coal seams are located under moderate or deep cover depth (greater than or equal to 200m), and most thick coal seams (88%) have a low or moderate gas content (less than 10m3 gas per tonne of coal). In addition, 48% of thick seam reserves have short length along the strike (less than 1000m) due to the appearance of many faults in the coalfield which may influence the application of mechanized mining equipment. Based on the results of thick seam analysis, this study evaluated the underground methods, currently applied around the world, that could potentially be applied to thick seams under Quangninh’s geo-mining conditions. These methods (room and pillar, high reach single pass longwall, multi-slice longwall, hydraulic mining, and longwall top coal caving) all offer possible solutions to the problems of thick seams. Around the world, all five methods have been used to exploit thick seams, with varying degrees of success being reported. There are a number of issues that must be considered in assessing the overall feasibility of a particular mining method. For this reason, this study has focused primarily on the operational viability of these mining methods under the current mining conditions in Quangninh. 226

The room and pillar method was considered to have a limited application in Quangninh coalfield due to a high percentage of coal loss, and the unsuitability of its application in high cover depth and inclined seams. The blasting gallery method, another modification of the room and pillar method, was also concluded to have limited application for extracting inclined thick coal seams in Quangninh’s conditions because of the difficulties in driving a large number of roadways along the seam dip and the high percentage of coal loss. The High reach Single Pass Longwall method was also considered to have limited applicability in the Quangninh coalfield at the present time due to a number of factors. These are: its large size and heavy weight, the high capital cost of mining equipment; the instability of the high face supports due to the dipping factor; and equipment relocation and transportation in mines with steeply dipping conditions. The multi-slice longwall method was also found to have many issues when applied to Quangninh. These issues are: mining under the goaf area, the stability and protection of roadways, and the difficulty in operating separated slices in coal seams with high variations in thickness. For these reasons, this mining method has only a limited application for the coal seams that need the pre-fracturing of the immediate roof prior to the extraction of the lower slices in thick coal seams. Hydraulic mining was identified to have significant potential in complicated geological conditions where mechanized methods face operational difficulties or are economically unfeasible. However, hydraulic mining would face many operational challenges in the Quangninh coalfield under the current level of technology and the infrastructure existant in the mines. The first potential issue that may arise when applying hydraulic mining in Quangninh, is the low technology in coal processing, currently operating, that faces difficulties in beneficiating a large proportion of fine coal product created from hydraulic mining. Coal products with a high percentage of fine coal, if not beneficiated, will have a high percentage of dilution and result in a low market value. The other operational issue is the complications in the transportation of the mixture of coal and water from the mine site to the coal preparation station and the transportation of the recycled 227 water back to the mine site. This is due to the long distance between the coal preparation station and the mine sites. The difficulty in the treatment of spillage and contaminated runoff water discharged from hydraulic mining is also a issue when applying this method in the Quangninh coalfield. In addition, the high rate of roadway development per tonne of coal production is also a challenge under the current technology in Quangninh. In the future, when the technology of coal processing and roadway development are improved, it is believed that hydraulic mining will have further chances to prove its viability in Quangninh’s geo-mining conditions. The LTCC method has been proved to have the greatest potential for the extraction of thick seams in the Quangninh coalfield, as it overcomes the disadvantages of the multi-slice longwall and the high reach single pass longwall methods. Compared to the multi-slice longwall method, the LTCC method significantly reduces development costs per tonne due to more coal being extracted from the upper section during mining at the lower section of a thick seam. The LTCC method enables up to 80 - 85% seam extraction in the 5 - 9m thickness range, including thick seams with high variation in thickness, which the multi-slice longwall method has difficulty in exploiting. Compared to high reach single pass longwalling, the LTCC method offers a lower face height, resulting in smaller and less expensive equipment, and better face conditions. Hence, the LTCC method is considered the most suitable method for the extraction of thick coal seams in Quangninh.

9.1.3 Operational issues associated with the application of the LTCC method in steeply dipping conditions

This study has identified a number of operational issues which may arise when applying the LTCC method for the extraction of inclined thick seams in the Quangninh coalfield, and offered practical solutions to these problems through analyses of current operational practice. The first is the direction of the longwall face. Compared to a flat coal seam, 228 when extracting an inclined thick seam, the face can retreat in one of three different directions: up-dip, down-dip or along the strike. Among those three directions, the face retreating along the strike has operational advantages over the face retreating up-dip or down-dip. The first advantage is the heading gates (the maingate and the tailgate) of the panel for the face retreating along the strike are horizontal, which makes driving them is more convenient than those of the face retreating up-dip or down-dip of the seam, where the heading gates are in steeply dipping conditions. For that reason, the longwall panel for the face retreating along the strike is likely to be longer and take less time in development than the panel developed along the seam dip. As a result, less frequency in face movement is anticipated for the face retreating along the strike compared to the face retreating up-dip or down-dip. Another advantage is the convenience of the transportation and gathering of supports and other mining equipment in the horizontal gate roadways. Therefore, from an operational point of view, it is recommended that for the extraction of inclined thick seams, the face should retreat along the strike. A critical review of the current level of the mechanization of thick seam mining in the Quangninh coalfield showed that over the last decade, its technology and management have been improved. These improvements resulted in a dramatic increase in coal production from 9.4 million tonnes in 1994 to nearly 35 million in 2005. However, this improvement was still limited in capacity and safety management. The increase in total coal production was mainly based on the increase in the number of longwall faces. In addition, the maximum output of the longwall face is low compared to the longwall faces in developed countries. The main reason for the limits in coal production and safety management is the inappropriate mining equipment used. Mining operations require a number of processes undertaken by manual work, resulting in low production and productivity. In the coming year, the Vietnamese economy’s demand for coal product will increase rapidly and this requires an improvement in the mining industry. As a result, there is a need to improve the underground mining sector, especially in underground thick seam mining. In order to ultimately prevent potential prob- 229 lems and realise the potential revenue in the thick seams in Quangninh, it is necessary to improve the understanding of the operational issues associated with the application of mechanized mining equipment for the LTCC face in the steeply dipping conditions of the Quangninh coalfield. Of the two shield designs (European and Chinese designed shields) which are used in mechanised LTCC faces, the Chinese designed shields are considered appropriate for LTCC faces. The European designed shields were proved to have good stability and provide a safe working environment for operators. However, they produce a large concentration of dust on the face during the caving process because the caving chute is located at the middle height of the canopy. The Chinese shield design, on the other hand, attracts a greater recovery rate because the rear caving canopy is long and the window for top coal drawing is large. In addition, due to the caving point being located at a lower height of the face, the creation of dust by the caving process is much reduced compared to the European designed shields. For the face retreating along the strike, some potential problems may exist when using the LTCC method to extract thick coal seams in steeply dipping conditions compared to horizontal seams. The biggest problem faced by the operations of the LTCC face in steeply dipping conditions is the instability of the supports and other equipment (tilting and slippage) during operations in steeply dipping conditions. Another problem is transporting the supports and other equipment in different dipping conditions. As a result, the supports in steeply dipping conditions should have sufficient lateral stiffness to maintain stability against any loading parallel to the face line (to prevent support tilting and slippage). In addition, they must provide adequate bearing pressures at the roof and floor interfaces to avoid excessive fracturing, compacting or movement of the strata that would compromise ground control or support stability. The support should have the ability to provide adequate roof and goaf cover, to avoid the hazardous infiltration of roof material into the working area. The support also should have a reasonable weight so that transport of mining equipment in mine or the movement of the longwall face is economically reasonable. 230

In addition, the transition between the face ends and the gate roadways are considered weak points that may create a labour intensive and safety issues, and also reduce the advance rate. As a result, the transition between the face ends and the gate roadways is recognized as one of the big issues for the LTCC method applied to the extraction of inclined thick seams and needs a considerable amount of further investigation.

9.1.4 Geotechnical issues associated with the application of the LTCC method in steeply dipping conditions

In terms of geotechnical issues, this study focused on investigating the factors influencing the caving mechanism of top coal during the LTCC operation. A detailed literature review of previous studies in cavability and stability assessment showed that the major factors influencing the cavability of the top coal are: coal strength, top coal thickness, the cover depth, persistence and intensity of discontinuities, the roof strata, and the moisture sensitivity. However, the effect of the seam dip on the caving mechanism of the top coal and above rockmass has not been investigated thoroughly. As a result, this study focused on investigating the effect of the seam dip, combined with other factors, on the caving mechanism of top coal in the LTCC method. The first attempt to improve understanding of the effect of seam dip on the caving mechanism of top coal in the LTCC method was to investigate the effect of the orientation of the longwall face on the caving mechanism of the top coal and above rockmass. The study was undertaken with reference to results of numerical analysis (UDEC software),based on discontinuum methods. Different geometric models were simulated to represent a cross-section at the centre point of the longwall face retreating in three directions: along strike, up-dip, and down-dip. Results from the study found that, for the extraction of inclined thick seams, better cavability of the top coal is anticipated for the face retreating along the strike than that of the face retreating down-dip or up-dip of the inclined seam. As a result, it is recommended that for the extraction of inclined thick seams, the 231 longwall face should retreat along the strike. In order to investigate the differences in the caving mechanism of top coal between flat and inclined coal seams when the face retreats along the strike, the 3DEC programme, a three dimensioned numerical code, was also undertaken. The aim of the study was to gain a greater insight into the effect of seam dip, coupled with other factors, on the cavability of the top coal for the LTCC method. Numerical modelling analysis showed that in the case of the face retreating along the strike, better cavability of top coal for flat coal seams is predicted compared to inclined seams. In a comparison of the cavability in different parts of the top coal along the longwall face, results illustrated that for flat seam cases, better cavability of the top coal is predicted at the middle of the longwall panel, and a similar cavability of the top coal near both endgates of the longwall face (near the maingate and tailgate) is expected, while for inclined seams, better caving performance in the uphill portion of the longwall face is anticipated than in the downhill portion. Another geotechnical aspect studied by this research was to investigate the differences in distribution of the abutment stress in the chain pillar in the next panel, down-dip of the longwall face already mined. FLAC software, a finite differential element code, was used for this analysis. The result from the FLAC analysis concluded that the location of the abutment stress in the chain pillar, downdip side of the longwall panel is influenced by the dipping factor. For flat coal seams, the peak stress area is located within the coal seam with different values of horizontal stress, while for inclined seams, the location of the abutment stress area is largely dependant on the ratio between the horizontal and vertical stress. If the horizontal stress is smaller than the vertical stress, the abutment stress area moves towards the roof rockmass area above the coal seam. As the horizontal stress increases, the abutment stress gradually moves downward to the coal seam area. However, all results showed that the peak vertical stress of an inclined seam is smaller than that of a flat coal seam. This result indicated that, under the same geological conditions, the gate roadways located in the chain pillar, downdip of the longwall panel of an inclined thick seam are subjected to less 232 re-distribution of the vertical stress than that of a flat coal seam. Results from the FLAC analysis further confirmed the conclusions from the 3DEC analysis, that the seam dip has an effect on the caving mechanism of the top coal and above rockmass when extracting inclined thick seams by the LTCC method. The FLAC analysis also showed that for flat coal seams, better cavability is anticipated in the middle portion of the longwall face, while for inclined seams, better cavability is predicted in the uphill portion of the longwall face than in the downhill portion. In addition, better cavability of the top coal and above rockmass is anticipated in flat rather than inclined coal seams.

9.1.5 Cavability assessment method for extracting inclined thick seams by the LTCC

Two important factors that influence the success of the LTCC method is the accurate estimation of the cavability of the top coal in the designing stage so that the mine planners can predict the efficiency of the LTCC method when applying it in certain geo-mining conditions and identification of the potential risks that need to be managed. In addition to these, the impact of various parameters on the cavability of top coal can be understood. This study also developed an assessment method for predicting the cavability of top coal (the CPI value) based on six geological parameters, which were: coal strength, top coal thickness, seam dip, spacing of the bedding planes, spacing of the joints/cleats, and the cover depth of the seam. The categorisation of cavability prediction was classified into four categorises ranging from 1 (excellent cavability) to 4 (very poor cavability). The determination of the CPI was achieved by conducting numerical analysis on the effect of the six parameters influencing the cavability of top coal, combined with back analyses of the caving performance witnessed in the LTCC practices. While it is recognised that several other parameters might also influence the defined cavability categories, the chosen parameters allowed for both a systematic investigation of the caving mechanism and the prediction of cavability assessment design options. 233

9.2 Recommendations for further research

Hydraulic mining, identified in this study, is one prospective method with a high potential for complicated geological conditions that needs to be further studied. Further research should address the mechanics of the waterjet (particularly in jet pressure and jet impact force) of the hydraulic monitor so it can both break the high strength coal seam in Quangninh coalfield, and also to reduce dilution by avoiding breaking the weak immediate roof or the floor located next to the coal seam. In addition to this, improvements in coal processing, especially the process of de-watering the fine coal from the mixture of water and fine coal, and the treatment of slippage and contaminated runoff water should be addressed. In parallel, a study on improvements in roadway development also needs to be undertaken in the Quangninh coalfield so that the rate of roadway development can meet the rate of extraction by the hydraulic mining method. It is believed that when these issues are improved, hydraulic mining will have further chances to prove its potential in Vietnam in a large range of thick seam resources. In terms of stress distribution in the chain pillar of the inclined seams, this study was limited in investigating the differences in stress distribution in the chain pillar between flat and inclined seams by reference to results from numerical analysis. Further study should be undertaken to identify the extent of the difference in the stress field or stress concentration in the longwall gate-ends (the maingate and tailgate) between flat and inclined seams. In addition, there is a need to further study on the relationship between the stress distribution in front and besides of a longwall face. It was identified in this study that the transition between the face ends and the longwall face are the weak points during mining operations, especially in roof control and maintaining the stability of supports in the inclined face. Further study is needed to investigate roof control at the transition between the longwall face and the gate roadways. It is believed that the cavability of the top coal is influenced by numerous parameters. The assessment method for predicting the cavability of the top coal 234 developed in this study only covers some of the major parameters of the geological/geotechnical properties of top coal, while some important factors that may influence the cavability of top coal have not been examined. Some examples of the factors that have not been investigated are: the persistence of discontinuities, joint/cleat orientations, the effect of moisture content in the top coal or rockmass above, the effect of horizontal stress, and the characteristics of immediate and main roof strata. Future studies may need to investigate those factors as adjustments to the categorisation of the cavability assessment method. The back analysis of the cavability assessment method in this study was limited by the database from Chinese LTCC face practices, which were primary flat coal seams, with only some cases for inclined thick seams in the Quangninh coalfield. To increase the reliability of the developed cavability assessment method, more case histories of the LTCC face in the Quangninh coalfield should be added to the cavability method. It is important for individual cases to include three major components: geological and geotechnical information, caving performance from observation, and cavability assessment and ratings based on the cavability system. References

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Fish functions used in numerical modelling analysis

This appendix contains FISH functions (a programming language embedded within numerical programme to enable performing user deﬁned tasks) which were used in the 3DEC, FLAC analyses.

A.1 Fish functions to generate the support elements in 3DEC modelling

Fish function to determine the location of the support elements ef chock gridpoint can grid x = b x(canopy) can grid y = b y(canopy) base grid x = b x(base) base grid y = b y(base) end

Fish function to set the support elements against the roof def chock up command

250 251 free reg 313 ; (reg 313 representing the canopy and the base of support) endcommand chock gridpoint first point = can grid y second point = can grid y limit y = 0.2*cos(alfa*degrad) cc = b gp(canopy) bb = b gp(base) vel z1 = -10*sin(alfa*degrad) vel y1 = 10*cos(alfa*degrad) vel z2 = 10*sin(alfa*degrad) vel y2 = -10*cos(alfa*degrad) loop while cc # 0 gp xvel(cc)=0 gp zvel(cc)=vel z1 gp yvel(cc)=vel y1 cc = gp next(cc) endloop ; << endloof of cc >> loop while bb # 0 gp xvel(bb)=0 gp zvel(bb)=vel z2 gp yvel(bb)=vel y2 bb = gp next(bb) endloop ; << endloop of bb >> distance = abs(abs(second point) - abs(first point)) loop while (distance) < limit y command step 5 endcommand chock gridpoint second point = can grid y 252 distance = abs(abs(second point) - abs(first point)) endloop chock gridpoint cc = b gp(canopy) bb = b gp(base) loop while cc # 0 gp xvel(cc)=0 gp zvel(cc)=0 gp yvel(cc)=0 cc = gp next(cc) endloop ; << endloof of cc >> loop while bb # 0 gp xvel(bb)=0 gp zvel(bb)=0 gp yvel(bb)=0 bb = gp next(bb) endloop ; << endloop of bb >> command bound reg 313 yvel 0 bound reg 313 xvel 0 bound reg 313 zvel 0 fix reg 313 step 1 endcommand end

Fish function to lower the support elements def chock release command free reg 313 253 endcommand chock gridpoint first point = can grid y second point = can grid y limit y = 0.2*cos(alfa*degrad) cc = b gp(canopy) bb = b gp(base) vel z1 = 10*sin(alfa*degrad) vel y1 = -10*cos(alfa*degrad) vel z2 = -10*sin(alfa*degrad) vel y2 = 10*cos(alfa*degrad) loop while cc # 0 gp xvel(cc)=0 gp zvel(cc)=vel z1 gp yvel(cc)=vel y1 cc = gp next(cc) endloop ; << endloof of cc >> loop while bb # 0 gp xvel(bb)=0 gp zvel(bb)=vel z2 gp yvel(bb)=vel y2 bb = gp next(bb) endloop ; << endloop of bb >> command fix reg 313 endcommand distance = abs(abs(second point) - abs(first point)) loop while (distance) < limit y command step 5 endcommand 254 chock gridpoint second point = can grid y distance = abs(abs(second point) - abs(first point)) endloop chock gridpoint cc = b gp(canopy) bb = b gp(base) loop while cc # 0 gp xvel(cc)=0 gp zvel(cc)=0 gp yvel(cc)=0 cc = gp next(cc) endloop ; << endloof of cc >> loop while bb # 0 gp xvel(bb)=0 gp zvel(bb)=0 gp yvel(bb)=0 bb = gp next(bb) endloop ; << endloop of bb >> command bound reg 313 yvel 0 bound reg 313 xvel 0 bound reg 313 zvel 0 step 1 fix reg 313 step 1 endcommand end 255

Fish function to advance the support elements def chock advance command free reg 313 endcommand chock gridpoint first point = can grid x second point = can grid x cc = b gp(canopy) bb = b gp(base) loop while cc # 0 gp xvel(cc)=30 gp zvel(cc)=0 gp yvel(cc)=0 command print cc endcommand cc = gp next(cc) endloop ; << endloof of cc >> loop while bb # 0 gp xvel(bb)=30 gp zvel(bb)=0 gp yvel(bb)=0 bb = gp next(bb) endloop ; << endloop of bb >> command fix reg 313 endcommand distance = abs(abs(second point) - abs(first point)) loop while (distance) < 1.0 256 command step 3 endcommand chock gridpoint second point = can grid x distance = abs(abs(second point) - abs(first point)) endloop chock gridpoint cc = b gp(canopy) bb = b gp(base) loop while cc # 0 gp xvel(cc)=0 gp zvel(cc)=0 gp yvel(cc)=0 cc = gp next(cc) endloop ; << endloof of cc >> loop while bb # 0 gp xvel(bb)=0 gp zvel(bb)=0 gp yvel(bb)=0 bb = gp next(bb) endloop ; << endloop of bb >> command bound reg 313 yvel 0 bound reg 313 xvel 0 bound reg 313 zvel 0 fix reg 313 step 1 endcommand end 257

A.2 Fish functions to generate the top coal caving in FLAC modelling def caving ;;define xbegin, xending for zones in i ;;define ybegin, yending, y roof and jtopcoalzn for zones in j array ixx(5000) jyy(5000) searchingzone = 0 ; loop while searchingzone # 1 xbegin = xbegin xending = face advance xxbegin = xbegin xxending = xending total yield zone = 0 zn all yield = 0 row yield zone = 0 removed zone = 0 yielding zone = 0 loop jj (ybegin,yending) xsearching = 1 loop ii (xbegin,xending) ;searching to behind the support if model(ii,jj) = 1 then if xsearching = 1 then xxbegin = ii xxending = ii xsearching = 2 else if ii >= xxending then xxending = ii endif endif ;xsearching = 1 258 endif ;end model = 1 if model(ii,jj) # 1 then if xsearching = 1 then if syy(ii,jj) >= 0.0 then ;tension(ii,jj) then total yield zone = total yield zone + 1 ;all zones row yield zone = row yield zone + 1 ; for row ii search zn all yield = zn all yield + 1 ;zone search xxbegin = ii xxending = ii xsearching = 2 endif ;syy(ii,jj) if syy(ii,jj) < 0.0 then if state(ii,jj) = 1 then total yield zone = total yield zone + 1 ;all zones row yield zone = row yield zone + 1 ; for row ii search zn all yield = zn all yield + 1 ;zone search xxbegin = ii xxending = ii xsearching = 2 endif ; state(ii,jj) if state(ii,jj) = 2 then total yield zone = total yield zone + 1 ;all zones row yield zone = row yield zone + 1 ; for row ii search zn all yield = zn all yield + 1 ;zone search xxbegin = ii xxending = ii xsearching = 2 endif ; state(ii,jj) if state(ii,jj) = 3 then total yield zone = total yield zone + 1 ;all zones row yield zone = row yield zone + 1 ; for row ii search 259 zn all yield = zn all yield + 1 ;zone search xxbegin = ii xxending = ii xsearching = 2 endif ; state(ii,jj) endif ;syy(ii,jj) < 0 endif ;xsearching = 1 if xsearching # 1 then if syy(ii,jj) >= 0.0 then total yield zone = total yield zone + 1 ;all zones row yield zone = row yield zone + 1 ; for row ii search zn all yield = zn all yield + 1 ;zone search if ii >= xxending then xxending = ii endif end if ;syy(ii,jj) if syy(ii,jj) < 0.0 then if state(ii,jj) = 1 then total yield zone = total yield zone + 1 ;all zones row yield zone = row yield zone + 1 ; for row ii search zn all yield = zn all yield + 1 ;zone search if ii >= xxending then xxending = ii endif endif ;end state(ii,jj) if state(ii,jj) = 2 then total yield zone = total yield zone + 1 ;all zones row yield zone = row yield zone + 1 ; for row ii search zn all yield = zn all yield + 1 ;zone search if ii >= xxending then xxending = ii 260 endif endif ;end state(ii,jj) if state(ii,jj) = 3 then total yield zone = total yield zone + 1 ;all zones row yield zone = row yield zone + 1 ; for row ii search zn all yield = zn all yield + 1 ;zone search if ii >= xxending then xxending = ii endif endif ;end state(ii,jj) endif ;end syy < 0.0 end if ;end xsearching # 1 endif ;end model # 1 end loop ;ii loop ;;removing the caving zones if row yield zone # 0 then jlower = jj - 1 loop iii (xxbegin,xxending) zone below = model(iii,jlower) if zone below = 1 then if model(iii,jj) # 1 then command model null i=iii j=jj end command total zone del = total zone del + 1 ixx(total zone del) = iii jyy(total zone del) = jj removed zn = removed zn + 1 yielding zone = yielding zone + 1 if jj <= jtopcoalzn then removed cl = removed cl + 1 261 else removed rc = removed rc + 1 endif ;jtopcoal endif ;model endif ;zone below end loop ;iii row yield zone = 0 ;for next row endif ; end row yield zone end loop ;jj loop if removed zn = 0 then searchingzone = 1 endif if removed zn # 0 then get equilibium = 1 total steps = 0 loop while get equilibium = 1 total steps = total steps + 2000 command step 2000 endcommand uu = unbal if uu <= 5000 then get equilibium = 2 endif if total steps >= 40000 then get equilibium = 2 endif endloop ; end get equilibium end if ;end if removed zn # 0 if yielding zone = 0 then searchingzone = 1 262 endif ;; backfill the caving area if removed zn > 0 then xxending = face advance xsearching1 = 1 jcuting = zn j bottom loop iiii (xbegin,xxending) if xsearching1 = 1 then if model(iiii,jcuting) = 1 ;check null zones in row xfill begin = iiii xfill ending = iiii xsearching1 = 2 endif ; end model endif ; end xsearching1 = 1 if xsearching1 # 1 then if model(iiii,jcuting) = 1 then if iiii >= xfill ending then xfill ending = iiii endif endif ; end model endif ; end xsearching1 end loop ;loop iiii xcuting = xfill ending - xfill begin + 1 bulk fac = 1.5 rock caving =(removed cl*0.2+removed rc)*bulk fac caved zn=(removed cl*0.75)+(removed rc)+(3*xcuting*0.95) command print rock caving caved zn print xfill ending xfill begin endcommand 263

if rock caving >= caved zn then y fillstart = ybegin y topcoal = jtopcoalzn xsearching2 = 1 x ending = face advance loop xx (xbegin,x ending) ilast = x ending - xx + xbegin if xsearching2 = 1 then if model(ilast,y topcoal) = 1 then x fill end = ilast xsearching2 = 2 endif ;end model endif ; end xsearching2 endloop ; loop xx ;; backfill in cutting coal section loop fill y (y fillstart,y roof) loop fill x (xbegin,x fill end) if model(fill x,fill y) = 1 then command model dy i=fill x, j=fill y prop den=2000 bulk=4.5e8 shear=2.1e8 i=fill x j=fill y prop coh=0.1e6 cap p=1e4 cptable 5 fric=25 ten=0.0 & i=fill x j=fill y table 5 delete table 5 0,1e4 0.01,1e5 0.03,3e5 0.05,7e5 0.08,2e6 & 0.1,6e6 0.11,1e7 0.12,2e7 0.13,5e7 endcommand endif ; end model endloop ;fill x endloop ;fill y ;; backfill in topcoal and rockmass sections 264 loop bkfill (1,5000) irow = ixx(bkfill) jcl = jyy(bkfill) if irow # 0 then if irow <= x fill end then if model(irow,jcl) = 1 then command model dy i=irow, j=jcl prop den=2000 bulk=4.5e8 shear=2.1e8 i=irow j=jcl prop coh=0.1e6 cap p=1e4 cptable 5 fric=25 ten=0.0 & i=irow j=jcl table 5 delete table 5 0,1e4 0.01,1e5 0.03,3e5 0.05,7e5 0.08,2e6 & 0.1,6e6 0.11,1e7 0.12,2e7 0.13,5e7 endcommand endif ;model endif ;irow endif ;irow #0 endloop ;bkfill removed rc = 0 ;only if after end loop more search removed cl = 0 searchingzone = 1 endif ;buff eff end if ;removed zn > 0 ;;; searchingzone = searchingzone + 1 ; endloop ;loop searchingzone # 1 command set sratio = 0.005 end command end Appendix B

Model results from FLAC analysis

This appendix contains plots resulting from FLAC analysis presented in Chapter 7

Figure B.1: Contours of vertical stress and the axial forces of liner elements for ﬂat coal seam, pillar width = 5m

265 266

Figure B.2: Contours of vertical stress and the axial forces of liner elements for ﬂat coal seam, pillar width = 10m

Figure B.3: Contours of vertical stress and the axial forces of liner elements for ﬂat coal seam, pillar width = 15m 267

Figure B.4: Contours of vertical stress and the axial forces of liner elements for ﬂat coal seam, pillar width = 20m

Figure B.5: Contours of vertical stress and the axial forces of liner elements for 300 dipping seam, pillar width = 5m 268

Figure B.6: Contours of vertical stress and the axial forces of liner elements for 300 dipping seam, pillar width = 10m

Figure B.7: Contours of vertical stress and the axial forces of liner elements for 300 dipping seam, pillar width = 15m 269

Figure B.8: Contours of vertical stress and the axial forces of liner elements for 300 dipping seam, pillar width = 20m Appendix C

Multiple regression analysis of the CPI method

C.1 Raw results of 3DEC modelling used in the multiple regression analysis

270 271

Table C.1: Raw results from 3DEC modelling

Young’s Seam Seam vertical Horizontal Cover CPI modulus dip thickness spacing spacing depth (%) (MPa) (degree) (m) (m) (m) (m) 1.35 0 1.50 0.5 0.5 500 54.8 1.35 15 1.50 0.5 0.5 500 51.0 1.35 30 1.50 0.5 0.5 500 48.0 1.35 45 1.50 0.5 0.5 500 43.8 1.35 0 1.50 0.75 0.5 500 44.4 1.35 0 1.50 1.0 0.5 500 33.3 1.35 0 1.50 0.375 0.5 500 55.18 1.64 0 1.50 0.5 0.5 500 52.1 1.64 30 1.50 0.5 0.5 500 44.8 1.64 45 1.50 0.5 0.5 500 42.7 1.98 0 1.50 0.5 0.5 500 53.1 2.83 0 1.50 0.5 0.5 500 51.9 3.68 0 1.50 0.5 0.5 500 35.1 1.35 0 1.50 0.5 0.5 400 49.7 1.35 0 1.50 0.5 0.5 300 44.9 1.35 0 1.50 0.5 0.5 200 35.8 1.35 0 1.50 0.5 0.375 500 56.9 1.35 0 1.50 0.5 0.30 500 70.2 1.35 0 1.50 0.5 0.75 500 7.5 1.35 0 1.00 0.5 0.5 500 48.9 1.35 0 2.00 0.5 0.5 500 44.0 1.35 0 1.50 0.5 0.5 500 54.8 1.35 0 0.50 0.5 0.5 500 61.0 1.35 0 2.50 0.5 0.5 500 40.1 272

C.2 Multiple regression analysis

The parameter studies in Chapter 6 has ploted the relationship between the CPI against each parameter in multiple regression. The plot of the CPI against individual parameter help determine an appropriate form of those parameter in regression model. It was assumed that the model of the CPI has an expectation function given as:

β CPI = β0 + β1E + β2α + β3T 7 + β4Sv + β5Sh + β6H (C.1)

Equation C.1 is a nonlinear regression model. According to Ryan (Ryan, 1997) three possible options are commonly suggested:

1. Fit a linear regression model with at least one nonlinear term;

2. Transform CPI and ﬁt a simple linear regression model; or

3. Rearch for a nonlinear regression model that ﬁts the data.

Ryan (Ryan, 1997) states that since nonlinear regresion produces additional complexities, option 1 and 2 are usually prefered. For that reason, option 1, which is to transform the model ﬁt a simple linear regression was chosen in this regression analysis.

In order to transform the nonlinear model to the linear model, the value β7 in Equation C.1 must be estimated. Based on the graph ploting the relation between the CPI and the seam thickness in Figure 6.12, it is assumsed that β7 = -0.2291. Hence, Equation C.1 is redeﬁned as:

−0.2291 CPI = β0 + β1E + β2α + β3T + β4Sv + β5Sh + β6H (C.2)

Give T’ = T−0.2291, Equation C.1 can be rewriten as:

0 CPI = β0 + β1E + β2α + β3T + β4Sv + β5Sh + β6H (C.3)

Equation C.3 is a multiple linear regression where the parameters β1, β2, β3,

β4, β5, β6 need to be determined. The determination of those variables can be carried out by using matrix algebra. 273

    CPI β β E β α β T 0 β S β S β H  1   0 1 1 2 1 3 1 4 v1 5 h1 6 1         0   CPI2   β0 β1E2 β2α2 β3T2 β4Sv2 β5Sh2 β6H2           CPI  =  β β E β α β T 0 β S β S β H   3   0 1 3 2 3 3 3 4 v3 5 h3 6 3           ....   ....         0  CPI22 β0 β1E22 β2α22 β3T22 β4Sv22 β5Sh22 β6H22

Results from multiple regression are shown in Table C.2, C.3, and C.4.

Table C.2: Regression Statistics

Parameter Value Multiple R 0.919699 R Square 0.845847 Adjusted R Square 0.79144 Standard Error 5.43421 Observations 24

Table C.3: ANOVA table

df SS MS F Signiﬁcance F Regression 6 2754.621 459.1035 15.54668 4.64E-06 Residual 17 502.0209 29.53064 Total 23 3256.642 274

Table C.4: Results for multi-regression analysis

Regressor Coeﬃcients Standard Error t-statistic P-value Constant 73.3849 20.7395 3.5384 0.0025 Young’s Modulus (E) -3.0849 2.0939 -1.4733 0.1590 Seam Dip (α) -0.0948 0.0799 -1.1867 0.2517 Seam thickness (T’) 48.2146 18.4663 2.6109 0.0183

Vertical spacing (Sv) -29.7547 9.9254 -2.9978 0.0081

Horizontal spacing (Sh) -132.1712 15.8411 -8.3436 0.0000 Cover depth (H) 0.0336 0.0159 2.1072 0.0503